rotary drilling and blasting in large surface mines .pdf

Rotary Drilling and Blasting in Large Surface Mines

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Rotary Drilling and Blasting in Large Surface Mines

Bhalchandra V. Gokhale

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Cover Illustration:

Atlas Copco Pit Viper 351E drilling 311 mm holes to a depth of 19 m in the Aitik Mine, Sweden, owned by Boliden Mining Co. Courtesy of Atlas Copco, 2009

CRC Press/Balkema is an imprint of the Taylor & Francis Group, an informa business © 2011 Taylor & Francis Group, London, UK Typeset by Vikatan Publishing Solutions (P) Ltd, Chennai, India Printed and bound in Great Britain by Antony Rowe ( A CPI-group Company), Chippenham, Wiltshire All rights reserved. No part of this publication or the information contained herein may be reproduced, stored in a retrieval system, or transmitted in any form or by any means, electronic, mechanical, by photocopying, recording or otherwise, without prior permission in writing from the publisher. Innovations reported here may not be used without the approval of the authors. Although all care is taken to ensure integrity and the quality of this publication and the information herein, no responsibility is assumed by the publishers nor the author for any damage to the property or persons as a result of operation or use of this publication and/or the information contained herein. Published by: CRC Press/Balkema P.O. Box 447, 2300 AK Leiden, The Netherlands e-mail: [email protected] www.crcpress.com – www.taylorandfrancis.co.uk – www.balkema.nl Library of Congress Cataloging-in-Publication Data Rotary drilling and blasting in large surface mines / editor, Bhalchandra V. Gokhale. p. cm. Includes bibliographical references and index. ISBN 978-0-415-87878-4 (hardback) 1. Blasting. 2. Rotary drilling. 3. Strip mining. I. Gokhale, B.V. (Bhalchandra V.) TN279.R68 2011 622’.292–dc22 2010045162 ISBN: 978-0-415-87878-4 (Hbk) ISBN: 978-0-203-84139-6 (eBook)

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Dedication

Dedicated to, My father, who gave me his genes, My mother, who brought me up and showered me with love till my 50s, Both my elder brothers, who supported me after the early demise of my father, My wife, who has been bearing with me for four decades, My daughter, son in law, son and daughter in law who have been making my life worth living, My grand daughter, who has kept me active.

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Contents

Foreword Preface Credits for illustrations 1 Surface rock excavation scenario 1.1 1.2

1.3 1.4 1.5

1.6

Introduction Need for surface rock excavation 1.2.1 Recovery of minerals or crushed rock 1.2.2 Creation of unobstructed space Strength of rock mass and classes Drill and blast operations Options to drill and blast 1.5.1 Water jetting 1.5.2 Machine digging 1.5.3 Ripping 1.5.4 Rock breaking Oilwell drilling

2 An overview of blasthole drilling 2.1 2.2 2.3 2.4 2.5

Introduction Basic concepts of drilling Peculiarities of blasthole drilling Comparison of drilling methods Choice of a blasthole drilling method

3 Properties of rocks 3.1 3.2 3.3

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Introduction Earth and its interior Geological cycle 3.3.1 Igneous rocks

xxv xxvii xxxi 1 1 1 1 3 3 4 7 7 7 8 11 12 15 15 16 22 23 27 31 31 31 34 35

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Contents

3.3.2 3.3.3

3.4

3.5 3.6 3.7

Sedimentary rocks Metamorphic rocks 3.3.3.1 Contact metamorphism 3.3.3.2 Regional metamorphism 3.3.3.3 Dynamic metamorphism 3.3.3.4 Impact metamorphism Formation of rock mass 3.4.1 Volcanic activities 3.4.1.1 Extrusive structures 3.4.1.2 Hypabyssal structures 3.4.2 Plutonic structures 3.4.3 Surface activities Soils, rocks and their profile Occurrence of minerals Rock specimen properties 3.7.1 Density 3.7.2 Compressive strength 3.7.3 Tensile strength 3.7.4 Hardness 3.7.4.1 Shore hardness 3.7.4.2 Vickers hardness 3.7.5 Toughness 3.7.6 Brittleness 3.7.7 Coefficient of internal friction 3.7.8 Swell factor 3.7.9 Abrasiveness 3.7.9.1 Cerchar abrasiveness test 3.7.9.2 Miniature drill test 3.7.9.3 Bit wear test

4 Brief history of rotary blasthole drilling 4.1 4.2 4.3 4.4 4.5

Introduction Era of shot hole drilling Rotary drilling Rotary blasthole drilling Truck mounted rotary blasthole drills

5 Rotary blasthole drilling bits 5.1 5.2

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Introduction Drag bits 5.2.1 Blade bits 5.2.2 Claw bits 5.2.3 Feed force and rotation of drag bits

36 37 38 38 38 38 40 40 40 40 40 43 43 44 46 47 47 50 52 53 54 54 54 55 57 57 57 58 60 61 61 61 63 71 80 83 83 83 83 85 85

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Contents

5.3

5.4

5.5

5.6 5.7

Tricone bits 5.3.1 Types and nomenclature 5.3.2 Geometrical design aspects 5.3.2.1 Fluid circulation 5.3.2.1.1 Air cooled tricone bits 5.3.2.1.2 Sealed bearing bits 5.3.2.2 Size and shapes of teeth 5.3.2.3 Cone size 5.3.2.4 Number of teeth 5.3.2.5 Cone orientation 5.3.2.6 Top connections 5.3.3 Metallurgy 5.3.3.1 Materials 5.3.3.2 Heat treatment 5.3.4 Summary of design features 5.3.5 Bit load and rotation of tricone bits 5.3.6 IADC code Dull bit analysis 5.4.1 Bit observation 5.4.1.1 Observation during drilling operation 5.4.1.2 Observation after bit withdrawal 5.4.1.3 Observation at final bit rejection Bit records 5.5.1 General data 5.5.2 Formation data 5.5.3 Blasthole drill and accessories details 5.5.4 Bit performance details 5.5.5 Bit discard details Bit failure analysis and remedy Discarding drill bits

6 Rotary blasthole drilling accessories 6.1 6.2

6.3

6.4

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Introduction Drill pipes 6.2.1 Integral drill pipes 6.2.2 Fabricated drill pipes 6.2.3 Choice of a drill pipe Stabilizers 6.3.1 Replaceable sleeve stabilizers 6.3.2 Welded blade stabilizers 6.3.3 Integral blade stabilizers 6.3.4 Roller stabilizers Crossover subs

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x

Contents

6.5

6.6

6.7

Shock absorbers 6.5.1 External shock absorber 6.5.2 In the hole shock absorbers Miscellaneous accessories 6.6.1 Thread protectors 6.6.2 Lifting bails and hoisting plugs 6.6.3 Blasthole plugs 6.6.4 Rotary deck bushing 6.6.5 Drill stem wrench 6.6.6 Bit breaker 6.6.7 Blasthole inclinometer 6.6.8 Blasthole camera 6.6.9 Feed force measuring kit 6.6.10 Blasthole dewatering pump 6.6.11 Recovery tools 6.6.12 Blasthole sampler 6.6.13 Laser measuring instruments Miscellaneous safety items 6.7.1 Close fitting shirt and long pants 6.7.2 Safety glasses 6.7.3 Safety toed shoes 6.7.4 Ear plugs or ear muffs 6.7.5 Safety vest 6.7.6 Respirator 6.7.7 Rain coat 6.7.8 Face shield

7 Rotary blasthole drills 7.1 7.2 7.3

7.4 7.5

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Introduction Assemblies in rotary blasthole drills 7.2.1 Assembly placement Types of rotary blasthole drills 7.3.1 Special purpose rotary blasthole drills 7.3.2 Conventional rotary blasthole drills 7.3.2.1 Crawler mounted rotary blasthole drills 7.3.2.2 Carrier mounted rotary blasthole drills 7.3.2.3 Comparison of crawler vs. carrier Layouts of rotary blasthole drills Details of assemblies 7.5.1 Assemblies in the lower group 7.5.1.1 Carrier undercarriage 7.5.1.2 Crawler undercarriage

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Contents

7.5.2

7.5.3

7.5.4

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7.5.1.2.1 Slewing crawler undercarriage 7.5.1.2.2 Pivoted crawler undercarriage 7.5.1.2.3 Crawler track components Assemblies in the upper group 7.5.2.1 Main frame 7.5.2.2 Leveling jacks 7.5.2.3 Prime mover 7.5.2.4 Compressor 7.5.2.5 Hydraulic system 7.5.2.6 Bit lubricating system 7.5.2.7 Automatic lubrication system 7.5.2.8 Radiators and oil coolers 7.5.2.9 Machinery house 7.5.2.10 Dust control system 7.5.2.10.1 Wet dust control 7.5.2.10.2 Dry dust control 7.5.2.11 Operator’s cab 7.5.2.12 Driver’s cab 7.5.2.13 Fuel tanks 7.5.2.14 Transformer and cable reel Assemblies in the mast group 7.5.3.1 Mast 7.5.3.2 Mast raising cylinders and mast braces 7.5.3.3 Mast ladder 7.5.3.4 Pipe changer 7.5.3.4.1 Carousel pipe changers 7.5.3.4.2 Single pipe changers 7.5.3.5 Bit changer 7.5.3.6 Angle hole attachments 7.5.3.7 Rotary head 7.5.3.8 Feed mechanism 7.5.3.8.1 Double drum and wire rope 7.5.3.8.2 Hydraulic cylinder and wire rope 7.5.3.8.3 Hydraulic cylinder and chain 7.5.3.8.4 Fixed sprockets and chain 7.5.3.8.5 Rack and pinion with chain 7.5.3.8.6 Chainless rack and pinion 7.5.3.9 Auxiliary winch 7.5.3.10 Tool handling jib 7.5.3.11 Centralizer Special purpose items 7.5.4.1 Language name plate 7.5.4.2 Lighting

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Contents

7.6

7.7

7.5.4.3 Fast fuel fill system 7.5.4.4 Depth indicator 7.5.4.5 Fast retract system 7.5.4.6 Tow hook 7.5.4.7 Fire extinguishers 7.5.4.8 Automatic fire suppression system 7.5.4.9 Welding outlet 7.5.4.10 Automatic leveling system 7.5.4.11 Remote propel control 7.5.4.12 Air conditioners 7.5.4.13 Heaters 7.5.4.14 Hydraulic test station 7.5.4.15 Video camera system Extreme cold operation devices 7.6.1 Engine starting aid 7.6.1.1 Glow plugs 7.6.1.2 Manifold flame heater 7.6.2 Batteries 7.6.3 Double wall machinery house 7.6.4 Machinery house heater 7.6.5 Operator cab heater 7.6.6 Other enclosures and heaters 7.6.7 Protective coating on steel components Comparison of types of blasthole drills

8 Compressed air and air compressors 8.1 8.2

8.3

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Introduction Compressed air 8.2.1 Gas laws 8.2.1.1 Boyle’s law 8.2.1.2 Charles’ law 8.2.1.3 Guy Lussac’s law 8.2.1.4 Joule’s law 8.2.1.5 Poisson’s law 8.2.1.6 Amagat’s law 8.2.1.7 Avogadro’s law 8.2.1.8 General gas law 8.2.2 Power required for air compression Flow of compressed air 8.3.1 Compressed air flow in steel pipes 8.3.1.1 Simplified formulae 8.3.1.1.1 Harris formula 8.3.1.1.2 Engineering toolbox formula

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Contents

8.3.2 Compressed air flow in hose pipes 8.3.3 Compressed air flow in pipe fittings 8.3.4 Effect of ups and downs in the flow path 8.3.5 Effect of curvature in the flow path 8.3.6 Compressed air flow through nozzles 8.3.7 Leakage of compressed air 8.4 Compressors used on blasthole drills 8.4.1 Sliding vane rotary compressors 8.4.2 Rotary screw compressors 8.4.3 Discharge and pressure control 8.4.3.1 Load/no load control 8.4.3.2 Modulation control 8.4.3.3 Load/no load and modulation control 8.5 Measurements of compressed air 9 Mechanics of rock fracture under a drill bit 9.1 9.2

9.3

9.4

9.5 9.6

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Introduction Rock fracture in drilling 9.2.1 Modes of fragmentation 9.2.2.1 Mechanical energy 9.2.2.1.1 Crushing 9.2.2.1.2 Impact crushing 9.2.2.1.3 Scratching 9.2.2.2 Heat energy 9.2.2.2.1 Spalling 9.2.2.2.2 Fusion and vaporization 9.2.2.3 Chemical energy Basic theory of solid fracture 9.3.1 Mohr-Coulomb theory of shear failure 9.3.2 Griffith theory of tensile failure Fracture of rock in drilling 9.4.1 Penetration of an indenter 9.4.2 Penetration of a wedge 9.4.3 Formation fracture below a drill bit 9.4.4 Specific energy Drillability of rocks Estimation of penetration rate 9.6.1 Protodyakonov approach 9.6.2 Paone and Bruce approach 9.6.3 Bauer, Calder and Workman approach 9.6.4 Specific energy approach 9.6.5 Microbit test approach 9.6.6 Indenter test approach

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Contents

10 Flushing the blasthole 10.1 10.2 10.3

10.4

10.5 10.6

10.7 10.8 10.9

Introduction Why compressed air? Schematics of flushing 10.3.1 Gravitational force 10.3.2 Buoyant force 10.3.3 Drag force 10.3.4 Terminal velocity 10.3.5 Laws for settlement of particles in fluid Formulation of desired bailing velocity 10.4.1 Size of fragments 10.4.2 Density of fragments 10.4.3 Roundness of fragments 10.4.4 Roughness of fragments 10.4.5 Rate of fragmentation 10.4.6 Annular space 10.4.7 Inclination of the blasthole 10.4.8 Quantity of water injection 10.4.9 Other approaches Formulation of compressed air pressure 10.5.1 Oil injection Choosing a drill pipe 10.6.1 Drill pipe dimensional parameters 10.6.2 Drill pipe surface treatment 10.6.3 Size and shape of drill cuttings 10.6.4 Bailing velocity 10.6.5 Drill pipe wall thickness Discarding a drill pipe Choosing nozzles for tricone bits Choosing the right compressor

11 Effect of altitude and severe weather 11.1 11.2

11.3

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Introduction Atmosphere 11.2.1 Effects of altitude on atmosphere 11.2.2 Effects of humidity on atmosphere Effect of scarce air in drilling 11.3.1 Diesel engines 11.3.2 Electric motors and transformers 11.3.3 Hydraulic system 11.3.4 Air compressors 11.3.5 Blasthole flushing 11.3.6 Drill lubrication

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Contents

11.4

Effect of severe weather 11.4.1 Stormy winds 11.4.2 Heavy rainfall 11.4.3 Heavy snowfall 11.4.4 Extreme hot weather 11.4.5 Extreme cold weather

12 Computers in rotary blasthole drilling 12.1 12.2

12.3 12.4

12.5

12.6

12.7

12.8

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Introduction Drill operation 12.2.1 Tramming 12.2.2 Leveling 12.2.3 Drilling and pipe handling Modes of blasthole drill computerization Drilling knowledge systems 12.4.1 Formation logging units 12.4.2 Analog drilling recorders 12.4.3 Electronic drilling recorders 12.4.4 Parameters of drilling knowledge data Positioning by GPS 12.5.1 Location recognition by GPS 12.5.2 Movement of drill by using GPS Computerized drill systems 12.6.1 Measurements while drilling 12.6.1.1 Independent parameters 12.6.1.1.1 Time 12.6.1.1.2 Depth 12.6.1.1.3 Rotary speed 12.6.1.1.4 Weight on the bit 12.6.1.1.5 Torque 12.6.1.1.6 Vibrations 12.6.1.1.7 Compressed air pressure 12.6.1.2 Calculated parameters 12.6.1.2.1 Rate of penetration 12.6.1.2.2 Blastability index 12.6.2 Automation systems Hardware for drill computerization 12.7.1 System unit 12.7.2 Display monitor 12.7.3 Input device 12.7.4 Software Advantages of drill computerization 12.8.1 Advantages of drilling knowledge system

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Contents

12.8.2 12.8.3 12.8.4 12.8.5

Advantages of drill automation system Advantages of GPS positioning system Advantages of integrated mining system Advantage of added safety 12.8.5.1 Blasthole depth limiting input 12.8.5.2 Tram interlock for pipe in hole 12.8.5.3 Centralizer damage prevention 12.8.5.4 Leveling jack interlock 12.8.5.5 Carousel damage prevention 12.8.5.6 Breakout wrench damage prevention

13 Concepts for rotary blasthole drill design 13.1 13.2 13.3 13.4 13.5 13.6 13.7 13.8 13.9 13.10 13.11

375

Introduction Primary requirements of the drill Size and weight of the drill Rotary speed and torque Feed force and speed Pullout force and speed Breakout torque Hydraulic leveling jacks Ground load bearing Compressor discharge and pressure Engine power

375 375 376 378 383 384 384 385 385 386 386

14 Cost analysis of rotary blasthole drilling

389

14.1 14.2 14.3 14.4

14.5

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369 369 371 372 372 372 372 373 373 373

Introduction Working pattern of a blasthole drill Factors related to blasthole drilling cost Owning cost 14.4.1 Purchase expenditure 14.4.2 Yearly taxes, duties and levies 14.4.3 Salvage value 14.4.4 Calculation of owning cost 14.4.4.1 Single installment 14.4.4.2 Multiple installments Operating cost 14.5.1 Cost of maintenance and repairs 14.5.2 Cost of consumables 14.5.2.1 Cost of power 14.5.2.2 Drill lubricants 14.5.2.3 Oils 14.5.2.4 Water

389 390 393 393 393 396 396 397 398 398 399 399 400 401 401 403 403

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Contents

14.6 14.7 14.8 14.9

14.5.2.5 Tires 14.5.2.6 Power cable 14.5.3 Cost of accessories and bits 14.5.3.1 Drilling accessories 14.5.3.2 Tricone bits 14.5.4 Cost of operating labor Overhead cost Total blasthole drilling cost General considerations in cost estimation Example of cost estimation

15 Procurement of rotary blasthole drills 15.1 15.2

15.3

Introduction Pre procurement considerations 15.2.1 Which drilling method? 15.2.2 What type of drill? 15.2.3 How many blasthole drills? Procurement methodology 15.3.1 Draft of specification for the drill 15.3.2 Erection and commissioning of the drill 15.3.3 Drill manuals 15.3.4 Operation and maintenance training

16 Tips for operating rotary blasthole drills 16.1 16.2 16.3 16.4 16.5 16.6 16.7 16.8 16.9 16.10 16.11

Introduction Knowing the environment Moving the rotary blasthole drill Setting up the rotary blasthole drill Operating a rotary blasthole drill Controlling dust emissions Transporting the drill Maintenance activities Periodical check ups Drilling records Safety precautions

17 Mechanics of blast fragmentation 17.1 17.2 17.3 17.4

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Introduction Explosion process Formation of a crater by a blast Blast fragmentation mechanism

xvii

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Contents

17.4.1 17.4.2 17.4.3 17.4.4 17.4.5 17.4.6 17.4.7 17.4.8

Crushing of rock Radial cracking Circumferential cracking Spalling Radial push by gases Cracking by flexure Collision of fragments Separation of beds

18 Properties of explosives 18.1 18.2 18.3

Introduction Explosives Properties of explosives 18.3.1 Velocity of detonation 18.3.2 Detonation pressure 18.3.3 Blasthole pressure 18.3.4 Strength 18.3.4.1 Weight and bulk strength 18.3.4.2 Absolute and relative strength 18.3.5 Heat of explosion 18.3.6 Specific gas volume 18.3.7 Sensitivity 18.3.7.1 Sensitivity to shocks 18.3.7.2 Sensitivity to friction 18.3.7.3 Sensitivity to heat 18.3.7.4 Sensitivity to detonator strength 18.3.7.5 Gap sensitivity 18.3.8 Handling and transport safety 18.3.9 Brisance value 18.3.10 Charging density 18.3.11 Toxic fumes 18.3.12 Water resistance 18.3.13 Hygroscopicity 18.3.14 Storage life 18.3.15 Volatility 18.3.16 Material coexistence 18.3.17 Minimum hole diameter

19 Thermochemistry of explosives 19.1 19.2 19.3

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Introduction Chemical nature of explosives Reactions of explosive chemicals

440 440 441 442 442 442 444 444 445 445 445 449 449 451 452 452 452 453 453 453 453 454 454 454 454 455 455 455 455 456 458 459 459 460 460 460 463 463 463 463

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Contents

19.3.1 19.3.2 19.3.3 19.3.4

Oxygen balance Volume of products of explosion Heat of explosion Strength of explosive

20 Explosives – their history and composition 20.1 20.2 20.3

20.4

Introduction Brief history of explosives Types of explosives in a blasthole 20.3.1 Detonator (blasting cap) 20.3.2 Primary explosive 20.3.3 Booster 20.3.4 Main explosive 20.3.4.1 Nitroglycerin 20.3.4.2 Nitrocellulose 20.3.4.3 Ammonium nitrate 20.3.4.4 Sodium nitrate 20.3.4.5 Fuel oil 20.3.4.6 Wood pulp 20.3.4.7 Sulfur 20.3.4.8 Antacid Explosive mixes used in mine blasts 20.4.1 Dry blasting agents 20.4.1.1 Poured ANFO 20.4.1.2 Packaged ANFO 20.4.1.3 Heavy ANFO 20.4.2 Slurry 20.4.2.1 Bulk slurry 20.4.2.2 Cartridged slurry 20.4.3 Emulsions 20.4.4 Dynamite 20.4.4.1 Granular dynamite 20.4.4.2 Gelatin dynamite

21 Tests on explosives 21.1 21.2

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Introduction Tests for measurements of VOD 21.2.1 D’Autriche method 21.2.2 Chronograph method 21.2.3 Fiber optic sensor method 21.2.4 SLIFER method 21.2.5 Optical measurement method

xix

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Contents

21.3

21.4

21.5

21.6

Tests for measurement of strength 21.3.1 Ballistic mortar test 21.3.2 Cratering test 21.3.3 Underwater test 21.3.4 Plate dent test 21.3.5 Traulz lead block test 21.3.6 Cylinder compression test Tests for measurement of sensitivity 21.4.1 Shock sensitivity test 21.4.2 Heat resistance test 21.4.3 Cap sensitivity test 21.4.4 Gap sensitivity test 21.4.5 Friction sensitivity test 21.4.6 Electrostatic discharge sensitivity test Tests related to storage of explosive 21.5.1 Effect of wetness test 21.5.2 Internal ignition test 21.5.3 Material compatibility test 21.5.4 Vacuum stability test Miscellaneous tests 21.6.1 Critical diameter test 21.6.2 Critical height test 21.6.3 Bullet impact test 21.6.4 Koenen test

22 Blasting consumable and accessories 22.1 22.2

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Introduction Blasting consumables 22.2.1 Initiation transmission line 22.2.1.1 Safety fuse for transmitting fire 22.2.1.2 Electric transmission wires 22.2.1.3 Detonating cord for transmitting detonation 22.2.1.4 Detonating cord for transmitting shock 22.2.1.5 Hercudet tube for transmitting fire 22.2.2 Detonator 22.2.2.1 Delay element 22.2.2.2 Electric detonators 22.2.2.2.1 Solid pack electric detonator 22.2.2.2.2 Fusehead electric detonator 22.2.2.2.3 Exploding bridgewire detonator 22.2.2.3 Non-electric detonators 22.2.2.4 Electronic detonators

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Contents

22.3

22.4

Blasting instruments 22.3.1 Testing instruments 22.3.1.1 Blaster’s multimeter 22.3.1.2 Blaster’s ohmmeter 22.3.1.3 Blaster’s tagger 22.3.2 Initiating instruments 22.3.2.1 Safety fuse initiator 22.3.2.2 Electric detonation initiator 22.3.2.3 Generator type blast initiator 22.3.2.4 Capacitor discharge type blast initiator 22.3.2.5 Detonation wave initiator 22.3.2.6 Shock wave initiator 22.3.2.7 Hercudet system initiator 22.3.2.8 Electronic blast initiator 22.3.3 Measuring instruments 22.3.3.1 Hot hole meter 22.3.3.2 Burden measuring instrument 22.3.3.3 Seismograph 22.3.3.4 VOD meter 22.3.3.5 High speed camera Blasting tools and miscellaneous items

23 Hazards to and from blasting 23.1 23.2

23.3

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Introduction Hazards to blasting process 23.2.1 Presence of ground structure 23.2.2 Seepage of ground water 23.2.3 Hot blastholes 23.2.4 Hole deviation Hazards of blasting process 23.3.1 Misfires 23.3.1.1 Faulty safety fuse installation 23.3.1.2 Faulty electric blasting circuits 23.3.1.3 Faulty detonating cord circuits 23.3.1.4 Faults of exploder or faulty operation 23.3.1.5 Unnoticed ground water inflow 23.3.2 Ground vibrations 23.3.2.1 Nature of ground vibration waves 23.3.2.2 Prediction of ground vibration levels 23.3.2.2.1 Langefors formula 23.3.2.2.2 Scaled distance formula 23.3.2.3 Damage by ground vibrations

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Contents

23.3.3

23.3.4 23.3.5 23.3.6

Air blast 23.3.3.1 Prediction of air blast pressure 23.3.3.2 Damage by air blast pressure Fly rocks Air pollution Environmental changes

24 Properties of rock masses 24.1 24.2

24.3

24.4

Introduction Rock specimen properties and blasting 24.2.1 Influence of rock strength on blasting 24.2.2 Influence of rock density on blasting 24.2.3 Influence of rock porosity on blasting 24.2.4 Specimen blastability Properties of rock masses and blasting 24.3.1 Voids 24.3.2 Folds, unconformities and bedding planes 24.3.2.1 Shooting with the dip 24.3.2.2 Shooting against the dip 24.3.2.3 Shooting along the strike 24.3.3 Faults and joints Classification of rock masses 24.4.1 Classification by visual observation 24.4.2 Classification by index 24.4.2.1 RQD based classification 24.4.2.2 RMR based classification 24.4.2.3 RTQI based classification 24.4.3 BI index based classification 24.4.3.1 Index proposed by Hansen (1968) 24.4.3.2 Index proposed by Hainen and Dimock (1976) 24.4.3.3 Index proposed by Ashby (1977) 24.4.3.4 Index proposed by Langefors (1978) 24.4.3.5 Index proposed by Lilly (1986) 24.4.3.6 Index proposed by Ghose (1988) 24.4.3.7 Index proposed by Gupta (1990) 24.4.3.8 Index proposed by JKMRC (1996) 24.4.3.9 Index proposed by Han, Weiya and Shouvi (2000)

25 Methods and patterns of charging blastholes 25.1 25.2

Book.indb xxii

Introduction Mechanized blasthole charging system

545 546 547 549 550 552 553 553 553 554 554 555 555 555 556 556 558 558 559 560 560 560 560 562 563 564 564 565 566 566 566 566 568 568 569 569 571 571 571

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Contents

25.3

25.4

25.2.1 Trucks for loading ANFO or heavy ANFO 25.2.2 Trucks for loading slurry or emulsion 25.2.3 Safety features of bulk delivery system Blasthole charging pattern 25.3.1 Type and placement of primer 25.3.2 Direction of propagation of detonation 25.3.2.1 Top priming 25.3.2.2 Bottom priming 25.3.2.3 Multi point priming 25.3.3 Continuous side initiation 25.3.4 Air decking 25.3.5 Priming under special rock mass conditions 25.3.5.1 Well defined ore layer 25.3.5.2 Heavy water seepage in blasthole 25.3.5.3 Hard boulder in soft bed 25.3.5.4 Cavities in the rock mass Drilling and firing patterns 25.4.1 Drilling patterns 25.4.2 Firing patterns 25.4.2.1 “V” pattern 25.4.2.2 Echelon delay pattern 25.4.2.3 Flat face pattern 25.4.2.4 Channel delay pattern 25.4.2.5 Sinking hole pattern

26 Design of a surface blast 26.1 26.2 26.3

26.4

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Introduction Blastholes in a mine bench Types of blasting in surface mines 26.3.1 Conventional bench blasting 26.3.2 Secondary blasting 26.3.3 Cast blasting 26.3.4 Presplitting 26.3.5 Snake hole blasting 26.3.6 Rip rap blasting What is involved in design of a blast 26.4.1 Powder factor 26.4.2 Blasting direction 26.4.3 Number of free faces 26.4.4 Blasthole diameter 26.4.4.1 Desired rate of production 26.4.4.2 Desired fragmentation

xxiii

572 573 574 575 576 580 581 582 582 583 583 586 586 586 587 587 588 589 592 592 593 593 593 595 597 597 597 598 598 599 599 599 601 603 603 603 604 604 605 605 606

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xxiv

Contents

26.5 26.6 26.7 26.8

26.4.4.3 Properties of rock specimen 26.4.4.4 Presence of geological structures 26.4.5 Blasthole depth 26.4.5.1 Combination of shovel and dumper 26.4.5.2 Walking dragline 26.4.5.3 Combination of wheel loader and dumper 26.4.6 Blasthole inclination 26.4.7 Burden 26.4.8 Blasthole spacing 26.4.9 Subdrilling 26.4.10 Stemming height 26.4.11 Size of the blast Calculation of burden Relationship of blasthole parameters Design of conventional surface blast Design of other types of blasts 26.8.1 Rip rap production blasting 26.8.2 Snake hole blasting 26.8.3 Cast blasting

Appendix 1 Appendix 2 Appendix 3 Appendix 4 Appendix 5 Appendix 6 Appendix 7 Appendix 8 Appendix 9 Appendix 10 Appendix 11 Appendix 12 Appendix 13 Appendix 14 Appendix 15 Index

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Properties of atmospheric air at high altitudes Air properties at various temperatures Bailing velocities in blasthole annulus Air pressure loss in steel pipes Air pressure loss in hose pipes Air pressure loss in pipe fittings A note on tungsten carbide and other hard metals Scales of hardness – measurements and their conversion Manufacturers of rotary blasthole drilling equipment Details of rotary blasthole drills manufactured all over the world Details of drill pipes used for rotary blasthole drilling Details of stabilizers and other miscellaneous items Details of shock absorbers Properties of some rocks Bibliography

606 606 609 609 610 611 612 616 617 618 620 620 621 625 627 630 630 630 632 639 643 647 651 657 661 667 675 679 687 693 699 707 709 717 737

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Foreword

I joined Bucyrus-Erie in 1974 after finishing my graduate studies. BE assigned me working on the design and test of rotary blast hole drills in the U.S., Canada and South Africa. In 1976 we had offered BE 45R (and also 40R which was soon to be discontinued) rotary blasthole drill to Malanjkhand Copper Project in India. To follow up this potential business I, as a young Sales Manager of BE, visited the mining site near Balaghat, with another young sales executive Mr. B.V. Gokhale from our Indian distributor. I vividly remember, Mr. Taneja, the director of the project, stating in our last meeting, “Mr. Johnson, we can’t purchase 45R because it is too large and too expensive for our job and your 40R does not have angle hole drilling attachment which is so important for our job. However, I must say that you and Mr. Gokhale have certainly taught us many aspects of rotary blasthole drilling in our discussions. Many thanks for that.” BV and I traveled throughout India to sites operated by Hindustan Copper, KIOCL, and Coal India at a later date. We had discussion on different types of equipment. I found BV not only had a superior knowledge of blast hole drills but also rigid frame haul trucks, dragline, electric mining shovels etc. When BE were considering starting manufacture of their shovels in India, we found BV as the most suitable individual to work as consultant to BE. Based on some of his work we were able to form a joint venture for electric mining shovels with Heavy Engineering Corporation and had signed a MOU for blast hole drill manufacturing. BV then conducted an in depth study of the Indian Coal sectors and equipment demands for the next 20 years. His study involved visit to all sites in the country for coal mining and accessing the precise long term demand for coal for power generation. His study convinced BE to establish a subsidiary operation in India. I was convinced beyond doubt that knowledge of BV in the domain of blasthole drilling was of highest caliber. Sometime during 1984 he mentioned to me that he is going to write a book on blasthole drilling. I knew he had all the necessary talent and knowledge for the venture. I wished him best luck and was really eager to see the book because there was no book on the subject of blasthole drilling. The nearest to the subject was “Drilling of Rocks” by McGregor but was grossly outdated within a decade after its publication in 1967.

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xxvi

Foreword

I was naturally pleased to see his book “Blasthole Drilling Technology” published in 2003. It does contain lot of useful material on the subject. But as the book was published by him, without backing of an international publisher, it has remained somewhat unknown. Now he has come up with this “Rotary Drilling and Blasting in Large Surface Mines”. I have read a few chapters of the book and I find that BV has put many innovative ideas on various topics. I can whole heartedly recommend this book to everyone directly or indirectly involved in blasthole drilling. It covers every aspect of rotary blasthole drilling. To a mining engineering student it gives complete knowledge of the machine and methods of its use without spending hours of field work. For a driller it will give deep insight into the various factors that affect the drilling process and several associated topics. To a mining superintendent it shows how to choose machine drill pipes and other high value consumables. For him the elaboration of cost calculations is of great importance. To a design engineer with a drill manufacturer it reveals several concepts that form the basis of the actual drilling process, and guiding principles for the design of a rotary drill. In short, by reading this book everyone involved with rotary blasthole drilling, will get a deep insight in the subject and will become acquainted with many ideas that can be used to get better results in drilling operations. Later in the book BV has also covered blasting techniques. Whereas I don’t have adequate knowledge to comment on the blasting part, after quickly reading it, I certainly feel that he has described the explosives, the process of rock fracture in blasting, thermochemistry of explosive and the design of a blast for surface mining operations in most appropriate manner. With a worldwide sales network CRC Press/Balkema are well placed to spread the knowledge of relatively untouched subject of rotary blasthole drilling and blasting through this book. I wish grand success to both BV and CRC Press/Balkema. Dean Johnson, Ex VP Terex Corporation. Currently Capital Goods Consultant for Global Sales and Distribution of Open Pit Mining and Construction Equipment.

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Preface

Nearly seven years ago I wrote and published my book titled ‘Blasthole Drilling Technology’. Many readers of the book, found it very useful because it contained more detailed information about rotary and rotary percussion drilling than almost all other books. Some of the readers, however, did express that I should have written two separate books – one on rotary blasthole drilling and the other on rotary percussion blasthole drilling. When I thought over this suggestion, I did realize that these two modes of blasthole drilling are vastly different from each other in terms of equipment and techniques. The only commonality is that both can be effectively used for the blasthole drilling application. Even for such application rotary percussion drilling is best suited for drilling small diameter blastholes, whereas, rotary drilling is best suited for drilling larger blastholes. Therefore, initially I decided to write a book only on rotary blasthole drilling. As I thought further over the matter I realized that in ‘Blasthole Drilling Technology’ blasting was covered only through one chapter and hence gave only rudimentary knowledge of it to the readers. In other words the book was subject-oriented and not application-oriented. In reality blasthole drilling and blasting are so inseparable that every book on blasthole drilling must cover blasting to a considerable extent. Considering all these aspects I have written this book. The book is named ‘Rotary Drilling and Blasting in Large Surface Mines’ because rotary blasthole drilling is almost exclusively carried out in large surface mines. The book also covers blasting to sufficient depth. Quite a bit of text matter as well as illustrations in this book have been repeated from ‘Blasthole Drilling Technology’ but there have been many additions to the text matter and illustrations. This surely makes it worthwhile to purchase this book even if one has a copy of ‘Blasthole Drilling Technology’. This book is surely not a work of art but a book pertaining to specific techniques. The purpose of the book is to make the readers understand the techniques of drilling large diameter blastholes by rotary drilling method, subsequently filling them with appropriate explosives in apt manner and finally blasting them to get desired fragmentation of the rock mass. When it comes to understanding, a picture is always worth thousands of words. Keeping this saying my in mind I have included nearly 400 illustrations in this book. Almost 60% of these illustrations are based on the images that I have retrieved from internet websites. As it was necessary to change these images to suit the context.

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xxviii

Preface

I have modified almost all the images. I have endeavored my best efforts in obtaining permission for us of the original image. In this regard I must honestly confess that without use of such images this book would not have been complete. On pages immediately following the table of contents of this book in a section titled ‘Credits for Illustrations’ I have given due credit to the organizations or individuals for initial use of the mother image. In case some name has been missed please consider such errors to be inadvertent. Readers are sincerely requested to bring all such error to my notice through a letter, so appropriate action will be taken at the time of preparing the next edition of this book. During the early years of my interest many manufacturers and suppliers fed me with their product and technical literature to turn me towards the field of drilling. I am thankful to them. Some names that I vividly remember in this regard are, Tamrock, Atlas Copco, Ingersoll Rand, Joy Manufacturing Co., Gardner Denver. I worked in Greaves Cotton Co. Ltd., Voltas Ltd., Bucyrus Erie Co. and Harnischfeger Corporation. As a product manager my job was to understand technological aspects of different types of construction and mining equipment and answer technical queries of customers. Without the generous attitude of organizations mentioned in earlier paragraphs this book would not have seen the light of day. From late 1990s Internet became the source of my information. It is ever expanding. There are many individuals, without whom this book would not have been possible. In this regard the first name that comes to my mind is Ms. Germaine Seijger, Senior Acquisition Editor of CRC Press/Balkema. Apart from her there are many individuals from CRC Press/Balkema who have taken lot of efforts in greatly elevating the standard of this book. Next in the line is Mr. Dean Johnson, Ex VP Terex Corporation. Right from the days in 1977 when our friendship began, till now he has been sending me very valuable articles, books etc. on drilling so as to be well informed in the subject. Perhaps more important is the fact that he has always kept my enthusiasm burning, I have personally not met him since 1986 but on his own he has gifted me two most recently published books on blasthole drilling. I cannot forget the names of Prof. A.K. Ghose, Former Director of ISM Dhanbad, Dr. A.K. Chakraborty and Mr. G.K. Pradhan, Joint Director of PCRA. After reading my book ‘Blasthole Drilling Technology’ these respected persons were largely instrumental in spontaneously awarding me with a Lifetime Achievement Award for Outstanding Contribution in Drilling Engineering through Indian Mining Journal at RockEx 2005 held in Nagpur. This was a great moral booster for me. It kept my interest in the field alive to culminate in this book. A technical book is always backed by technical data. While preparing the table of appendix 10 many persons gave the needed information promptly. I am thankful to them. Ms. Amanda Finn – Measurement Devices Ltd., Mr. Brian Fox – Atlas Copco Drilling Solutions, Mr. K.S. Ramchandran – Revathi Equipment Ltd., Mr. Tharmarajah Ramchandran – Sandvik Mining and Construction, Mr. Greg Scott – Bucyrus International Inc., Mr. Fred Slack – Schramm Inc.

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Preface

xxix

I believe, that every technical book written by any author must be read by at least one external editor who can make the book easily “digestible” to the readers. For my book this was very ably done by Dr. James Rainbird. He not only polished my language but with his superb memory and “hawk” eye went on to point out a few anomalies as distantly placed as chapters 3 and 22. I am greatly indebted to him. Finally, I thank my wife Suneeti, son Amit, daughter in law Anjali for smilingly bearing the neglect shown towards them while writing the book. B.V. Gokhale Mumbai, August 8, 2010

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Credits for illustrations

Literally hundreds of illustrations have been included in this book so as make the contents easily understandable to the readers. Nearly 60% of the illustrations are based on images gathered by me from various sources viz. books, periodicals and particularly internet websites of many manufacturers and suppliers. I have made changes in the images to suit the contents in the book. Thanks are due to all these individuals or organizations for such indirect contribution. Following is a list of individuals or organizations who originally have published the mother images.

Name of individual or organization

Figures for which credit given

AB Nordbok Acker Drill Co. Allis Chalmers American Coldset Corporation Aquila Mining System Atlas Copco

4.1 5.1 8.11-8.14 6.8 12.5, 12.6, 12.9, 12.10, 12.11, 12.13 2.2-2.5, 2.7, 4.10, 5.5, 5.6, 5.8, 6.5, 6.9, 7.1, 7.4, 7.6, 7.28, 7.31, 7.34-7.36, 7.41, 7.43, 7.57, 7.58, 8.5, 10.4, 10.5, A12.1, Table 5.13 7.23 4.12, 4.13 22.11, 22.20 22.20 5.2, 5.3 2.8, 4.8, 4.18, 4.19, 7.2, 7.8, 7.9, 7.22, 7.25, 7.38-7.40, 7.42, 7.50, 7.59, 6.2, 6.4, 8.3, 13.1, 26.3, 26.7 1.9, 7.10, 7.15 3.10 7.51 12.3, 26.14 26.27-26.30 4.14, 4.16 2.6, 6.1, 6.6 6.14, 6.15 22.7, 22.12, 22.14, 22.15 3.7

Autolub Lubin Baroid Inc. Blasting Accessories Inc. Blasters Tool and Supply Co. Inc. Blue Demon Bucyrus International Inc. Caterpillar Christine Rasch+Dr. Wolfgang Kollenberg Christopher Bills CR Books D’Appolonia Dave Lang Drillco Smithtool Duraquest Dyno Nobel ELE International

(Continued )

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xxxii




Elvex Corporation Explosiva Fluid Control Products Inc. Geolograph GeoSonic GeoVision Gulf Publishing Hamlyn Publications implosionworld.com International Drilling Services J. Trent Bruce Joy Manufacturing Co. Keith Haddock Kennmetal Laser Measuring Instruments Inc Laval Measurement Surveys LIM SA Lincoln M&P Flange and Pipe Protection Mammoth Technologies Measurement Devices Ltd. Mining Technology International Modular Mining System Nordex Explosives Ltd NTNU – Norwegian Univ. of Sci. and Tech. Oil Tool International Services Pvt Ltd Olympus Orica P&H Mining Prof Hustrulid Pulsar Measuring Co. R. Smith and Sons Inc. Research Energy of Ohio Inc. Sandvik Schlumberger Schramm Inc. SDS Corporation Security BDS Smith Drilling Co. SA Smithtool Steve (Hanna Equipment Co.) Thomas Instrument Inc. ThyssenKrupp Fordertechnik TLC Engineering Solutions + Vibronics Transtronics Trumph Drilling Tools Twin Disk US Army US Dept. of Interior+US BM+OSHA Ultra Enviro-Systems (P) Ltd. United Nations Environmental Program

6.38 25.3 7.52 12.2 22.17 6.26 4.2-4.5, 4.7, 4.11 3.1, 3.2, 3.5 22.13 A12.2, A12.5 8.15 4.15, 7.5, 7.7 4.6, 4.20 5.4 6.33 6.27 6.23, 12.4 7.24 8.17 6.21 6.31, 6.34-6.37 6.12, 6.13 12.8 25.4 3.11, 3.14, 17.10 A12.8 22.19 25.1 7.26, 7.30, 7.32, 7.37, 7.47, 7.56, 8.6, 26.10, 26.11 25.2 6.25 6.18 22.11 4.21, 7.16, 7.48, 23.7, 26.6, A7.1 6.22 7.13, 7.14, 7.27, 7.49 6.32 5.9, 6.28, 8.18 6.16 5.7, 5.11, 5.12, 5.14, 5.17, 5.18, 5.29, 5.21 1.8 22.11 1.7 21.2 6.24 A12.7 7.21 25.20 7.54, 9.16, 9.24, 20.1, 20.2, 22.2, 22.18 22.18 1.12 (Continued )

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Book.indb xxxiii



University of Western Australia Untraced Uralmash Vermeer Corporation Vibronics Inc. World Mining Equipment www.all-security-system.com www.geology.ebr.state.nc.us www.soundwaves.usgs.gov

26.14 23.4, 23.9, 23.10, 25.18 5.10 1.10 21.2 7.20 7.53 1.11 1.6

xxxiii

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Chapter 1

Surface rock excavation scenario

1.1

INTRODUCTION

Excavation of rock at the surface of the ground becomes necessary for several reasons. Drill and blast is the most commonly adopted method for such excavations. In essence, this consists of drilling an array of holes in the rock mass, filling them with explosives and then initiating an explosion so that the release of enormous energy breaks the large rock mass into small pieces. This may appear quite simple, but the techniques of drilling the holes and blasting them have advanced very greatly through the experience gained in the past four centuries over which this excavation method has been used. Excavation of rock below the surface of the ground is also essential for various purposes. The techniques, methods and equipment used for underground excavation differ considerably from those used for surface rock excavation. Apart from drill and blast, other alternative techniques and corresponding equipment are used for surface rock excavation. This chapter is devoted to essential background knowledge of surface rock excavation techniques.

1.2

NEED FOR SURFACE ROCK EXCAVATION

Surface rock excavation can mean an excavation carried out by means of equipment that stands on the surface of the ground. However, in this book its meaning is taken as the excavation of the rock lying below the surface of ground at shallow depth. The need for surface rock excavation is felt for three basic reasons as follows. 1 2

Recovery of minerals or crushed rock Creating an unobstructed space These points are elaborated further below.

1.2.1

Recovery of minerals or crushed rock

Many minerals of commercial importance, called ores, are located at shallow depths below the ground surface. Mineral deposit is the term used for a rock mass that

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2

Rotary drilling and blasting in large surface mines

contains a mineral of commercial importance in economically large proportions. A mineral deposit is usually surrounded by other minerals that have no economic importance. These are called waste. Since they often lie above the mineral deposits, as in surface coal mines, they are also called overburden. When mineral deposits lie near the surface of the ground, as illustrated in Figure 1.1, the volume of waste to be removed is comparatively less. In such cases, the whole rock mass is fragmented by one of the rock excavation methods. It is very easy to visually differentiate and pick up ore from such a fragmented rock mass with loading equipment and to transport it by hauling equipment to the place designated for further processing of the ore. To ensure continuity of operation of ore recovery, it also becomes necessary to transport the waste to a dumping place. This is sometimes carried out by using different types of equipment. All the activities narrated in the above paragraphs are together called surface mining, or open pit mining, or opencast mining. Large pieces of rocks are used in the construction of dams, breakwaters and the substrate of road and railway embankments. Smaller crushed rock pieces of different sizes are used in large quantities for making concrete used for the construction of objects such as dams, bridges, harbor wharves, buildings etc. For making very strong and durable structures, the concrete must have great strength. This is possible only when the pieces of rock used in making concrete have a high compressive strength. Therefore, crushed rock is recovered from rock masses in specific areas. The activities carried out for getting crushed rock are similar to those adopted in surface mining. Since no specific mineral is involved and the whole rock mass is of importance, all these activities together are called quarrying rather than mining.

Overburden

Ore Bodies

Figure 1.1 Ore bodies below a hillock.

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1.2.2

3

Creation of unobstructed space

Instances in which it is necessary to make suitable space, by removing obstructions and unevenness of the ground, are met in virtually every type of construction. Some common examples are described below. Huge skyscrapers constructed in many cities have to stand and remain stable on the ground for over a hundred years. A sufficiently competent rock mass, capable of bearing the enormous load imposed by such buildings, is met only at depths in the ground. This makes it necessary to excavate a large rock mass for construction of the foundations. Foundations of bridges have to reach to great depths and have to penetrate into competent rock to a certain depth, because only at such depth is the load-bearing capacity of the ground sufficiently high. The danger of the whole foundation drifting under heavy water currents is also minimized when foundations are constructed in this manner. Roads and railways have to pass through hilly regions at several places. Since locomotives and vehicles have limited grade ability, at many places it is safer and economical to make a long cutting or to construct a tunnel along the road or railway route rather than make a long circuitous route. Depending upon the topography and the width of the road or number of railway tracks, the cutting can be of enormous size. Cuttings of very large dimensions are required to be excavated for water supply or navigation canals. Concrete dams have to be anchored well in the rock mass beneath the ground surface in the river basin as well as at both sides. For properly anchoring a large concrete dam, excavation can spread over a very large area and to great depths. In all the examples mentioned above, surface excavation of the rock has to be carried out by one of the excavation methods. Often a rock mass measuring several million cubic meters is removed in the excavation process.

1.3

STRENGTH OF ROCK MASS AND CLASSES

The word ‘rock’ will be repeatedly used in forthcoming chapters. The meaning of this word is viewed differently by geologists and engineers. To a geologist, every aggregate of minerals means rock, because he is more concerned with the formation, stratification, mineral contents etc. of the rock, rather than its strength. A civil engineer looks at the rock from the viewpoint of strength, load-bearing capacity etc., as he is more concerned with either making something by using rock pieces or breaking a rock mass to create space and construct a structure at the place. A mining engineer is concerned with both the above viewpoints. The properties of rocks from an engineering viewpoint have been elaborated in chapter 3 because they are more closely connected to drilling. The properties of rock masses assume larger importance in the realm of blasting and are, therefore, detailed in chapter 23. The properties of rock masses are however, for the better understanding of various aspects, presented in the first and second chapters, so the concept of the strength of rock must be introduced here first.

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4


Table 1.1 Strength based rock classification. Qualitative name

Strength in MPa

Strength in psi

Extremely Hard Very Hard Hard Moderate Soft Very Soft Extremely Soft

More than 345 207 to 345 103 to 207 69 to 103 34 to 69 7 to 34 Less than 7

More than 50000 30000 to 50000 15000 to 30000 10000 to 15000 5000 to 10000 1000 to 5000 Less than 1000

When referring to the strength of rock, one is most commonly referring to its compressive strength. Let us presume that a cylindrical sample of rock, having a length of 100 mm and diameter of 60 mm is subjected to a steadily increasing axial compressive load and breaks when the load reaches 200 kN. In such cases the strength of the rock can be calculated by dividing the breaking force by the cross sectional area of the sample. For the specific values of the parameters mentioned here, the compressive strength Sc will be: Sc = 200*103/(π*602/4) = 70.735 N/mm2 i.e. Sc = 70.735 MPa For giving some idea of the strength of rocks, adjectives such as soft, hard etc. are arbitrarily applied before their names. The arbitrary nature of these terms can be removed to a good extent by following the nomenclature in Table 1.1.

1.4

DRILL AND BLAST OPERATIONS

Drill and blast operations have a unique place in mining and civil engineering. When the volume of hard rock mass to be excavated is very large, the use of mechanical cutting methods is snail-slow and ineffective. Blasting, however, can quickly fragment even the hardest known rocks over a large area. Breaking of rock mass into fragments requires energy. For any specific weight of rock mass the surface area is inversely proportional to the size of the fragments formed in fragmentation. When rock mass is blasted by the drill and blast method, the confinement of the explosive in the blastholes generates extremely high pressure and a huge quantity of heat. When this pressure and heat passes on to the surrounding rock mass, the rock mass just behind the wall of the blasthole gets pulverized into very fine particles. This happens because in that region new cracks are developed due to high energy concentration. However, at greater distances only the most vulnerable pre-existing cracks expand and join together to form a parting plane. Many such planes give rise to very large fragments in comparison to those formed by the other methods of breaking rocks.

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5

Naturally, blasting proves to be the most economical method of primary rock fragmentation. Figure 1.2 shows the effectiveness of different rock fragmentation methods. The lower and upper curves of the shaded area indicate operations of high and poor efficiency respectively. For a typical surface rock excavation, many blastholes are drilled systematically in a vertical or near vertical direction, filled with explosives and then blasted. These blastholes have specific diameter, depth and distance between them. After removal of the fragmented rock, depending upon the extent of excavation, many step-like benches, as shown in Figure 1.3, are formed.

Energy Required to Break in kW/kg

60 Medium Rotary Blasthole Drill (Chips)

52.5

Extra Large Rotary Blasthole Drill (Chips)

45

Road Header

37.5

Surface Miner

30

Full Face Tunneling

22.5

Ripping

15

Blasting

7.5 0

0

25

50

75 100 125 150 175 200 225 250

Mean Fragment Size in mm

Figure 1.2 Average fragment size achieved by different fragmentation methods.

Bench 1

Bench 2 Bench Height

Figure 1.3 Benches formed in mining process.

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6


Various parts of a bench and its surroundings are shown in Figure 1.4. The top openings of blastholes are in a straight line parallel to the crest of a bench; this line is called the row. Only one row of blastholes is shown in Figure 1.4 but in actual practice 3 to 8 rows of blastholes are drilled and filled with explosive to carry out a blast. More detailed nomenclature of a bench is given in Figure 1.5

Backbreak

Ideal Crest Line Crest or Bench Edge Overburden Bench Top

Blastholes

Overbreak Overhang Bench Face

Sockets

Toe Bench Floor Stump

Burden Below Grade

Figure 1.4 Bench nomenclature.

First Row

Spacing Bench Edge

Bench Top Burden

Bench Height = K Bench Floor Blasthole Bottom Level Reduced Level A Blastholes

Reduced Level of the Bench

Mean Sea Level

Figure 1.5 Bench and blasthole terminology.

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1.5

7

OPTIONS TO DRILL AND BLAST

When all is said and done, blasting is a hazardous and dangerous operation. Due to its harmful effects like vibrations, flying rocks, poisonous gases etc., it is prohibited in areas that are in proximity to human populations, important structures etc. Even if drilling and blasting is carried out in a remote area, it remains a dangerous operation for those who work on the site or live in the vicinity. Before concluding that blasthole drilling and blasting is needed, the necessity and feasibility of other ‘no blast’ methods of rock excavation have to be thoroughly examined. Excavation methods that require no blasting are: 1 2 3 4

Water Jetting Machine Digging Ripping Rock Breaking

In general, if the formation can be described in one or more of the following terms, in most of the cases it can be excavated by one of the above ‘no blast’ methods. Soil, Loess, Talus, Gravel, Peat, Loam, Sandy Clay, Alluvium, Sand Bed, Soft Coal, Shingle, Weathered Limestone, Weathered Rock.

1.5.1 Water jetting Water Jetting is used for excavation in semi-consolidated or loosely cemented mineral masses, that can be termed as soil or extremely soft rocks. In this method a high pressure water jet is directed at the rock mass. The soft rock simply crumbles under the impact of water. Broken rock fragments flow with the water. Gold mines, which are often located in the alluvial areas near a river, or soft sedimentary rock formations on the slopes of hills, use this excavation method. Since the water used for breaking the rock mass cannot easily be recovered, a huge quantity of water is needed for such operations. A typical water jetting operation on hill slopes is shown in Figure 1.6.

1.5.2

Machine digging

If well consolidated and cemented rocks, e.g. iron ore-bearing laterite, have low strength binding between the particles, they can be directly dug by means of rope or hydraulic shovels, provided a wall-type face is available for such digging. Deposits of lignite coal are often horizontal and have a layer of soft sandstone above them. Bucket wheel excavators are preferred for excavation in such lignite mines. Figure 1.7 shows a bucket wheel excavator in action. Buckets are fixed to the rotating wheel of the excavator. As the wheel rotates, the buckets scrape or rip the soft overburden which fills the buckets. As the buckets tilt, the material collected in the bucket falls on the conveyor belt system of the excavator. The conveyer belts then carry the material to the dumping area. Horizontal beds of lignite are thus brought into the open. This method proves economical and effective due to its continuity in excavation.

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8


Figure 1.6 Water jetting in gold mining.

Figure 1.7. A giant bucket wheel excavator.

1.5.3

Ripping

When the rock mass is too hard for excavation by water jetting or machine digging, a third available alternative is ripping. Ripping has been used in the farming industry for thousands of years under the name plowing. Ripping means deep scraping the ground with one or many teeth attached to a moving machine. This method proves very economical in soft and medium-hard formations with compressive strength less than about 70 MPa.

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9

There are many ways in which ripping is carried out. Most of the crawler dozers can be equipped with an optional single shank or multi shank ripper attachment. This attachment is fixed at the back of the dozer and is operated hydraulically to force the tooth or teeth in the ground to the desired depth. This depth is suitably chosen according to machine power and formation hardness. Once this is done the dozer is moved. As the dozer moves, the tooth or teeth break the surrounding rock like a plow in farming. A single tooth behind a large dozer can penetrate even 1 m depth in medium soft ground. Light duty ripping can be carried out with motor graders by using a ripper attachment. Figure 1.8 illustrates a dozer, equipped with a three shank ripper attachment, busy ripping in a medium soft formation. Nowadays large dozers, equipped with powerful engines and strong single or multi-shank ripper attachments, are encroaching on the traditional drill and blast territory. Formation rippability can be decided on the basis of the seismic wave velocity. A well known chart universally adopted for the purpose, is reproduced in Figure 1.9. If formations composed of other types of rocks are encountered, verification of their rippability, either by means of field tests or reliable information on ripping in similar formations, is desirable. If the formation is highly fissured and jointed, ripping becomes easier. During the process of ripping some large size boulders are often encountered. Such boulders can be broken with hydraulic hammers attached to a hydraulic excavator. For a long time, machines having a rotating drum with long teeth have been used for scraping the top bituminous layers of roads. This is also an example of ripping. Recently, similar but very heavy and powerful machines have been introduced in the mining industry under the name ‘Surface Miner’. Some small mines and civil engineering contractors opt for such surface miners. One such surface miner is shown in Figure 1.10.

Figure 1.8 Three shank ripper attachment behind a crawler dozer.

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10


Seismic Wave Velocity in m/s 0

1000

500

1500

2000

2500

3000

3500

4000

4500

Topsoil Clay Glacial Till Igneous Rocks Granite Basalt Trap Rock Sedimentary Rock Shale Sandstone Siltstone Claystone Conglomerate Breccia Cliche Limestone Metamorphic Rocks Schist Slate Minerals and Ores Coal Iron Ore Rippable

Marginally Rippable

Non Rippable

Figure 1.9 Formation rippability chart.

Figure 1.10 A surface miner.

Surface miners are equipped with a rotating drum on their front side. The drums have tungsten carbide teeth fitted into the fixtures provided for the purpose. As the drum is pushed into the ground by means of hydraulic cylinders provided for the purpose, and is rotated by means of the hydraulic motors within the drum, the teeth

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11

scrape or break the ground. As the teeth of surface miners are of small size, compared to those of a ripper, the broken rock fragments are also of small size. Compared to dozers with a ripper attachment these machines are less effective on hard rocks but more effective on soft rocks.

1.5.4

Rock breaking

A hand held paving breaker is a very commonly used tool for breaking pavements in cities. In essence it is a percussion hammer. When the rock mass to be broken is hard and large in volume, the energy output of the paving breaker falls very short of requirements. Therefore, when large scale hard rock excavation is to be carried out, very large size percussion hammers mounted on the booms of hydraulic excavators are used, as shown in Figure 1.10. Hydraulically driven percussion hammers require very low energy input due to their efficient power transmission as compared to the compressed air hammers. A hydraulic excavator is chosen for mounting these hammers because the hydraulic excavator is easily available, works on hydraulic energy, is very versatile in placing the hammer at any alignment and can reach tall heights. All that has been said above may induce readers to draw a conclusion that in soft and medium-hard rippable formations no blasthole drilling and blasting is necessary. This is not true. Overburden in some coal mines can have marginal rippability. Coal can also have marginal rippability. Even if both overburden and coal in coal deposits are rippable, blasthole drilling and blasting are often practiced for overriding considerations as follows. 1

2

3

4

A sizable part of the sedimentary formation may be non-rippable. In such cases, if bucket wheel excavators are used they may either prove unsuitable or the hard portion has to be blasted by drilling closely spaced small diameter blastholes so that the resulting rock fragments are sufficiently small, and their excavation and transportation can be done by using the bucket wheel excavators. It is to be noted that by their design and construction, bucket wheel excavators are suitable only in easily rippable formations. When the formation in a coal mine is difficult to rip and the coal seam is very thick, it is often more advantageous to use a dragline for overburden removal and shovels for coal removal. In such an event blasthole drilling and blasting becomes necessary in both overburden and coal. Dozer ripping can be done in relatively hard formations. However, when such ripping is done the broken formation remains in a horizontal layer. Loading of waste from such broken layers into the hauling equipment, without unacceptable air pollution, is difficult. Moving the fragmented rock mass by using dozers over a long distance is often impracticable.

One important factor, to be taken into account before resorting to drilling and blasting or ripping, is the cost of crushing. Fragment size distribution achieved by drilling and blasting is far different than that by machine digging or ripping.

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Figure 1.11 Hydraulic hammer attached to an hydraulic excavator.

Small size fragments can be obtained by ripping with multi shank rippers. Similar adjustments can also be made in blasting where the diameter of the holes as well as burden (the distance between two adjacent rows), and spacing ( the distance between two adjacent blastholes in a row), can be reduced to give smaller blast fragments. In any case for large scale excavation operations, where the rock is non-rippable, drilling and blasting is the only available alternative. Needless to say nearly 80% of the excavation in mining activities is carried out by drilling and blasting.

1.6

OILWELL DRILLING

One more important and very useful mineral that is abundantly available in the crust of earth is crude petroleum oil. In terms of the importance attributed to the mineral obtained, oil gets far more preference than other minerals that are in solid form. Oil is retrieved from the earth’s crust by means of drilling holes into the oil reserves. Since oil is in fluid form, it can be easily pumped by pumps that are placed on the ground. For this reason, there is no necessity of blasting in the oilwell drilling industry. Rotary drilling techniques are invariably used in current oilwell drilling practice. The reasons are: 1 2 3

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Oil reserves are very often at great depths and it is almost impossible to reach to such depths by using other drilling methods. Progress achieved by other drilling methods is very slow. Many techniques such as hole deviation, well cementation, blowout prevention etc. are not possible without using many accessories that can be used only with rotary drilling.

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13

Due to the economic preference given to the oil industry over the mining industry, far more rotary drilling research is carried out by the oilwell industry than is carried out by the mining industry. However, rotary oilwell drilling vastly differs from rotary blasthole drilling. Similarities between the two are limited only to the fact that both use rotary drilling Table 1.2 Comparison of oil well drilling and rotary blasthole drilling.

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Aspect

Oil well drilling

Blasthole drilling

Hole Depth

Holes are very deep. Depth often exceed 5000 m.

Holes are shallow. Depth rarely exceeds 50 m.

Hole Diameter

Holes may start with large diameter, sometimes even 600 mm, but the diameter reduces to about 200 mm at maximum depth.

Holes have uniform diameter from top to bottom. Diameter ranges between 152 to 445 mm.

Hole Spacing

There is no defined pattern of hole placement.

Holes are always drilled in a well defined rectangular or triangular pattern.

Hole Straightness

Holes are rarely straight. Often they are purposefully deviated.

Holes are required to be straight. Hole deviation is undesirable and avoided.

Need for Casings

Holes are always cased to considerable depth. Casing within casings i.e. telescoping casings are usual. Huge quantity of tubular steel is required for casings.

Holes are never cased.

Drilling Rig

Rig is in modular form. Several pieces of equipment such as diesel engines, derrick, drawnwork, mud pump etc. are assembled at the site of the hole and after completion of the hole, all the pieces are dissembled. Moving from one hole to the other requires 20 to even 100 truckloads and takes 15 days upward.

Rig is composite and is termed a rotary blasthole drill. It contains everything that is needed for drilling the hole. It is moved from one hole to the other in about two minutes and does not require any other carriers. Drill rods are added to the drill string automatically by pipe changer.

Flushing Medium

Water is a usual medium of flushing. Often many different additives are mixed with water for specific purposes. In very few cases gas is also used as medium.

Air is invariable flushing medium. Small quantity of oil and water is mixed for bit cooling, bit lubrication as well as dust suppression.

Drill Pipes

Mostly integral drill pipes are used. High tensile strength is the most important property. Pipes are external upset and internal flush.

Mostly fabricated drill pipes are used. High surface hardenability is the most important property. Pipes are external flush.

Testing in the Hole

Many tests are carried out in the hole at various depths by pausing the drilling.

No test is carried out by pausing the drilling. In fact there are no tests.

Drilling in Sea

Holes are often drilled in sea from specialized ships, jackups, submersibles etc.

Holes are always drilled on land by the drill standing on the surface of ground.

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14


Figure 1.12 Scenario of a typical oil well drilling operation on land.

techniques and tricone bits attached to the drill string composed of drill rods. Even these accessories, used in these two domains of drilling, are not interchangeable. Of several points of dissimilarities only a few are mentioned in Table 1.2. Due to such differences, the research carried out in rotary drilling by the oilwell industry is of very limited use for the mining industry. A typical scenario of oilwell drilling operation is shown in Figure 1.12. Oilwell drilling equipment usually comprises many components that are carried to the worksite in several trucks and assembled for drilling operations.

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Chapter 2

An overview of blasthole drilling

2.1

INTRODUCTION

The most effective, and hence most usually adopted, method of destruction of an object is by blasting. This method invariably aims at breaking the object into several small pieces. Explosives are used for the purpose of breaking the object. An explosive is a substance that generates a very high volume of gases and a huge quantity of heat within a very short time span of a few milliseconds as soon as it is ignited. When gases are generated by ignition in this way, very high pressure is exerted on the surroundings. At the same time, the strength of the surrounding material also reduces considerably due to the heat. As a result of this, cracks develop in the material and the material gets fragmented. When the place of generation of gases is confined, the gases cannot escape to atmosphere and the intensity of the pressure increases to a considerably higher level. This results in maximum destruction. For the most effective destruction of surrounding material, a hole is drilled in the material to be destroyed. After completion of drilling, the hole is charged (i.e. filled with explosive material) in a certain predefined manner. A detonator, connected by two copper wires, is kept in the hole while the explosive is filled. The hole is then sealed from the top. Once the appropriateness of all the precautionary measures is verified, an electric current is sent through the wires. The electric current heats the wire in the detonator and the explosive material in the detonator explodes. This explosion causes further detonation in the explosive filled in the hole. The hole drilled for filling explosives in this manner is called a blasthole. The process of drilling such holes is called blasthole drilling. A machine used for drilling the hole is called a blasthole drill or often simply a drill. Except cases like breaking a boulder in field etc., a blasthole is seldom drilled in singular. They are drilled in numbers, perhaps in hundreds, one after another, then charged and blasted simultaneously. The fragmented rock mass formed by blasting is almost always moved away, either to create space, as in civil engineering projects or to extract a mineral of interest, as in mining or quarrying. If the sizes of the fragments are small they can be moved easily. In order to have small size of rock fragments after a blast, blastholes are required to be located in a properly planned manner. This requires careful design to determine the diameter, depth and inclination of the holes, and also layout and other relevant parameters of the holes. Rotary drilling is one of the two methods currently used for drilling blastholes in the ground. The other method is called rotary percussion drilling. This book deals with rotary blasthole drilling and covers it from the theoretical as well as practical aspects.

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16

2.2


BASIC CONCEPTS OF DRILLING

In actual practice, drilling is carried out by using a drill bit connected at the bottom of a drill string that consists of drill pipes, as shown in Figure 2.1. A drill string often consists of some other items, not shown in the figure, that have specific purposes. The drilling bit is meant to interact with the rock. When energy in some form is passed to the rock from the drill bit the rock fractures. The drill string enables a continuous supply of energy to the drill bit so that rock fracture continues at progressive depths. As the bit interacts with rock and breaks it, rock cuttings are formed. These cuttings are moved away by means of a circulating fluid that is pumped to the bit through the central bore in all the drill string components. When fluid escapes from the bit and starts moving up to the ground surface through the annular space, as shown in the figure, it drags the cuttings along with it and deposits them on the ground surface. Thus, the bit keeps on breaking fresh portions of the rock.

Drilling Fluid Moving Up through the Annular Space Between Hole Walls and Drill Pipes

Drill Pipe

Drilling Fluid Moving Down through the Central Bore of the Drill Pipes

Drill Bit

Figure 2.1 Drilling Schematics.

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17

The drill string is also given rotary motion so the bit rotates and its cutting points cover a complete cross sectional area to form a cylindrical blasthole. The drilling process can commence, continue, and be complete only when the following essential technical requirements are fulfilled. 1

2 3 4

5

6 7 8 9

The drilling bit, meant for breaking the rock formation, must have adequate hardness, heat resistance, strength, toughness, etc., so it can keep on breaking the formation over a long period of time. Energy, in one of the many possible forms, must be supplied to the bit to cause rock disintegration. The drill bit must be rotated to ensure a circular hole, irrespective of bit cross section. Components in the drill string must also have specific properties to effectively transfer feed force, rotary torque, percussion blows etc., without undue deformations. The rock cuttings formed in the process of formation disintegration must be taken out of the hole as soon as possible, so they do not waste energy in getting crushed once again. The drill should be able to supply sufficient energy to the drill bit through the drill string components to cause formation fracture. The drilling bit must be advanced in the intended direction of the drillhole. The walls of the hole, formed as the drilling progresses, must be prevented from collapsing inside the hole. After attaining the desired drillhole depth, it must be possible to withdraw the drill string composed of drill bit, drill pipes and other accessories.

Combinations of means of fulfilling these essential requirements give rise to different drilling methods. Besides the above technical requirements, it is also necessary to fulfill certain other environmental and economic needs. These are, I Cuttings of very small size – in the form of dust – generated during the process of drilling, should be prevented from mixing with atmospheric air and polluting it. II The process of drilling should continue at optimum speed so as to keep the cost of the operation to a bare minimum. III Noise generated by the drilling equipment must be kept to an acceptable low level. Theoretically, there are many forms of energy that have been applied to a rock mass for its disintegration in laboratories, but in current field practice only mechanical energy is used. Two forms of mechanical energy, viz. percussive and rotary, are used for drilling. In either case the drill string and the drill bit are rotated. Accordingly, the drilling methods are called rotary percussion drilling and rotary drilling. A schematic of both these drilling methods is in Figure 2.2. In rotary percussive drilling the energy generated at the drill is in the form of repeated blows like those of a hammer used in manual chiseling. These blows create

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Heavy Feed Force Without Percussive Blows

Low Feed Force With Percussive Blows

High Torque Low Torque

Large Diameter Drill Pipes

Small Diameter Drill Rods Drill Bit With Rotating Cones

Drill Bit Without Moving Parts

Figure 2.2 Schematics of percussion and rotary drilling.

shock waves in the drill string. The shock waves that travel to the bit are passed on to the rock mass through the cutting edges or points on the bit. As a result, cracks are formed in the rock mass and result in formation of rock chips. In rotary drilling the drill pushes the bit into the rock mass with very heavy force. When this heavy feed force is transmitted to the rock mass through the cutting points of the drill bit, cracks develop and rock chips are formed. In both rotary and rotary percussion drilling, cuttings are removed from the hole by a fluid circulated from bottom to top. A drill contains many assemblies meant for various purposes. In a rotary percussion drill the assembly used for generating blows is called the hammer. Hammers are of two types viz. top hammers and down-the-hole hammers. Figure 2.3 shows rotary percussion drilling with these hammers. A top hammer is so called because it always remains at the top of the drill string and hence out of the hole. It creates impacts by causing to and fro movement of a piston contained within it. In addition, it also generates rotary motion through a motor within it. The piston movement or rotary motion are caused by hydraulic fluid or compressed air (now obsolete) depending upon the type of drill. Top hammers give fastest penetration rates in hard rocks. They are mostly used for drilling holes for different purposes on construction sites, small opencast mines,

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19

Feed Frame Feed Motor

Rotary Head

Boom Hoisting Cylinder

Top Hammer Feed Frame Hoisting Cylinder Boom

Drill Pipe Crawler Mounting DTH Hammer

Drill Rod Drill Bit

Figure 2.3 Schematics of top hammer and down the hole hammer drilling.

quarries and underground mines. The diameter of holes that can be drilled with top hammers is usually within 38 to 152 mm. The efficiency of top hammer drilling reduces rapidly as more drill rods are added to reach greater depth. Drilling holes deeper than 30 m with a top hammer is difficult because of the energy lost at the connections of the drill rods. Some top hammers are rated for drilling holes to 50 m depth by use of special drill string components. A down-the-hole hammer, usually abbreviated as DTH hammer, is attached to the bottom of the drill string. A drill bit is attached to the lower end of the DTH hammer. A DTH hammer always goes inside the hole along with the bit. The movement of the piston in the DTH hammer is caused by compressed air. The rotary movement for drilling is generated by the rotary head in the drill and is passed on to the drill bit through drill string components and DTH hammer. DTH hammers give fast penetration rates in hard rocks. They are also used for drilling holes for different purposes on construction sites, opencast mines, quarries and underground mines. The diameter of holes that are usually made with DTH hammers is within 100 to 200 mm. Many large hammers, used with specially built drills, are capable of drilling holes of diameter up to 900 mm and in case of specially built hammers even to 1500 mm. Holes of depth 1000 m or even more can be drilled by DTH drilling as the impact energy used for drilling is generated right above the bit and is, therefore, not lost in transmission.

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Drill bits used with top or DTH hammers are made from alloy steel. On one side they have tungsten carbide inserts. For most of the bits used now, the inserts are in the form of buttons because such bits are found to be superior to chisel shaped inserts arranged in cross or X shape. On their other side, top hammer bits have female threads that enable them to be attached to the drill rods. DTH bits have splines and collars for attachment to the hammers. Tungsten carbide insert buttons are shrink fitted in the holes made in the steel body, whereas the chisel-shaped inserts are brazed into pre-machined grooves by means of a suitable brazing alloy. Top hammer drilling bits are small in size and are firmly coupled to the drill rods by means of threads. DTH bits are loosely coupled to the DTH hammers and have splines to ensure rotation as well as a collar at their top end to ensure that the bits are not detached from the hammer while drilling is carried out or when they are retracted after completion of the hole. Figures 2.4 and 2.5 show some bits used in top and DTH hammer drilling respectively. Schematic of rotary blasthole drilling is shown in Figure 2.6.

Button Bit

Cross Bit

X Bit

Figure 2.4 Bits used in top hammer drilling.

Splines

Collar Cross Bit

X Bit

Button Bit

Figure 2.5 Bits used in down the hole hammer drilling.

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21

Mast Drill Head Shock Absorber Mast Braces Machinery House Operator’s Cab

Crawler Mounting

Drill Pipes Stabilizer Drill Bit

Figure 2.6 Schematics of rotary blasthole drilling.

In a rotary blasthole drill, the rotary motion is generated at the drill head by hydraulic or electric motors. The head is pushed down through the feed mechanism to exert the necessary feed force. Drills used for rotary blasthole drilling are very large and heavy, as they have to exert a very high feed force and high torque on the bit through the drill string. Crawler mounting is best suited for the large, tall and heavy drills and for traveling in rough ground. The components, from top to bottom, in a rotary drill string are a shock absorber, a saver sub, drill pipes, a stabilizer and a drill bit. Rotary drilling is carried out by using tricone roller bits in soft to very hard rock mass. In a soft rock mass, drag bits or claw bits are also used. These three types of bits are shown in Figure 2.7. All the three types of bits use tungsten carbide inserts fixed onto a steel body either by shrink fitting, brazing, or mechanical pin or circlip fitting. Tricone bits have three rotating cones fixed into the bit legs through roller and ball bearings. Unlike rotary percussion bits, they have moving parts. Under the heavy force exerted onto the bit, the buttons or teeth on the rotating cones of the bit penetrate into the rock mass. With the heavy torque, they crush, scrape or plow the rock mass to create cuttings. These cuttings are moved out by drilling fluid circulated through the central bore of the drill string components. When rotary or rotary percussion drilling methods are used for blasthole drilling, it is customary to use compressed air as the drilling fluid because it is far more efficient in cutting removal as compared to water or other types of drilling fluids.

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Blade Bit

Claw Bit

Tricone Bit

Figure 2.7 Rotary drilling bits.

2.3

PECULIARITIES OF BLASTHOLE DRILLING

Compared to the other purposes of drilling, such as waterwells, oilwells, exploration etc., blasthole drilling has some peculiarities as follows, 1

2

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Blastholes are drilled at the same location. In the case of huge mining projects there may be a need to drill many million blastholes within an area of say ten square kilometers. Most of such mining projects last for fifty or more years. Some mines are productive even for a century. When a set of blasthole drills becomes inefficient after years of use it is replaced. In many mines, not only one or two but several such blasthole drill generations are replaced one after another. In the case of tunnel construction all the blastholes lie along the alignment of the tunnel. Here too, millions of blastholes are localized between the two ends of a tunnel. For other purposes of drilling, such as oil or water wells only a few holes are drilled at one location. The drilling work may last for a few weeks or months at the most. The location of the next drill hole may be far away, sometimes even thousands of kilometers. Blastholes are very near to each other. In surface mining operations several hundred blastholes are drilled at each of the benches. The distance between the blastholes is often from 3 to 12 m. In the case of underground mining or tunneling Operations, many blastholes are drilled at the face of tunnel. The distance between two blastholes can be as low as half meter or even lesser. Therefore, a drill is not required to be moved but the feed frames of the drill can be moved and suitably aligned for drilling each of the blastholes. Movement of the drill becomes necessary for drilling blastholes in the next face after the blasted fragments are removed. This distance is often from 3 to 5 m. In rock faces formed after blasting, blastholes are drilled in a very specific pattern, that is more or less the same as previous. For other purposes of drilling, even if few holes are required to be drilled at the same location, the distance between holes can be very large. Further, in many cases there is no definite pattern for drilling such holes.

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3

4

5

6

7

23

Blastholes are shallow in depth. In tunneling, the blastholes are horizontal and rarely exceed 6 m in length. In underground mining blastholes may be reach 30 m in length or depth. In the case of surface mines, blasthole depths vary from 10 m to 50 m depending upon the mining method and the machinery used for moving the rock. For other purposes of drilling, such as oil or water wells, the hole depth is commonly in terms of hundreds of meters. Particularly in oilwell drilling, hole depth can be between one to seven kilometers or even more, the deepest being over 10 km. Blastholes are drilled in rock masses that have a high degree of uniformity. In surface mining the blastholes are drilled in a more or less uniform rock mass. This rock mass may be stratified but the properties of the strata do not vary greatly. In underground mining, and particularly in tunneling, the rock mass characteristics may greatly differ from one place to the other but the variation within a group of, say, 100 blastholes is very small. For other purposes of drilling, the variety of rock masses can be very large. In very few cases, if any, the rock mass at two locations of the hole is the same. Blastholes are drilled in the same environment. In surface or underground mining as well as tunneling, the holes are required to be drilled at almost the same depth and surroundings. For other purposes of drilling, two holes may be in radically different surroundings. For example, one waterwell may be in a busy locality near a sea port, whereas the next hole may be at a hill station at high altitude. No testing is done in blastholes. In the case of blastholes, no testing is involved. The holes are drilled, charged with explosives and fired. After removing the fragmented rock by using loading and hauling equipment, the process of drilling is continued once again over the freshly exposed area. In a few instances it may be essential to carry out a few tests in blastholes but such tests are carried out only after the hole is drilled and the drill is moved to next blasthole location. For other purposes of drilling, a lot of ‘in hole’ testing may have to be done in the midst of the drilling process. For hole path orientation test, permeability test, borehole logging etc. special tools have to be lowered into the hole. Blastholes are always straight. Barring unavoidable deviation, blastholes are intended to be in a straight line. In other purposes of drilling, holes may be purposefully deviated and so are on a curved path.

In view of all these peculiarities of blasthole drilling, the drills, components and tools are designed with an objective of giving fastest economical output in terms of meters drilled.

2.4

COMPARISON OF DRILLING METHODS

Rotary percussion and rotary drilling methods described above have been compared in Table 2.1. The comparison is made on a technological basis from the blasthole drilling viewpoint.

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Table 2.1 Comparison of blasthole drilling methods on the basis of essential requirements. Rotary percussion blasthole drilling

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Criteria of comparison

Top hammer drilling

Down the hole hammer drilling

Rotary blasthole drilling

Means of Achieving Formation Disintegration

By means of blows generated at the drilling machine located above the ground.

By means of blows generated in the down the hole hammer placed just above the drill bit in the drill hole.

By means of static pressure given by the feed force exerted by the drilling machine placed on ground.

Type of Drill String

Relatively slender as it comprises of drilling rods coupled together by means of sleeves.

Relatively rigid because it comprises of large diameter drill pipes that have male and female threads on their ends. No sleeve is required for coupling the pipes.

Rigid because it comprises of even larger diameter and heavier drill pipes that have male and female threads on their ends. No sleeve is required for coupling the pipes.

Transfer of Impacts and Rotary Motion

Both impacts and rotary movement are transferred through drill string.

Only the rotary movement is transferred through drill string. Impacts are generated at the hammer in the hole just above the bit.

Very heavy static feed force instead of impacts and rotary movement are transferred through the drill string.

Practical Hole Dia. Range

38 mm to 150 mm.

89 mm to 250 mm.

150 mm to 445 mm.

Practical Hole Depth Range and Susceptibility to Hole Deviation

0 to 18 m. Highly susceptible to deviation at higher depths due to high slenderness ratio of the drill rods.

0 to 100 m. Hole deviation is relatively small because drill pipes have low slenderness ratio.

0 to 60 m. Hole deviation is relatively small because drill pipes have low slenderness ratio.

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Possibility of Angle Hole Drilling

As holes are drilled with light equipment they can be drilled at any angle.

As holes are drilled with medium equipment drilling holes in direction other than vertical or near vertical is difficult.

Drilling equipment is very heavy.Vertical or slightly inclined holes are possible.

Effect of Depth on Penetration Rate

Penetration rate is considerably reduced with depth because of loss of percussion energy. It is absorption by drill rods.

No significant loss of penetration rate as the hole depth increases because no energy is lost in transmission.

Penetration rate is unaffected by increase in the depth of the hole because impact energy is not used.

Means of Cutting Removal from the Hole

By means of compressed air circulated through the drill string. Sometimes water is also used as circulation medium.

By means of compressed air circulated through the drill string.

By means of compressed air circulated through the drill string.

Magnitude of Feed Force Required on Drilling Tool

Small force is essential to ensure that the bit is in contact with the formation.

Small force is essential to ensure that the bit is in contact with the formation.

Very heavy force is essential to ensure that the bit remains in contact with the formation and crushes the formation beneath it.

Effect of Extreme Cold Weather

As hydraulic oil is the medium of power transfer, great difficulties are experienced due to oil viscosity changes.

As compressed air is the medium of power transfer, difficulties in cold weather are easily surmountable.

Electric drills operate satisfactorily. In hydraulic drills difficulties are experienced due to oil viscosity changes.

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26


Table 2.2 presents a qualitative comparison between top hammer and downthe-hole hammer drilling methods in typical limestone quarries. Faster penetration rate is one of the most important criteria of a drilling method. Figure 2.8 illustrates the penetration rates experienced while drilling about 150 mm blastholes in quarries by Rotary, DTH and Top Hammer drilling methods. The variation of penetration rates is of a qualitative nature.

Table 2.2 Qualitative comparison of TH and DTH drilling. Criteria of comparison

TH drilling

DTH drilling

Diameter of Blastholes in mm Penetration Rate Achieved Straightness of the Blasthole Suitability for Drilling Deep Blastholes Production Capacity per shift Low Fuel Consumption Economic Service Life of Drill String Suitability for Drilling in Difficult Conditions Suitability for Drilling in Good Conditions Simplicity of Drilling From Operator’s Viewpoint

75–127 ♦♦♦♦ ♦♦♦ ♦♦♦ ♦♦♦♦ ♦♦♦♦♦ ♦♦♦ ♦♦♦ ♦♦♦♦♦ ♦♦♦♦

127–300 ♦♦♦ ♦♦♦♦♦ ♦♦♦♦♦ ♦♦♦ ♦♦♦ ♦♦♦♦ ♦♦♦♦ ♦♦♦♦ ♦♦♦♦♦

Hard Monolithic Rock Mass DTH

Rotary Drifter

Increasing Penetration Rate


Soft Rock Mass

DTH Rotary

Increasing Rock Strength

Drifter



Hard Fractured Rock Mass

DTH Drifter Rotary


Figure 2.8 Penetration rates experienced in drilling about 150 mm blastholes in limestone.

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27

It must be borne in mind that the trend of penetration rates shown in Figure 2.8 does change when the rock to be drilled is different and particularly when the holes are of larger or smaller diameter.

2.5

CHOICE OF A BLASTHOLE DRILLING METHOD

Once it is decided that blasthole drilling is to be carried out in preference to other ‘no blast’ methods of rock excavation described in chapter 1, the next logical question is “Which should be the blasthole drilling method?”. The main factors that influence the choice of drilling method are: 1 2 3 4

Period over which operations continues. Diameter of the hole. Depth of the hole. Characteristics of the formation to be drilled.

The most important factor that affects the choice of method is the continuity of operations. In the case of civil engineering, the quantum of blasthole drilling to be carried out is relatively very small as compared to mining. To be specific, a civil engineering contractor may rarely have a need for continuous blasthole drilling at one location for more than two years. Thus, for a civil contractor it is wise to select many smaller machines rather than few large machines for his civil engineering projects, because after completion of the project the smaller machines can be distributed on many separate small jobs. Blastholes in civil engineering are rarely larger than 127 mm in diameter and almost never larger than 200 mm. The depth of a blastholes is also limited to about 20 m. Civil engineering firms, therefore, usually choose top hammer drilling for blastholes of smaller diameter and DTH drilling in case of larger blastholes. In mining choosing the right drilling method is very critical as the blasthole drilling operations continue for decades and blasthole drills specifically designed for a method are to be procured before commencing the operation. Choice of incorrect drills can be disastrous as reselling drills at a fair price is difficult. When blasthole drilling is to be carried out it in the soft sedimentary overburden of opencast coal mines, the diameters of the hole range between 152 to 381 mm and depths range between 15 to 55 m. In soft thick coal beds blasthole diameters range between 152 to 200 mm and depths between 10 to 20 m. For soft sedimentary formations suitability of a blasthole drilling method is illustrated in Figure 2.9. In many opencast metal mines the formations are hard and the variation in hardness is over a very wide range. Here both overburden and ore are removed by using shovels and dumpers. For this reason the blasthole depths lie between 10 to 20 m and diameters can be from 100 to even 445 mm. For all such cases the blasthole drilling method can be chosen on the basis of the illustration in Figure 2.10. In underground coal or metal mines, blasthole diameters are between 51 to 100 mm but in certain cases can be as high as 152 mm. The depths of blastholes can be from 2 to 30 m.

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28


Top Hammer Rotary with Drag or Tricone Bit Rotary with Tricone Bit (Rarely Needed) Rotary Drag Bit

Rotary with Tricone Bit

Soft Very Soft 50

0

100

150

200

250

300

350

400

450

Blasthole Diameter in mm

Figure 2.9 Suitability of blasthole drilling method in very soft and soft formations. TH Drilling DTH Drilling TH or DTH Rotary with Tricone Rotary with Drilling Bit or DTH Tricone Bit Extreme Hard Very Hard Hard Medium Hard 0

50

100

150 200 250 300 Blasthole Diameter in mm

350

400

450

Figure 2.10 Suitability of blasthole drilling method in medium to very hard formations.

In coal mines it is very common to use rotary drag bit drilling. In hard coal formations top hammers are often used for drilling blastholes. When holes of diameter 152 mm to 30 m depth are to be drilled, as in some metal mines, DTH hammer is the most usual choice because the size of the drilling machines used for DTH drilling is small. In tunneling or shaft sinking operations, diameters of blastholes are 38 to 64 mm. The depths of the blastholes, in a very few instances, can be as much as 25 m but in most of the cases depths are between 2 to 6 m. The top hammer rotary percussion drill is almost exclusively used in all such operations. Two important points must always be remembered while choosing the right drill for blasthole drilling. Rotary drilling requires exerting a heavy feed force in the downward direction on the drill string. Hence the drills are very heavy. If they are not heavy, their ability in drilling large diameter blastholes remains limited and must be investigated cautiously. Percussive drills need not exert heavy feed force, but when drilling upward holes the weight of the piston acts adversely to the blow energy. Percussive drill performance must be judged in the light of this fact.

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29

The performance of blasthole drilling equipment can be judged by means of carrying out certain laboratory tests. However, due to many influencing factors, it cannot be totally relied upon, while tackling an important issue like choosing a blasthole drill. Field trials by using hired equipment before taking a final decision about the suitable blasthole drilling method is a far better alternative, even if it is expensive. In this chapter both rotary percussion and rotary drilling methods have been described briefly. As the subject of this book is rotary blasthole drilling, in all forthcoming chapters all the elaborations are presented from the a viewpoint of rotary blasthole drilling.

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Chapter 3

Properties of rocks

3.1

INTRODUCTION

Blasthole drilling is a part of earth excavation processes. Needless to say, some basic knowledge of earth science, and particularly the properties of rock and rock masses, is very essential to a person engaged in blasthole drilling. Earth science is a very wide subject and has many branches within it. Branches that are important in the context of blasthole drilling are geology, mineralogy, meteorology and rock mechanics. This chapter presents essential knowledge of Geology and mineralogy. Topics on rock mechanics are covered in chapter 9. Meteorology is elaborated in chapter 11. The properties of rock masses have been covered in chapter 23 as they are more relevant to blasting than drilling.

3.2

EARTH AND ITS INTERIOR

Earth is one of the planets revolving around the sun in an elliptical orbit, having a mean diameter of 1.496 × 106 km. The shape of the earth is almost spherical. Earth rotates around its axis and completes one rotation in one day. The two points at which the earth’s rotational axis crosses its surface are called poles. According to the magnetic properties of the earth these poles are named south and north. An imaginary circle, perpendicular to the earth’s rotational axis, that lies on the surface of the earth and thus divides the earth into two equal hemispheres, is called an equator. The equatorial diameter of the earth is 12756.6 km and the polar diameter is 12713.8 km. The volume of the earth is 1.08323 × 106 km3. The total surface area of the earth is calculated to be 510 × 106 km2 of which land covers only 149 × 106 km2 i.e. 29.22%. The remaining 70.78% area is covered with water. Average land height is estimated to be about 840 m above the mean sea level. The average depth of the water below the mean sea level is 3804 m. Various experiments have proved that the average density of the earth is 5517 kg/m3 and accordingly the mass of the earth works out to be 5.976 × 1024 kg. The interior of the earth consists of three zones as illustrated in Figure 3.1. The innermost zone, spherical in shape with a radius of about 3400 km, is called the core. The core itself is subdivided into an inner core and an outer core. The outer core is

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Inner Core Outer Core Inner Mantle Crust

Outer Mantle

Figure 3.1 Interior of the earth.

believed to be composed of molten iron and nickel and has a density of 10000 kg/m3. The inner core is believed to be in a solid state and has a density of 17000 kg/m3. The next outer zone, called the mantle, has a thickness of about 2900 km. The mantle is again subdivided into two zones viz. inner and outer mantle. The inner and outer mantle have a density of 5700 kg/m3 and 3300 kg/m3 respectively. The outer mantle consists of rock-forming minerals. The outermost zone is called the crust. It is contained in a shell of thickness varying between 5 to 40 km. As illustrated in Figure 3.2, the thickness of the crust is smaller below the oceans and seas and is much larger below the middle portion of the land—particularly mountains. These portions of the crust are called oceanic crust and continental crust. Earth’s crust is also divided in sial and sima. Sial, i.e. outer shell of the crust, is mainly composed of silica and aluminum and hence the acronym sial. Sial is present below the continental shelf and shallow sea/ocean floor. Its thickness varies between 0 to 13 km. Sial is absent below the deep ocean floors. The average density of sial shell is 2700 kg/m3. It is mostly composed of granitic rocks mainly formed from quartz and orthoclase. Sima is the inner shell of the crust. As it comprises silica and magnesia, it is named by the acronym sima. Sima is present everywhere i.e. below the sial as also the ocean floors. Sima thickness increases towards the center of the continental crust. Average density of sima is about 2940 kg/m3. It is mostly composed of basic rocks e.g. gabbro and basalt. Most of the elements found in earth’s crust are in compound form of which oxides are the main type. The proportions of various oxides found in the earth’s crust is given in Table 3.1. The proportion of elements that form these compounds is given in Table 3.2. Oxides chemically combine with each other, or with other substances, to form minerals that are the main constituents of rock. A mineral is an inorganic substance having definite chemical composition and properties. Earth’s crust contains more than 2000 minerals.

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Properties of rocks

Mountains

33

Oceanic Crust Ocean

Sial

Sima Mantle

Figure 3.2 Earth’s crust.

Table 3.1 Oxide proportion in earth’s crust. Common name of oxide

Chemical formula

% Wt. in earth’s crust

Silica Corundum Calcium Oxide Ferrous Oxide Sodium Oxide Magnesium Oxide Ferric Oxide Potassium Oxide Water Titanium Dioxide Other Oxides

SiO2 Al2O3 CaO FeO Na2O MgO Fe2O3 K2O H2O TiO2 –

59.1 15.2 5.1 3.7 3.7 3.5 3.1 3.1 1.3 1.0 1.2

Table 3.2 Element proportion in earth’s crust.

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Element name

Chemical symbol

% Wt. in earth’s crust

Oxygen Silicon Aluminum Iron Calcium Sodium Potassium Magnesium Titanium Hydrogen Other Elements

O Si Al Fe Ca Na K Mg Ti H –

47.0 27.0 8.0 5.0 3.5 3.0 2.5 2.0 0.5 0.17 1.33

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34


3.3

GEOLOGICAL CYCLE

One of the fundamental laws in respect of matter states that matter can neither be created nor destroyed but can only be altered. Accordingly, all the land that has come into existence, is actually as a result of transformation through many activities. Such activities are going on even now and will never stop. Figure 3.3 shows a flow diagram which indicates all such activities. These activities are geologic in nature i.e. they commence, continue and complete within the earth and its surroundings. Being cyclic in form they are called the ‘Geological Cycle’. Magma is the most primary matter in the geological cycle. It is a very hot, viscous, siliceous rock melt containing many minerals, water and gases. In addition to silica and oxygen, that are the major constituents of magma, it also contains elements such

Magma Mixing of Magma with Water in Volcanic Activity

Mixing of Magma with Air in Volcanic Activity

Solidification of Magma After Igneous Intrusions and Extrusions

Matter Suspended in Air

Matter Suspended in Water

Biological Activity

Biological Activity

Sedimentation and Deposition Sedimentation and Deposition

Igneous Rocks

Weathering and Transportation

Anatexis i.e. Regeneration of Magma by Fluxing of Pre Existing Rocks Biological Activity

Metamorphism Biodeposits

Biological Activity

Sediments and Soils

Diagenesis

Diagenesis

Weathering and Transportation Anatexis i.e. Regeneration of Magma by Fluxing of Pre Existing Rocks



Sedimentary Rocks



Metamorphism

Metamorphic Rocks

Figure 3.3 Geological cycle.

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Properties of rocks

35

as potassium, sodium, calcium etc. Magmas are accumulated at great depth below the surface of the earth. Due to superheated steam and hot gases, magma always exerts very high pressure on its surroundings. Such high pressure sometimes exceeds the limit of strength of the surrounding mass, resulting in formation of cracks in the surrounding mass. Being in a fluid state, magma flows through such cracks and either intrudes below or extrudes above the surface of the earth. Depending upon the magnitude of the pressure and the rate at which the magma flows, violent volcanic eruptions or slow intrusions take place. When magma cools down, it transforms into different types of rocks. Such rocks, formed after magma intrusion or extrusion, are called igneous rocks. Igneous rocks are found at great depths below all the surface of earth i.e. land or oceans. Various weathering agents e.g. water, wind, sea, river, glacier, biological activities, trees etc. cause continuous disintegration of igneous rocks. Resultant materials, that are in unconsolidated granular form, are called sediments. Sediments are often moved away by transporting agents such as streams, rivers, wind. These sediments eventually get deposited at other locations. When these sediments get cemented and/or consolidated, they turn into a hard and strong sediment mass that is called sedimentary rock. In volcanic activities, sometimes the sedimentary and igneous rocks get heated to very high temperature and are also subjected to very high pressure. Under such conditions the igneous and sedimentary rocks undergo transformation on a large scale to form metamorphic rocks. In some very large scale volcanic activities, all the three kinds of rocks are thrown from the place of their original formation to great depths below the earth’s crust. Inflow of enormous heat from the surroundings increases the temperature of these rocks to such a level that they melt and transform into magma. This process of magma formation is called anatexis.

3.3.1

Igneous rocks

Igneous rocks are composed of minerals that have silica as one of the main constituents. These minerals are quartz, orthoclase, plagioclase, hornblende, mica, augite and olivine. A mineral composition and texture-based genetic classification of igneous rocks is generally followed in geology. According to this basis of classification, there are four classes of igneous rocks, viz. acidic, intermediate, basic and ultrabasic. Acidic rocks have higher quartz contents than other igneous rocks. Quartz crystals can be separately observed in such rocks. The other major constituent of acidic rock is orthoclase. Besides this plagioclase, mica and hornblende are also present. Intermediate rocks may have some quartz and orthoclase but usually less than 10% each. The main constituents of intermediate rocks are plagioclase, mica and hornblende. Basic rocks do not contain any quartz. Their main constituent is plagioclase. Other constituents are hornblende, augite and olivine. Ultrabasic rocks are mostly formed from olivine and have some quantities of augite and plagioclase. Acidic rocks are generally lighter in color, whereas intermediate, basic and ultrabasic rocks are increasingly darker in that order.

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36


The density of acidic rocks is about 2600 kg/m3. For intermediate, basic and ultrabasic rocks it increases gradually to finally reach the level of 3300 kg/m3. The names of some igneous rocks are: granite, basalt, andesite, syenite, dolerite, gabbro, granodiorite, peridotite, ignimbrite, rhyolite, diorite, pumice, ash, eclogite, pitchstone. Figure 3.4 illustrates different classes of igneous rocks mentioned above. Igneous rocks are also classified on the basis of their texture i.e. the relationship between the grains forming the rock. It depends upon grain size, grain shape, degree of crystallinity and grain contact relationship. In this type of classification specific names of the rocks are used. Unfortunately the names of rocks are colloquial in nature and, therefore, there is no unification or standardization in the names of the rocks. Some rocks and their texture-based classification is shown in Table 3.3.

3.3.2

Sedimentary rocks

In the transportation of fragments formed by weathering and erosion, the solid materials get deposited. These deposits are in the form of layers and are called sediments. Sedimentary rocks are formed from these sediments. Therefore all sedimentary rocks are layered. Consolidation, cementation and crystallization turn sediments into sedimentary rocks.

Approximate Percentage of Principal minerals

Acidic

Intermediate

Basic

Ultrabasic

100

75

Quartz Plagioclase

50 Orthoclase

Mica

25

Augite

Olivine

Hornblende 0

Figure 3.4 Composition of igneous rocks. Table 3.3 Classification of igneous rocks. Nature of composition Grain and crystal

Rock type

Coarse Grain, Large Crystals Plutonic Medium Grain, Large Crystals

Acidic

Granite Porphyritic Granite Hypabyssal Granophyre

Medium Grain, Medium Crystals Fine Grained, Medium Crystals Fine Grained, Small Crystals Extrusive

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Intermediate Basic

Ultrabasic

Diorite Porphyrite

Peridotite

Quartz Porphyry Rhyolite Andesite

Gabbro

Dolerite

Basalt Trachyte

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Properties of rocks

37

Often particles in a sediment are of very small size. Due to their predominantly cohesive nature, these particles remain bonded together. Very small pore spaces within these particles are filled with water. When pressure gets exerted on such sediments the water from the pore spaces gets oozed out. In absence of water all the pressure is borne by the particles so the cohesive bonding between particles increases. Eventually a solid hard mass is formed. This mass is called sedimentary rock. The process of oozing out of water and the formation of a solid mass is called consolidation. In certain cases the sediments have particles formed out of very fine materials of a particular nature—mainly calcareous in their chemical composition. When these materials come in contact with water, a chemical reaction takes place and the particles stick together with a very high bonding force. The result is the formation of a hard mass. This process, called cementation, also gives rise to sedimentary rocks. Occasionally, the cementing materials occupy spaces between larger fragments deposited earlier. In such sedimentary rocks two distinct particle sizes are easily noticeable. The process of crystal formation from dissolved minerals often occurs in the shallow parts of the sea or in lakes, in desert areas etc., where evaporation is much higher than precipitation. The sea or lake contains dissolved minerals such as calcium bicarbonate and calcium sulfate. As evaporation takes place, water is lost and the dissolved minerals form crystals. These crystals settle on the bottom of the sea or lake. As evaporation continues, more crystals from and accumulate on the sea or lake floor. After some time interval, they transform into sedimentary rocks. Sedimentary rocks are found near the surface of the earth. They may also be composed of many different types of minerals. Their thickness is usually less than 1 km but can be as high as 10 km. The names of some sedimentary rocks are: breccia, conglomerate, sandstone, limestone, coal, lignite, chalk, dolomite, arkose, greywacke, shale. Sedimentary rocks are classified on the basis of particle size as shown in Table 3.4, or on the basis of their chemical composition as shown in Table 3.5.

3.3.3

Metamorphic rocks

Metamorphic rocks are formed by recrystallization or transformation of already existing rock masses or sediments. The process is called metamorphism. It requires high temperature or pressures or both. Table 3.4 Classification of sedimentary rocks on the basis of their particle size. Class of rock

Grain size

Name of the fragment class

Names of some of the rocks in the class

Rudaceous

Larger than 256 mm

Boulders

Conglomerate (Rounded Fragments) or Breccia (Angular Fragments )

Between 64 and 256 mm

Cobbles

Between 2 and 64 mm

Peebles

Between 1/16 and 2 mm

Sand

Arenaceous

Argillaceous Between ½56 and 1/16 mm Silt Less than ½56 mm

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Clay

Sandstones Siltstones Mudstones, Shale

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38


Table 3.5 Classification of sedimentary rocks on the basis of their chemical composition. Composition

Chemical sediment

Organic sediment

Calcareous Carbonaceous Siliceous Ferruginous Aluminous Phosphatic Saline

Oolithic Limestone, dolomite – Flint, Chert Oolithic Limestone, Laterite Bauxite Phosphorite Rock Salt, Gypsum, Anhydrite

Reef and Shelly Limestone, Chalk Peat, Lignite, Coal Diatomite, Radiolarite Bog Iron Ore – Guano, Bonebeds –

Metamorphism is of four different types as follows. 3.3.3.1

Contact metamorphism

In contact metamorphism very little pressure is involved. The metamorphism is mainly due to heat that transforms matter into metamorphic rocks. During the process of magma intrusion, the heat from magma is given to the adjoining rock. Very high temperatures at the contact between the intrusive body and surrounding older formation causes recrystallization. In this manner a layer of metamorphic rocks is formed around an intrusive body. In the case of small intrusive bodies, the thickness of such layers can be very small but for large intrusive bodies it can even be 2 to 3 km. 3.3.3.2

Regional metamorphism

Changes in the lower crust and mantle exert high pressures and medium temperatures in some parts of the earth’s crust. If the disturbances are very large, existing rocks may be thrown as deep as 10 to 15 km, where pressure and temperatures are very high. These conditions result in metamorphism in that region. The metamorphic rocks formed in this manner are not in layered form but in massive bodies lying within that region which often have dimensions exceeding tens of kilometers. 3.3.3.3

Dynamic metamorphism

Changes in the crust and mantle of the earth can develop cracks in the ground. Very often the layers on one side of the crack are moved up or down relative to others. Such movement generates very high pressure and temperature at the crack. These give rise to metamorphic rocks in thin layers. As the reason of the metamorphism is movement, it is called dynamic metamorphism. 3.3.3.4

Impact metamorphism

When a large meteorite hits the earth’s surface, the impact generates heat and pressure that results in the formation of metamorphic rocks in the immediate vicinity of the impact area.

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Properties of rocks

39

Names of some metamorphic rocks are: slate, schist, gneiss, marble, quartzite, skern, hornfels, amphibolite, mylonite, granulite, epidorite. Classification of metamorphic rocks is less rigidly adhered to than rocks of the other two types. A few rocks are recognized on the basis of their distinct textural and mineralogical features. Composite names are used to achieve greater precision in describing metamorphic rocks. In such composite names, the main constituent of the rock is usually mentioned first, e.g. Garnet-Mica Schist. Table 3.6 describes the generally accepted classes of metamorphic rocks. These classes are based the original rock from which they were formed. Table 3.6 Classification of metamorphic rocks on the basis of original rocks and formation process. Original rock Argillaceous Sediments

General description

Slate

Very fine grained with flat platy cleavages. Often dark in color. A product of low grade metamorphism.

Phyllite

Fine grained with undulating, lustrous cleavage planes. Commonly greenish in color. Exhibits higher metamorphism than slate.

Mica Schist

Medium to coarse grained with rough often puckered cleavage parallel to mica foliation. Middle to high grade of metamorphism.

Gneiss

Medium to coarse grained. Foliation of streaky bands of mica, quartz and feldspar. High grade metamorphism.

Granofels

Fine to coarse but even grained, unfoliated, often quartofeldsparthic. A product of medium to high grade metamorphism.

Arenaceous Sediments

Quartzite

Highly interlocking quartz grains. Usually lacks foliation. Occasionally contains mica crystals. Mostly white in color. High grade of metamorphism.

Calcareous Sediments

Marble

Contains interlocking calcite and dolomite grains. Sometimes contains calcium and silicate minerals. Usually white, gray or buff color. A product of high grade metamorphism.

Basic Igneous Rocks

Antinolite and Chlorite Schist

Greenish rocks with high degree of foliation. Undulated or puckering cleavage. Low to middle grade of metamorphism.

Amphibolite

Medium to coarse grained rocks. Usually found in blocks. Usually foliated and banded. Hornblende and plagioclase are the main constituents. High degree of metamorphism.

Eclogite

Medium to coarse grained rocks. Greenish red color due to red garnet set in a matrix of green pyroxene. Exhibits high grade of metamorphism.

Hornfels

Mostly fine grained rocks. Formed by contact metamorphism. Dark in color. Unfoliated.

Mylonite

Very fine grained flinty rocks. Thinly banded. Usually dark colored. Formed as a result of mechanical crushing and grinding.

Mixed Sediments or Acidic IgneousRocks

Various Sediments

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Metamorphic rock

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40

3.4


FORMATION OF ROCK MASS

Rock mass is formed at or near the surface of the earth by different activities that take place near the surface of the earth. Main activities in this regard are 1 2 3

Volcanic Activities Plutonic Activities Surface Activities

These have been briefly accounted for in earlier parts of this book. The following is a larger elaboration.

3.4.1 Volcanic activities Volcanic activities are those in which magma, which exists at depths within earth under very high pressure and temperature, finds a way to flow above or near the surface of earth. In this activity different types of rock masses are formed. In geological terms they are called structures. Two classes of volcanic structures are known. They are, 1 2

Extrusive Structures Hypabyssal Structures

3.4.1.1

Extrusive structures

Extrusive structures are formed when magma erupts above the ground surface. Such magma, that erupts on the surface of ground, is called lava. Some extrusive structures are lava plateaux, lava cones, hornitos, craters, calderas etc. Descriptions of these extrusive structures are given in Table 3.7. Usually lava cools very rapidly at the surface of the earth. Therefore, rocks found in these structures are fine grained. 3.4.1.2

Hypabyssal structures

When structures are formed by magma intrusion at shallow depths, they are termed hypabyssal structures. Some hypabyssal structures are volcanic plugs, dikes, laccoliths, sills etc. Descriptions of these hypabyssal structures are given in Table 3.8. As the cooling rate of magma in these structures is not very high, the rocks formed in them are medium grained. Occasionally, when the earth surface covering these structures gets eroded, they become visible on the ground surface.

3.4.2

Plutonic structures

Plutonic structures are formed by magma intrusions at great depth below the surface of the earth. Usually these structures are of very large size that ranges from ten to even many hundred kilometers. The rate of cooling of the magma is very slow. More often than not, these structures have layers of rocks. Due to the extremely slow process of

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Properties of rocks

41

Table 3.7 Extrusive structures. Name

Description

Lava Plateau

Lava plateau are formed on the surface of the earth by flow of magma through fissures. Fissures are thin but long and deep seated gaps at the surface of the earth. Magma that flows through these fissures is low in viscosity and flows over a very large surface. It cools rapidly and rocks formed in the process are basalt.

Lava Cone

Lava cones are formed when lava flows to the surface of earth through medium sized pipelike structures. Usually such lava has somewhat intermediate viscosity and, therefore, spreads less easily on the surface of the earth. In the process a hillock with gentle slope on all the sides is formed.

Hornito

Hornitos are formed when highly viscous lava flows to the surface of earth through large structures. Hornitos are tall but small diameter structures with steep slopes on the sides.

Crater

Craters are depressions in the ground. They are formed at the mouth of a volcano. Craters can also be formed when a meteorite hits the earth surface.

Caldera

Caldera are very large depressions on the surface of the earth. They are formed by coalescence of several small craters or repeated explosions of a volcano at adjacent places or ground subsidence of a large portion of earth surface in a huge hollow formed just below the surface of the earth during a violent volcanic eruption. Calderas can be very large in diameter – often tens of kilometers.

Table 3.8 Hypabyssal structures. Name

Description

Volcanic Plug

During volcanic intrusions magma often flows to the ground surface through circular feed channels. After the volcanic activity comes to an end such channels remain filled with solidified magma. Later the surrounding structure may get eroded, leaving behind the magma channel in vertical or near vertical position. This is called a volcanic plug.

Dike

Often magma flow cuts across the bedding planes of the older formations and forms sheet like structures. These are called dikes. Many different types of dikes – based on their shapes – are known e.g. parallel dikes, radial dikes, ring dikes etc.

Laccolith

Intrusive magma often takes the shape of a dome having flat underlying surface and arch like top surface. In appearance this structure looks like a mushroom. These are called laccolith. Laccoliths are called composite when they are formed from two different sources of magma. Similarly two or more laccoliths formed from the same magma source are called multiple laccoliths. Cedar tree laccoliths are the laccoliths lying one above another and resemble to the shape of a cedar tree. All these laccoliths are formed from the same magma source.

Sill

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When magma intrudes in between two bedding planes and forms a sheet like structure, of nearly same thickness and negligible concavity, it is known as a sill. Sills, like laccoliths, can be multiple or composite.

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42


cooling, the fluid state of magma lasts for a long time. Acidic magma, which has lower density, floats at the top and the basic magma settles down to greater depth. In this condition the upper rocks are coarse grained acidic—granites—and the lower rocks are coarse grained basic—gabbro. Various plutonic structures are batholiths, stocks, lopoliths, phacoliths etc. Descriptions of these extrusive structures are given in Table 3.9. Table 3.9 Plutonic structures. Name

Description

Batholith

Batholiths are gigantic structures having a cross sectional area of more than 100 km2. The structure goes on enlarging horizontally with increasing depth. There is no apparent bottom to the batholith. It can be considered as a part of inner crust. Stocks are similar to batholiths but much smaller in their physical dimensions. Structures smaller than 100 km2 area are classified as stocks. When the magma intrusions take a saucer like shape with the upper and lower concave surfaces as seen from the top, the resulting structure is called a lopolith. By their nature the lopoliths are formed by magma intrusions between two concave layers. The path of magma is somewhat narrow. Therefore, the type of magma that intrudes in lopoliths is basic in nature. Naturally, rocks formed in the lopoliths are coarse grained basic rocks. When magma intrudes in between two convex beds, as seen from the top, the resultant structure is called phacolith. In other words phacoliths are similar to lopoliths but inverted upside down.

Stock Lopolith

Phacolith

Table 3.10 Surface structures. Name

Description

Unconformity During the deposition of the sediments that give rise to the beds, there is a time interval during which no deposition takes place. Therefore, a surface formed during this time interval separates the old beds from new beds. On many occasions, due to change in the mode of deposition or change in the type of sediment, such surface is very distinct and is called unconformity. Angular, Parallel, Non Depositional and Hetrolithic are the four different types of unconformities. Joints During tectonic activities different types of stresses are developed on the beds. Rocks are capable of withstanding high compressive stresses but are weak in tension. Due to this, cracks i.e. joints develop at various places in the rock mass. The most important property of joints is that there is no apparent movement on two sides of the joint with respect to each other. A series of parallel joints is called as a set of joints. Joints are classified on the basis of mode in which they were formed. Three classes of joints viz. Shrinkage, Tectonic and Sheet are generally recognized. Faults Fractures in formation beds, along which there has been observable movement of the formation on each of the sides of the fracture plane, are termed as faults. Normal, Hinge, Pivot, Tear and Reverse are the five types of faults Folds During tectonic movements the plane beds are often converted into a wave form. Such structures are called folds. The trough like portion of a fold is termed as syncline, whereas, the crest like portion is named as anticline. There are nearly ten different types of folds.

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3.4.3

43

Surface activities

Denudation is a general term used to refer to the processes that alter the surface of the earth. It therefore consists of weathering, transportation and erosion. Weathering means natural disintegration and/or decomposition of the rock mass. Transportation means shifting of the disintegrated or decomposed material away from the place at which they were initially formed. While the broken materials are transported, they come in physical contact with other fresh uneroded formations. The friction developed between the two causes weathering of the fresh formation. This process is termed erosion. As all the above processes take place on the surface of the earth, they are called surface processes. Almost all the surface processes result in deposition of the material on the earth surface in layers and the structures formed in the process are unconformities, joints, faults, folds. A description of different surface structures is given in Table 3.10.

3.5

SOILS, ROCKS AND THEIR PROFILE

The term soil originates from the farming industry. In civil engineering it is defined as an unconsolidated, uncemented mineral mass containing air, water and other impurities of organic nature. Soil is also a product of weathering. It usually lies at or near the surface of the ground. When an excavation below the surface of ground is carried out in the form of a trench or vertical cut, different layers having different constituents and colors are visible on the wall. The composite structure of such layers is called a soil profile. Soil profiles depend upon many factors. Different soil profiles are seen in different regions. Some typical soil profiles are shown in Figure 3.5. In modern nomenclature the ‘O’ horizon is the topmost. It consists of fresh litter in its upper, and decomposed organic matter in its lower, part. It retains soil moisture. The ‘A’ horizon, having a dark blackish brown color, underlies the ‘O’ horizon and consists of a mixture of minerals and organic material. Leaching, i.e. chemical decomposition, removes minerals like Ca, Fe as well as decomposed organic matter to lower horizons. The ‘E’ horizon underlies the ‘A’ horizon. Usually it is a light-colored horizon with ongoing activity of leaching, or the removal of clay particles, organic matter, and/or oxides of iron and aluminum. In some forest regions the ‘E’ horizon has a high concentration of quartz giving the horizon an ash-gray appearance. Beneath the ‘E’ horizon lies the ‘B’ horizon. Most of the fine matter leached in the upper horizons is accumulated in the ‘B’ horizon. The accumulation of fine material leads to the creation of a dense layer in the soil. In some soils the ‘B’ horizon is enriched with calcium carbonate in the form of nodules or as a layer. This occurs when the carbonate precipitates out from capillary action of soil or water moving downward. The ‘C’ horizon is the next deeper horizon. It mainly consists of parent rock in weathered form. The lowermost horizon is the ‘D’ horizon and represents the parent rock in intact form. In some places some horizons are absent.

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Tundra Profile

A - Peat

Terra Rossa Profile A - Red Clay

O (Contains organic matter) A (Rich in humus)

B - Clay

B - Limestone

B (Rich in Mineral salts)

C (Subsoil and weathered rock)

Laterite Profile

Desert Profile A - Sand (No Dunes) A - Clay

B - Concentration Rich in salt D (Solid rock)

C - Leached Zone D D - Solid Rock

Figure 3.5 Soil profile under different environments.

3.6

OCCURRENCE OF MINERALS

A rock mass is formed by the assemblage of several minerals. Some of these minerals are of great economic importance because they are useful in many ways e.g. for extraction of metals, for use in different manufacturing processes, for medical use, for ornamental use etc. Mineral deposit is a term used for an assemblage of minerals where one or more, but not all, the minerals in the assemblage are of value. When a granite formation is used for production of tiles for house flooring, kitchen platform etc., it is not called a mineral deposit since it does not involve any specific mineral. But when a granite formation is excavated for obtaining pyrite contained in it, the granite formation is called mineral deposit. An ore body is the specific portion in a mineral deposit that fully or largely contains the mineral of economic importance. The remaining portion of the mineral deposit is considered as waste and is usually called overburden because it lies above the ore body. It has to be removed to gain access to the ore body. When an ore body contained in a mineral deposit has to be accessed it becomes essential to break the overburden or the surrounding rock mass and move it away. The ore itself has to be crushed before subjecting it to further chemical processes. All such activities together are called mining. When the rock mass is excavated as a whole, the activity is called quarrying. Examples of this are the rock mass excavation in broken form e.g. large stone pieces

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45

or crushed stone required in the construction industry or in dimensional block form for cutting tiles for house flooring etc. In nature, mineral deposits occur in many different forms. During the formation of igneous rocks, the magma often contains small quantities of minerals of copper, lead, zinc etc. These minerals do not find a place in the main rock mass that contains quartz, feldspar, mica etc. Therefore, these minerals get separated into ore bodies. Water superheated to very high temperatures is very active and can dissolve many economically important minerals to form a hydrothermal fluid. When an igneous rock mass forms, the hydrothermal fluid usually accumulates at the top of the rock mass as it has low density. After a certain time interval the water itself gets evaporated and the compounds crystallize to form a mineral deposit. On many occasions, the hydrothermal fluid mentioned above enters and spreads through cavities such as joints, fissures etc. This gives rise to vein minerals in the form of large crystals. Tourmaline, topaz, cassiterite etc. are some of the important vein minerals. In certain cases the hydrothermal fluid intrudes the existing rock bodies and forms a mineral deposit called a replacement mineral deposit. Minerals of lead and zinc are found in such deposits. The process of magmatic differentiation during the formation of intrusive and extrusive igneous bodies gives rise to pyritic deposits containing iron and sulfur. With the magma intrusion, hot gaseous substances flow in the magma mass. As the magma cools these gaseous substances deposit certain minerals, which till then are in a volatile state. The process of formation of such deposits is called pneumatolysis. The most important mineral contained in such deposits is Wolframite. At times, due to different degrees of metamorphism, not all the minerals change their form. In the process some minerals become liquid in their state and easily flow away. What remains is a skeleton of fibrous silicates commonly known as Asbestos. In certain instances, the formations in which metamorphism takes place involve carbon in the form of coal or carbonaceous gases. These get converted into graphite or diamonds. Mineral deposits are also associated with sedimentary rocks. The flow of water in a river may carry minerals of low specific gravity due to the limited velocity of water. Such minerals are mostly silicates. What remains in their place is a mineral mass of high specific gravity. These are called placer deposits. Gold, platinum, antimony etc. are associated with such placer deposits. Sea water contains many dissolved minerals. The proportion of these is about 31 parts per thousand. When sea water enters certain depressions at the time of high tide, it may remain entrapped in the depression for a long time. Evaporation of the water leads to formation of a thin layer of mineral deposit. When the process of such evaporation gets repeated with the next equal high tide the thickness of mineral deposit grows and eventually the deposit acquires a thickness of considerable magnitude. Such deposits are called evaporites. Potash, magnesium, gypsum etc. are associated with such deposits. Laterite is also a form of sedimentary mineral deposit. Laterite deposits are formed in tropical regions in the rainy season. Sediments formed are transported to flat surfaces where they penetrate into the ground. In the dry season some minerals

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containing the leached ions are drawn to the surface of the ground by capillary action. In the next rainy season these minerals are washed away and fresh sediments are redeposited. In this manner higher density minerals remain deep seated and form a large mineral mass. Iron ore or bauxite are the minerals found in such deposits.

3.7

ROCK SPECIMEN PROPERTIES

The objective of blasthole drilling is to drill a hole in the rock at the fastest possible penetration rate, while giving maximum life to the blasthole drill and components like drill bits, drillstring accessories etc. The progress of a blasthole is within a very small cylindrical path in a rock mass. Therefore, when it comes to blasthole drilling, the characteristics of the whole rock mass are less important as compared to the properties of a rock sample or specimen. In the rotary blasthole drilling process, the intact rock at the bottom of the hole is crushed or scraped and chips formed are blown out along with circulating compressed air. If any discontinuities, like joint, fracture etc. are perpendicular to the blasthole alignment they usually help the blasthole drilling process and result in faster penetration rates. However, if they are at an angle with the blasthole alignment and if the strength properties of the rock on two sides of the cracks or joints are far different, the blasthole can deviate from its intended alignment to an unacceptable level. From the standpoint of blasthole drilling, the most important property of intact rock is drillability, and from the viewpoint of blasting it is blastability. Drillability is a complex combination of several factors amongst which engineering properties of the rock specimen play a very important role. Some of these are density, compressive strength, tensile strength, hardness, toughness, brittleness, coefficient of internal friction, and abrasiveness. Drillability is treated in separate chapters in this book but the other engineering properties are briefly dealt with in the following subsections. Since intact rock specimens from different depths are usually obtained in the forms of cores in diamond core drilling, these cylindrical cores are used for most of the tests in a rock testing laboratory. The cores are carefully prepared before testing so that they have appropriate length and diameter. The ends of the test specimens are also carefully cut so the circular surface is exactly perpendicular to the core axis and is in a plane so that the specimen is subjected to uniform load and not point loads. When hard rock cores are to be shaped, diamond tools are used for the shaping. To prevent diamonds from burning, water is used as a coolant. The specimens so prepared are dried before testing. For shaping cores of soft rock that contain clay and other constituents that may be affected by water, tungsten carbide tools are used. Such tools do not require water as a coolant. In order to get truly a representative value of the property, many tests are carried out and the mean value of the property is taken as the correct representation. A few formulae are given in the following elaboration. As these formulae are only for the purpose of understanding the relation of various quantities, no units of measurements have been given for the quantities. In any case these formulae are unified and will give desired results when consistent units for all the terms and appropriate values of the constants in the formulae are used.

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3.7.1

47

Density

Density of rocks, usually measured in terms of kg/m3 or lb/ft3, is not so important from the viewpoint of formation fracturing, but is of great importance when it comes to flushing. As per Stoke’s law, discussed in one of the later chapters, chips of the rocks with low density, such as shale, coal etc., are bailed out from the blasthole very easily but chips of the rocks with high density such as iron ore, uranium ore etc., are bailed out with relatively great difficulty. Fortunately, the density of rocks is usually within a small range. When the bailing velocity of compressed air is properly selected by use of an appropriate compressor, the penetration rate does not reduce on account of the density of rocks. Densities of many different rocks are given in one of the appendices at the end of the book.

3.7.2

Compressive strength

In plain words, compressive strength means ability to resist compression. When a cylindrical rock sample is subjected to increasing axial compressive force as shown in Figure 3.6, its axial length goes on reducing and at the same time its diameter goes on increasing. Stress exerted on rock sample is calculated as σ = F/A = F/(π ∗ d2/4) The strain that results from the test is calculated as ε = ΔL/L Explanation of the symbols F, d, L is contained in Figure 3.6 itself. At certain values of compressive stress the rock sample fails. Thereafter the compressive strain keeps on increasing even when the axial load is not increased, or in fact even when it is decreased, resulting in unchanged or reduced compressive stress. Eventually, the rock sample crumbles into pieces. The maximum value of stress to which the rock specimen is subjected in such a test process, is called compressive strength. In most of such tests

ΔL

Compression Force F

Original Shape of the Rock Sample is in Gray Color

Top Plate Averaged Shape After Compression L Likely Failure Plane ΔL1

Bottom Plate

Figure 3.6 Uniaxial compression test.

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the rock sample is subjected to compressive pressure only in the axial direction and no pressure is applied on its cylindrical side except the atmospheric pressure. Therefore, the strength is termed as uniaxial compressive strength and abbreviated as UCS. Sometimes UCS is also taken as an acronym of unconfined compressive strength. UCS is measured in the laboratory in a machine that exerts force on the material sample as shown in Figure 3.7. The machine is called a Universal Testing Machine. It consists of special servo-controlled mechanisms for measurements. Rock samples used for the test are usually cylindrical in shape and the circular surfaces of the sample are very carefully polished to avoid point loading. The force exerted on the sample is increased at a certain rate and the material sample eventually fails. Brittle materials, like rock, get crushed into pieces. The compressive strength of the material is calculated by dividing the breaking force by the cross sectional area of the sample. UCS of a dry rock sample is always more than that of a wet rock sample. For some sedimentary rocks the difference is very high. The commonly adopted procedure for determination of UCS is as per American Standard ASTM D2938. Stress-strain curves plotted from the observations of unconfined compression tests on samples of quartzite, hornfels and marble are shown in Figure 3.8. Rock is said to be elastic when the strain is linearly proportional to stress, whereas when the relationship is non linear the rock is called plastic. Quartzite shows perfectly elastic behavior throughout up to a point B, and thereafter shows slightly plastic behavior up to the ultimate compressive strength at point C.

Figure 3.7 Rock core sample being subjected to UCS test.

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Properties of rocks

300 B

Stress in MPa

250 200

B1

49

C Quartzite

C1 Hornfels

150 100

2

C2

50

3

Marble

0

10

20 30 40 Strain in mm ×106

50

60

Figure 3.8 Stress strain curves for some rocks.

Since the length of the curve OB is several times longer than the length BC and even the zone BC is only slightly plastic, it is safe to say that quartzite is an elastic rock. Hornfels shows perfectly elastic behavior up to point B1 and thereafter significantly plastic behavior up to the ultimate compressive strength at point C1. Hornfels is considered an elastoplastic rock because the length of plastic zone B1C1 is of significant magnitude as compared to elastic zone OB1 and the plasticity in zone B1C1 is also of considerable magnitude. In the case of marble there is no elastic zone at all. Marble, therefore, is considered a plastic rock. When rock is subjected to compression, failure occurs due to shear stress at an inclined plane as shown in Figure 3.6. Some important properties of rock determined from the readings of uniaxial compression tests are the ‘Modulus of Elasticity’, ‘Modulus of Plasticity’ and ‘Poisson’s Ratio’. If rock shows elastic behavior then the modulus of elasticity within elastic zone is given by the equation E = (σ1 − 0)/(ΔL1 − 0) Modulus of plasticity in the plastic zone is calculated by the equation EPL = (σcom − σ1)/(ΔLcom − ΔL1) The meaning of the symbols used in the above two equation can be easily understood from Figure 3.9. Poisson’s ratio is the ratio of lateral strain to axial strain. In case of Figure 3.6 it equals μ = ΔL1/ΔL

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Stress in MPa

50

C1

σcom B1

σ1

0 0

ΔL1

Strain

ΔLcom

Figure 3.9 Stress strain curves for elastoplastic rocks.

In all the compression or tension tests carried on the rocks, the rate at which stress is applied has a great bearing on the ultimate compressive or tensile strength. All rocks show very high strength when the compressive force is applied rapidly. UCS of rocks has a reasonably good correlation with point load index and shore hardness. Often rock specimens (not rock masses) are classified on the basis of their UCS as indicated by Table 3.11. UCS of some commonly occurring rocks is given in Table 3.12 and also in Appendix 14. A special test, called triaxial compression test, is also sometimes carried out on soft rock samples. In such tests the rock sample is subjected to pressure on the cylindrical side of the sample. The value of compressive strength obtained in triaxial compression tests is higher than the UCS. The triaxial compression test has a far better resemblance of in-situ behavior of the rock, because in actual practice a portion of rock in the alignment of the blasthole is always surrounded by an adjacent portion of rock mass. Such rock mass invariably exerts horizontal pressure on the portion of rock in blasthole alignment. The commonly followed procedure for determination of CS by triaxial compression test is as per American Standard ASTM D2664.

3.7.3 Tensile strength The concept of tensile strength is almost identical to that of compressive strength, but the direction of the axial force acting upon the sample is exactly opposite. Here, as the sample is subjected to increasing tensile force, the diameter of the rock sample goes on reducing and the axial length goes on increasing. Tensile stress and tensile strain are observed in the tests in the same manner as those in compressive stress and strain. At a certain value of tensile stress, the rock sample fails. At this point the rock sample usually breaks apart in the case of elastic rocks. In some rocks, that show high plasticity, the tensile strain keeps on increasing even when the axial force is not increased. The value of the stress at such point is called tensile strength.

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Table 3.11 Rock classification on the basis of unconfined compressive strength. Uniaxial compressive strength in MPa

Point load Schmidt index in hardness by MPa L – hammer

Method of strength in the field

R5 – Extremely >250 Strong

>10

50–60

R4 – Very Strong

100–250

4–10

40–50

R3 – Strong

50–100

2–4

30–40

R2 – Medium Strong

25–50

1–2

15–30

R1 – Weak

5–25

–

<15

R0 – Very Weak

1–5

–

–

Extremely Weak

0.25–1

–

–

Rock material only chipped under repeated hammer blows Requires many blows of a geological hammer to break intact rock specimens Hand held specimens broken by a single blow of a geological hammer Firm blow with geological pick indents rock to 5 mm, knife just scrapes surface Knife cuts material but too hard to shape into triaxial specimens Material crumbles under firm blows of geological pick, can be scraped with knife Indented by thumbnail

Class

Examples of rocks in the class Fresh Basalt, Chert, Diabase, Gneiss, Granite, Quartzite Amphibolite, Sandstone, Basalt, Gabbro, Gneiss, Granodiorite Limestone, Marble, Phyllite, Sandstone, Schist, Shale Claystone, Coal, Concrete, Schist, Shale, Siltstone

Chalk, Rocksalt, Potash

Highly Weathered or Altered Rock

Clay Gouge

Unlike metals, where the compressive strength and tensile strength have comparable values, for most rocks the tensile strength has a much lower value than that of compressive strength. When the ratio of compressive to tensile strength is higher, the rock is more brittle. Tensile strength σTP of a rock sample is determined indirectly by a test, often called the ‘Brazilian Tensile Strength’ test and abbreviated as ITS. In ITS a rock sample having core diameter d and length L is subjected to a load as shown in Figure 3.10. The length L of the rock sample used for the test must not be more than 2 times its diameter d. The diameter d must also be larger than 22 mm. If the rock sample breaks at point load FP then the point load tensile strength is taken as σTP = 2 ∗ FP/(π ∗ L ∗ d)

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Table 3.12 Range of UCS of some common rocks. UCS range Rock name

App. MPa

psi

Diorite Gabbro Granite Andesite Basalt Rhyolite Trachyte Dolomite Limestone Conglomerate Sandstone Shale Gneiss Marble Quartzite Schist Slate Taconite

170–300 260–350 200–350 300–400 250–400 120–130 330–350 150–170 120–130 140–150 160–260 60–75 140–300 100–200 160–220 60–400 150–160 300–600

24500–43500 37500–50500 29000–50500 50500–58000 36000–58000 17500–19000 47500–50500 21500–24500 17500–18500 20500–21500 23000–37500 8500–11000 20500–43500 14500–29000 23000–32900 8500–58000 21500–23500 24500–87000

Measurements in the ITS test are done by an electronic servo-controlled machine. The procedure for the ITS test is given in ASTM D3967. Particular attention must be given to the placement of the sample in the ITS test when the rock samples are layered. In such rocks the tensile strength across the layers can be far higher than that along the layers.

3.7.4

Hardness

Hardness means resistance of a substance to penetration by another hard substance. Rock is composed of many minerals. Every mineral has a specific value of hardness, but due to differences in the percentages of minerals, the hardness of the same type of rocks differs within a certain range. Different scales for measuring hardness of materials exist. The basis on which these scales are formulated also differ. Appendix 8 at the end of this book presents a table for converting hardness from one scale to the other. Hardness is measured by many different types of tests, each using specific types of testing machines. Therefore, the value of hardness pertains to a specific hardness scale. For rock samples the most commonly used scales and tests for hardness are: Shore Rebound Hardness Vickers Indention Hardness Knoop Indention Hardness.

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53

Figure 3.10 Brazilian tensile strength test setup.

3.7.4.1

Shore hardness

The apparatus used for this test was originally made by the Shore Instrument and Manufacturing Company. The instrument is called a scleroscope, hence the hardness is often called scleroscope hardness. The apparatus consists of a steel rod i.e. ‘Hammer’

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having a diamond tip at one of its ends. This end is allowed to fall freely under gravity from a fixed height onto the polished surface of the material to be tested. After the impact the hammer rebounds freely. From several readings of the rebound height an average height of rebound is calculated. The hardness of the rock sample is determined by correlating the average rebound height with a chart made for the purpose. 3.7.4.2

Vickers hardness

In the Vickers hardness test a square-shaped diamond pyramid mounted on the tip of a rod is pushed into the polished surface of the rock sample with pre-adjusted force. The area of the indention made by the diamond is measured through a microscope. The hardness of the rock is then determined by correlating the measured area with a chart. Values of hardness of some important minerals are given in Table 3.13. Range of Vickers hardness values for some rocks are given in Table 3.14.

3.7.5 Toughness Toughness means resistance of the material to propagation of cracks. Tough rocks like basalt are difficult to break but may not have very high hardness. Toughness y, of rocks is calculated by the equation, y = (E/EPL) ∗ σcom where E = Modulus of elasticity EPL = Modulus of plasticity σcom = Unconfined compressive strength

3.7.6

Brittleness

Brittleness refers to relative ease of the material in its disintegration under impact loads. Highly brittle materials like glass break suddenly. Brittleness of rocks is evaluated by a test called the Impact Test. For the impact test, a sample of crushed rock is first sieved through a sieve with square openings of 16 mm. Pieces retained on the sieve are rejected. Pieces passing through the 16 mm sieve are further sieved through a sieve with square openings of 11.2 mm. From the aggregate retained on the 11.2 mm sieve about 0.5 kg sample is selected for the test. From this sample rock pieces without cracks are used for the test. The exact weight of the sample W1 is recorded. The sample is then put into a steel container having very thick walls. A heavy steel piece is then placed on the sample and it is subjected to impacts by a 14 kg weight falling through 25 cm height. The set up is shown in Figure 3.11. After 20 blows of the falling weight on the rock sample, the sample is carefully removed and again sieved through a sieve having 11.2 mm square opening. The weight of the sample retained on the sieve W2 is again recorded. The S20 value, which is an indicator of brittleness of rock, is calculated as S20 = 100 ∗ (W1 − W2)/W1

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Table 3.13 Hardness values for some minerals. Hardness values on different scales Mineral name

Chemical composition

Vickers

Knoop

Rosiwal

Corundum Quartz Garnet Olivine Hematite Pyrite Plagioclase Diopside Magnetite Orthoclase Augite Ilmenite Hiperstene Hornblende Chromite Apatite Dolomite Pyrrhotite Fluorite Pentlandite Sphalerite Chalcopyrite Serpentine Anhydrite Calcite Biotite Galena Chalcocite Chlorite Gypsum Talc Halite Sylvite

Al2O3 SiO2 Silicates of Fe-Mg-Al-Mn-Ca-Cr (Mg, Fe)2SiO4 Fe2O3 FeS2 (Na, Ca)(Al,Si) AlSi2O8 CaMdSi2O6 Fe3O4 KAlSi3O8 Ca(Mg, Fe, Al)(Al, Si)2O6 FeTiO3 (Mg, Fe)SiO3 NaCa2(Mg, Fe, Al)5(Al, Si)8O22 (Mg, Fe)Cr2O4 Ca5(PO4)3(F, Cl, OH) CaMg(CO3)2 Fe1-XS CaF2 (Fe, Ni)9S8 (Zn, Fe)S CuFeS2 Mg6Si4O10(OH)8 CaSO4 CaCO3 K(Mg, Fe)3(AlSi3O10)(OH)2 PbS Cu2S (Mg, Fe, Al)6(Al, Si)4O10(OH)8 CaSO4 · 2H2O Mg3Si4O10(OH)2 NaCl KCl

2300 1060 1060 980 925 800 800 800 730 730 640 625 600 600 600 550 365 310 265 220 200 195 175 160 125 110 85 65 50 50 20 17 10

1700 790

1000 141

3.7.7

395

7.3

85

4.08

32 12

0.85 0.82

Coefficient of internal friction

The coefficient of internal friction of a rock is an indicator of resistance to the sliding of two rock particles or pieces with respect to each other in a state where their surfaces are in contact with each other.

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Table 3.14 Vicker’s hardness for some rocks. Rock type

Vickers hardness

Amphibolite Andesite Anorthosite Basalt Black Shale Chromite Cooper Ores Diabase/Dolerite Diorite Epidotite Gabbro Gneiss Granite/Granite Gneiss Granodiorite Granulite/Leptite Green Schist Greenstone Hornfels Limestone Marble Metadiabase Metagabbro Micagneiss Micaschist Nickel Ores Norite Porphyrite Pyrite Ores Phyllite Quartzite Rhyolite Sandstone Serpentinite Shale/Silstone Skarn Sphalite Ores Tonalite Tuffite

500–750 550–775 600–800 450–750 300–525 400–610 350–775 525–825 525–775 800–850 525–775 650–925 725–925 725–925 725–925 625–750 525–625 600–825 125–350 125–250 500–750 450–775 500–825 375–750 300–550 575–725 550–850 500–1450 400–700 900–1060 775–925 550–1060 100–300 200–750 450–750 200–850 725–925 150–850

When rock breaks, the particles formed in the process always tend to slide along each other until they reach a stable condition. In situations such as calculations of the stability of a slope and calculations of the volume of the heaps formed by dumped material etc., the coefficient of internal friction gains importance.

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57

14 kg Weight

250 mm Fall

Guide

Aggregate

Container

Bottom Plate

Figure 3.11 Impact test for brittleness of rocks.

3.7.8

Swell factor

When intact rock, having volume V is broken, and the gross volume V1 of such broken pieces is measured, it is observed that V1 is larger than V. This increase in volume, called ‘Swelling’, results from the voids created by the haphazard placement of the broken pieces. The swell factor, SF, is measured in % and is given by the formula: SF = 100 ∗ (V1 − V)/V This factor is very important in calculations for the transportation of the blasted rock because blasted rock is always in swollen conditions.

3.7.9

Abrasiveness

Abrasiveness of a rock means the property of the rock to wear out (by friction) metal tools used for drilling or cutting. It is measured by any of the three tests viz. 1 2 3

Cerchar Abrasiveness Test Miniature Drill Test Bit Ware Index Test These tests are described in following paragraphs.

3.7.9.1

Cerchar abrasiveness test

The tool used for testing abrasiveness of rock is made from steel having a compressive strength of 200 kgf/mm2. The tip of the pin has a conical shape with cone angle

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of 90°. In the test this tip is placed on a carefully prepared, polished, plane surface of the rock sample with a piercing force of 7 kgf. The rock sample is then made to move over a distance of 10 mm in one second. The diameter of wear flat created at the tip is carefully measured in mm. Coating the pin tips with machinist’s blue dye prior to testing, makes the wear flat more visible. This procedure is repeated five times on a fresh surface of the rock sample that is undisturbed by the first scratch. A new pin is used every time. The whole procedure is repeated on 3 to 5 samples of the same rock. Cerchar Abrasivity Index, CAI, is then calculated by the equation: CAI = (10 ∗ Σdwearflat)/N where N = The number of observations d = Wearflat diameter in micrometers A schematic of the apparatus used for the Cerchar Abrasiveness Test is shown in Figure 3.12. CAI values for some rocks are given in Table 3.15. Tests carried out on many rock samples have shown that for most of the rocks having a CAI more than 0.7, there exists a correlation between CAI and Vickers hardness as indicated by the following equation. VH = 145 ∗ CAI 3.7.9.2

Miniature drill test

The miniature drill test actually simulates drag bit drilling on the rock sample.

Pull Handle

7 Kg Weight Hinge Steel Pin With 90° Apex Angle Limiting Screw

Rock Sample

Bench Vise Limiting Block

Figure 3.12 Apparatus for cerchar abrasiveness test.

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Properties of rocks

59

Table 3.15 CAI for some common rocks. Rock type

C AI value

Basalt Diabase Andesite Diorite/Syenite Granite Limestone Sandstone1 Sandstone2 Phyllite Mica Schist Mica Gneiss Felsic Gneiss Amphibolite Quartzite

1.7–5.2 3.8–5.4 1.8–3.5 3.0–5.6 3.7–6.2 0.1–2.4 0.1–2.6 2.3–6.2 1.3–4.3 1.8–5.0 1.8–5.0 3.7–6.3 2.8–3.7 4.8–7.3

The test setup is as shown in Figure 3.13. The rock sample used for the miniature drill test is of the size and shape of a brick commonly used in building construction. The sample is firmly held by clamps on the upper head of the testing machine as shown in the figure. The upper head assembly can freely move up and down in the slider tubes when its holding brake is released. By suitably adding weight on this head, the total weight of the head together with the brick sample, is made equal to 20 kg. Thus, a constant bit weight of 20 kg is used for drilling. A drill bit made from tungsten carbide, with a tip width of 8.5 mm and tip angle of 110°, is used in the test for drilling into the sample. The drill bit is fixed in the low speed rotation motor, adjusted to run between 175 to 200 rpm speed, with its tip pointing upwards as shown. Since the bit drills into the sample in an upward direction, the cuttings formed in the process of drilling fall down under gravity without any need for a flushing device. When the drill head assembly is lowered, the tip of the drill bit touches rock surface. At this point the rotation electric motor is rotated at a constant low speed to create a hole in the rock sample. A revolution counter on the test apparatus continuously indicates the number of revolutions made by the drill bit. When the drill bit makes 200 revolutions the drilling is stopped. The depth of the hole is measured correctly to 1/10th of a millimeter with the help of a vernier caliper. The SJ value for the rock equals 10 times the average depth found in the series of eight tests. To get a realistic reading of the SJ value for the rock, care must be taken to drill holes parallel to the rock foliation. In the case of coarse grained rocks, holes have to be drilled in all the grain types.

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60


8.5 mm 110 g

Rock Sample

Chuck Guide

Tungsten carbide drill bit

Guide 175−200 RPM

Figure 3.13 Miniature drill test.

Figure 3.14 Abrasion value test.

3.7.9.3

Bit wear test

The set up of the test used for evaluating the abrasion value of a rock is shown in Figure 3.14. In this test a piece of tungsten carbide, having the dimensions as shown in the magnified view of the bit in Figure 3.14, and having a Vickers hardness value of 1300 is used. The weight of the piece is carefully measured before starting the test. Crushed powder of the rock to be tested is continuously fed on a rotating disc at a feed rate of 80 g/min through a vibrating feeder. The drill bit is pressed on the powder with a force of 10 kg. The rock powder is sucked by a vacuum sucker after it has passed underneath the bit, so that any tungsten carbide powder separated from the bit and mixed with the rock powder is not used for further grinding of the bit in the next revolution. The disc is rotated at 20 rpm for 5 minutes. The abrasion value of the rock in the test is the loss of bit weight in mg. The test is repeated 8 times and the average weight loss is taken as the abrasion value for the rock.

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Chapter 4

Brief history of rotary blasthole drilling

4.1

INTRODUCTION

Blasthole drilling, as we see it today, has developed through need. For centuries past, the needs have been ever increasing, and in the future also the trend will continue to increase. The existence of water and minerals in the ground was known even in the prehistoric era. Wells were dug for extraction of water that percolated into them. When pumping water from small diameter holes became possible, the need for drilling deep holes to access the deep-seated water sources was felt. The percussion drilling method was used for water wells for over two thousand years till the recent past. Even metals have been known to the human race for more than seven thousand years. Ore, needed for making metals, was excavated from the surface mineral deposits by hand digging methods as late as the beginning of the seventeenth century. Georgius Agricola, the first of the great engineers of modern times, wrote his treatise ‘De Re Metallica’. It was published it in 1556. The work extensively dealt with the mining practice of those days and contained many woodcut illustrations. Many devices, such as windlass, hoist, horse power etc., are shown in the book, but there are no references to drilling holes or blasting techniques. As the book was meant to cover mining, it can be inferred that drilling was not carried out in the mining practice of those days.

4.2

ERA OF SHOT HOLE DRILLING

A shot hole means a blasthole of small diameter. Gunpowder was known to the Chinese as early as the ninth century. The technique of manufacturing gunpowder became known to the Europeans only in the eleventh century. ‘De Mirabili Potestate Artis et Nature’ written by Roger Bacon, published in 1242, gives details about the destructive power of gunpowder and a formula for making it. For many centuries the use of gun powder was restricted to war equipment like bombs, guns, cannons etc. The concept of drilling holes in rock, filling them with gunpowder and blasting them was first proposed in 1617 by Martin Weigel, a mining superintendent at Frieberg, Germany. Casper Windt is said to have used this concept successfully in 1623 at a German town Schemitz.

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62


One of the earliest articles that appeared on holes for the purpose of blasting was authored by Sir Richard Morey in the Philosophical Transactions of Royal Society in the year 1665. He described Monsieur Du Son’s method of rock blasting, along with the illustration of piercer – a drilling rod etc. His article reads, “… There is a tool of iron well steeled at the end, which cuts the rock 20 or 22 inches long or more and some 2.25 inch diameter at the steeled end the rest being somewhat more slender. The angle at the end is so shaped as to make it most apt to pierce the rock. The tool is to be held by hand in the place of the hole … so one may manage hammer while the other holds the tool … after the stroke of the hammer he that holds the piercer, is to turn it a little on its point so that edges at the point may strike all upon new surface … till the hole is 20 to 22 inch deep …” Sir Richard Morey further goes on to explain the method of stemming the hole and blasting it. Drilling deeper holes for prospecting became necessary with the increasing demand for minerals, particularly coal. On several occasions the depth of the holes for coal prospecting exceeded 6 m. For reaching this depth the idea of joining the drilling rods by means of screwed ends was born. The author of ‘Compleat Collier’, published in 1708, described the method used for exploration holes as, “… We have two labourers at a time, at the handle of the bore and they chop or pounce with their hands up and down to cut the stone or mineral, going round, which of course grinds them small so that finding your rod to have cut down four or six inches they lift up the rod, either all at once, as there conveniency for its lifts or by joynts fixing the key which is to keep the rods from falling down into the hole … taking off the cutting chisel, put or screw on the wimble or scoop which takes of the cut stuff be it what happens; and so by sight of the stuff … and so what consequently follows at any depth; for by addition of the joynts (which we screw on to the rod) we can decend to any depth”. Blasthole drilling became a faster process when two men hammered the ‘piercer’, and even faster when iron rods with chisel ends were initiated in Hungary in 1749. The idea of dropping some weight on a fixed steel bar was first put into practical use by Huthmann in 1683. In his ‘boring machine’ the weight was raised by a rope that was repeatedly pulled and released by two men. Similar machines with some differences were also later developed by Barthels in 1721 and still later by Gainsbrigg in 1803. The dawn of the industrial revolution began in Europe in the late eighteenth century. Activities in virtually every field increased on a very large scale. Many needs and consequent inventions, discoveries etc., directly or indirectly affected blasthole drilling technology. Some of the affecting factors were: 1

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Electric detonators had been invented in 1829 by Moses Shaw in New York. In 1831, William Bickford of Tuckingmill in England invented the safety fuse. Both these inventions made blasting operations far less dangerous as compared to the earlier practice of using black gunpowder for blasting where frequent premature explosions occurred.

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2 3

4 5

6

63

Use of positive displacement steam engines entered mining activities by the end of eighteenth century. Initially steam power was deployed for winches and haulage. Nitroglycerine was developed by Sobrero in 1847. It was far more powerful and devastating than black powder but very dangerous as it exploded with even slight frictional heat. Dynamite developed by Noble in 1866 was equally powerful but far less dangerous. The use of these explosives made it possible to dramatically increase mine productivity and advance rates in tunnels. Construction of many canals began in the first few decades of the nineteenth century. These projects needed excavation in rock where blasting was essential. Invention of the steam locomotive resulted in the laying of railways tracks in Europe, USA and many other countries, as a mean of quick transport. Construction of tunnels was needed to avoid long loops of railway tracks in the hilly regions. Use of compressed air for ventilation in tunnels was practiced in the early 1830s. As the heavy steam boilers could not be taken into tunnels, and as the steam quickly condensed in longer pipes, the use of compressed air instead of steam was introduced for power transmission. In 1844 Brunton suggested the use of compressed air for powering a rock drill and named it the ‘Wind Hammer’.

With all these and many other factors the demand for metals increased exponentially. To cope with such increased demand, drilling larger and deeper blastholes became essential. As the limitations of manual drilling were recognized, the introduction of mechanization in drilling was thought to be the best alternative. Efforts were made in that direction from the early nineteenth century. 4.3

ROTARY DRILLING

The origins of rotary drilling go back to nearly two thousand years before Christ. The earliest proof of human beings drilling holes in stones can be found in the quarries made for extraction of stone blocks for pyramids in Egypt. Holes as deep as 6 m and as large as 500 mm diameter were drilled in those days. Cores of basalt and obsidian, recovered from some of these quarries and kept in pyramids, have now been preserved in some of the well-known museums in the world. The length of some of these cores is as much as 500 mm. A closer examination of the cores indicates helical grooves on the outer surface of the cores. This leads to the conclusion that the drilling method employed must have been very similar to the calyx drilling, which is still practiced in developing countries as a labor intensive low cost form of core drilling. An artist’s impression, based on limited evidence and estimates, indicated that the assembly used by the Egyptians for drilling might have been like the one shown in Figure 4.1. Due to the primitive techniques used in those days the drilling speed was very slow in such rotary drilling. Hence rotary drilling was hardly ever used for drilling holes in the ground. Some references to rotary drilling being used for drilling water wells are found in France in the early eighteenth century. Until the end of the nineteenth century mines were small. In fact larger excavation activities were in the civil engineering domain. Even though the steam excavator was

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64


Figure 4.1 An Egyptian drill, 2000 BC.

invented by Otis in 1838, only a few of them were available to companies other than the manufacturer. Thus, the loading of blasted material into rail transport wagons was done manually. For this reason blastholes used to be of small diameter and shallow depth so they yielded small size fragments suitable for manual operation. In terms of their spacing and depth, blastholes differ radically from the holes drilled for oilwells and waterwells. It was, therefore, obvious that unless the heavy drilling mechanisms required for rotary drilling could become portable, the rotary drilling could not be used for blasthole drilling. The idea of introducing portability in large hole drilling machines was conceived way back in the 1860s, and the earliest portable machine appeared in 1867. It was said that this earliest rotary drilling machine was built by Henry Kelly – the founder of Kelly, Morgan and Company. For his efforts in reviving the rotary drilling method, Kelly is still remembered by his name given to the specialized drill rod with three or more grooves, alternately square or hexagonal in cross section, with which it is possible to impart rotation to the drill string. Kelly, Morgan and Company later became Armstrong Quam Company in 1904 and merged in 1933 with Bucyrus Erie Company, now known as Bucyrus International, Inc. Kelly apparently did not take a patent. His idea of portability was used by others, and the first portable percussion drilling machine of which there is a record was patented by Nelson in 1871. This is shown in Figure 4.2. During the 1870s and 1880s many manufacturers built portable cable tool drilling rigs. R. M. Downie was the first. The drill made by Downie is shown in Figure 4.3. In this drill the tripod type mast was not a part of the drill. A drill with built-in mast was developed by Downie during the early 1890s. The first such machine is shown

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65

Figure 4.2 Nelson’s portable cable tool drill, 1871.

Figure 4.3 Downie’s portable cable tool drill, 1880.

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in Figure 4.4. The trade name of these rigs was ‘Keystone’. Originally such rigs were made for drilling oil wells, but soon they were deployed for blasthole drilling. Spudder was a very common name for such percussion blasthole drills. One of the earliest all-steel spudder made by Armstrong is shown in Figure 4.5. First major project to use spudders was the construction of the Panama Canal. It began in 1904 and was completed in 1914. The authorities of the project initially purchased sample drills from twelve manufacturers and eventually purchased 218 drills made by Star Drilling Co. for the project. The other manufacturers of spudders in those days were Armstrong, Keystone, Austin and Columbia. An early self-propelled percussion blasthole drill on wheels was developed by Armstrong in 1917. One of these drills is shown in Figure 4.6.

Figure 4.4 Downie portable drill with A-frame mast, 1892.

Figure 4.5 Armstrong tow type churn blasthole drill, 1915.

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67

Figure 4.6 Armstrong self propelled churn blasthole drill, 1922.

Figure 4.7 Keystone churn drill with crawler tracks, 1933.

Crawlers were first developed for farm tractors by Benjamin Holt in 1904. They were very successful in slushy ground. Soon their potential in moving on rough ground was also realized and machines used in mines were mounted on crawler tracks. They were termed churn drills. One such churn drill, made by Keystone in the late 1920s, with crawlers on the drilling side of the machine is shown in Figure 4.7.

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68


Churns, as shown in Figure 4.8, were used in many mines till the mid 1950s, when their use sharply declined as rotary blasthole drills for large diameter blastholes were developed. For drilling in extremely hard and tough rocks, a drilling method, called ‘Jet Piercing’, that worked on the principle of thermal fragmentation, was also used successfully in some mines. These drills were jointly developed by Union Carbide and Bucyrus Erie in the early 1950s. A typical jet piercing drill looked like the one shown in Figure 4.9. It could give a penetration rate of the order of 5 to 6 m in an hour and required 170 kg fuel oil and 567 kg oxygen per hour. This method also was discarded in the early 1960s, after the newly developed tungsten carbide inserted tricone bits made it possible to drill large blastholes in hard rocks at much lower expense and in a far safer way.

Figure 4.8 Churn drills in a mine, 1937.

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Figure 4.9 Jet piercing drill of Russian origin.

Hard rock rotary drilling became possible after Sharp and Hughes developed the rolling cutter rock bit in the year 1909. It looked like the one shown in Figure 4.10. In the year 1912 the two cones of the bits were replaced by three cones. This wonderful tool gave tenfold faster penetration rates in hard formations as compared to the drag

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Figure 4.10 Sharp and Hughes two cone bit, 1909.

Figure 4.11 Moving rotary head in a patent filed by Ross, 1891.

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71

bits used till then. In 1925, with the invention of intermeshing cones, the two cone rock bits became acceptable in the oilwell drilling industry. By the second world war, rotary drilling was a standard method used in oil well drilling. The original idea of a moving drill head is quite old as can be seen from Figure 4.11 that represents a sketch from Ross’s patent of 1891. Such mechanisms were used for oilwell drilling rigs, where the tall derricks enabled use of a long drill string. Heavy drill collars were used right at the beginning for exerting a heavy weight on the tricone drill bit.

4.4

ROTARY BLASTHOLE DRILLING

The first use of rotary blasthole drilling was way back in the year 1865. The excavation of Mont Cenis Tunnel in the Alps, that joins French and Italian railways, began in the year 1854. Initially the work progressed at such a slow pace that the completion of the tunnel was expected only after about 40 years. Low drilling speed was the bottleneck in completing the tunnel quickly. As an attempt to increase the speed of drilling, a French engineer named Leschot, working on the project from the French side, developed a drill and a bit studded with diamonds to drill blastholes in the hard rock along the alignment of the tunnel in the year 1863. This drill could drill holes of diameter 50 to 75 mm. The patent drawing of the drill and the diamond bit in the form of a core barrel is shown in Figure 4.12. When both these devices were used together, the speed of drilling increased greatly but was still much less than expected. Drilling with the diamond bit and the drill was soon abandoned because the percussion drill, developed by Sommiller and being used on the Italian side of the tunnel, gave a much faster penetration rate. Thereafter, till today, diamond core drilling is rarely used for drilling blastholes, but the potential of this rotary drilling method was recognized in the USA as early as 1876. A huge steam driven diamond drill, shown in Figure 4.13, was exhibited in the Centennial Exhibition of Philadelphia State. Today, a greatly improved form of diamond drilling is very extensively used for oilwell drilling. Often blastholes are required to be drilled in hard rock right from the top of the ground surface. The heavy feed force necessary for such drilling cannot be generated by using heavy drill collars with a large, tall derrick that poses a problem in mobility. In the second half of the 1940s the problem was solved by as many as three manufacturers each in a different manner. In the late 1890s some manufacturers started making self-propelled rotary drilling rigs. These were made for exploratory drilling. One of the earliest rotary drills, made by Austin Manufacturing Company, is shown in Figure 4.14. In 1946 the Joy Manufacturing Company, a successor of Sullivan Machinery Company, introduced the first crawler-mounted rotary blasthole drill. This drill used a tricone bit rotated by a Kelly drill rod and rotary table mechanism. Water was used as a circulating fluid. The feed force required on the tricone bit was exerted through a chain pulldown mechanism. The water which accumulated in the blastholes was removed by lowering a pipe to the bottom of the hole and flushing the hole with compressed air.

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Figure 4.12 Drawing of the diamond drill and diamond studded core barrel submitted for second patent by Leschot in 1869.

The idea of the use of rotary drilling for blastholes worked well and many Joy blasthole drills were deployed in mines. In 1949 a mud pump on one of the blasthole drills working on the site of Michigan Limestone Co. failed and could not be repaired or replaced quickly. The company experimented with compressed air instead of water for a flushing medium. The results of this innovation were very encouraging. Very fast penetration rates could be achieved but the damage to the bearings of the tricone bits was excessive. Hughes Tool Company joined the drilling program and noticed that bit failure resulted from inadequate cooling of the bit bearings. They soon developed a tricone bit where compressed air was passed through passages made around the bearings. The idea worked and rotary drilling was established in blasthole drilling practice in 1949, with model 56BH made by Joy Manufacturing Company. This 56BH blasthole drill featured air circulation to the tricone rock bits through its drill string. The drill string was rotated through a rotary table, very much like the rotary table used in the oil well drilling industry. This drill is illustrated in Figure 4.15. The use of a Kelly in the drill was a limiting factor. Adding drill pipes to the drill string was rather difficult due to the presence of a Kelly and hence drilling deeper holes was very cumbersome.

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Figure 4.13 Sullivan P2 diamond core drill, 1876.

Figure 4.14 Self propelled rotary drill by Austin manufacturing company, 1903.

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Figure 4.15 Joy 56BH blasthole drill, 1946.

Joy 56BH was acceptable for shallow blastholes of about 10 m depth, but for matching the excavation capabilities of some large draglines and stripping shovels the need was for 250 mm diameter blastholes of more than 30 m depth. Bucyrus Erie Co., which enjoyed a very large market share for their churn drills, anticipated a challenge from the rotary drilling method and soon started developing a rotary drill. In 1949, they developed a drill where a drill head mounted on the top of a frame was pulled down by means of hydraulic cylinders as shown in Figure 4.16. The prototype never made it to the worksite, probably because Bucyrus anticipated something far better. In 1950 Wendell Reich developed a crawler-mounted blasthole drill with a moving rotary head driven by hydraulic motor. It was pulled up and down by a hydraulic cylinder and wire rope mechanism. These blasthole drills, due to their compressed air circulation arrangements, became quite popular and were used for drilling 150 to 165 mm blastholes in hard formations. In 1951 Bucyrus Erie Company developed a heavy crawler mounted blasthole drill model 50R and introduced it in practice in 1952. In the 50R the drill string was rotated by a top drive rotary head. This rotary head was moved up and down through a wire rope wrapped around a shipper shaft drum. It could be used for drilling blastholes of diameters up to 311 mm to depths of 39.5 m with a carousel for holding four drill pipes. Unlike Joy and Reich the Bucyrus drill was all electric. The only hydraulic actuators in the drill were the hydraulic leveling jacks. Figure 4.17 shows an early 50R drill. Electric drives used for other mechanisms offered extreme reliability in the machine. As the blasthole diameters possible with the 50R were most suitable for the production rates needed to match large shovels and draglines, the 50R became very popular. Soon smaller drills viz. 30R and 40R were added to the Bucyrus Erie lineup.

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75

Figure 4.16 Earliest version of Bucyrus rotary blasthole drill, 1949.

The extreme wear resistance and hardness of Tungsten Carbide (TC) inserts used for detachable bits was known from the field use of top hammer drilling. By taking a clue from such use, Morlan, Scott and Wood of Hughes Tool Company developed a tungsten carbide inserted tricone bit in 1951. It was initially called ‘Chert Bit’. As the bit did not have sharp teeth, like all the steel bits used to have till then, drillers doubted the capabilities of the new bit. However, to everyone’s surprise the Chert Bit gave almost the same penetration rate under the same operating conditions, and gave 3 to 10 times longer life. Soon, in 1959, chert bits with air circulation arrangements were also made for blasthole drilling. Today, TC inserted tricone bits are used in most hard rock blasthole drilling all over the world. Loss in electric power transmission is very low as compared to the equivalent hydraulic system. Thus, under most circumstances an all-electric drill is likely to be more efficient and reliable, but with high pressure hydraulics and improved components the difference has dwindled. For this reason the use of electric power for driving the components of rotary blasthole drills has now remained confined to very large blasthole drills. The decade of the 1960s saw a sudden spurt in demand for energy and metals. This resulted in an exponential increase in coal and metal mining activities in the USA and other parts of the world. The demand for rotary blasthole drills was so high that several companies started the manufacture of rotary blasthole drills. Leswell’s article published in 1977 lists as many as 12 manufacturers from the USA. Some of them like Portadrill, Davey/Rousselle made only truck-mounted rotary blasthole drills and eventually went out of the race because blasthole drilling work always looked for sturdy crawler mounting.

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Figure 4.17 Bucyrus Erie 50R introduced in 1952.

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77

Many manufacturers use hydraulic power rather than electric for small, medium and large drills and even the stubborn supporters of electric now use hydraulic power for driving the propel motors of their very large electric drills. Rotary blasthole drills can be divided into four classes, on the basis of the Feed Force (FF) they can exert on the drilling bit, as follows: Small Medium Large Extra Large

(100 kN < FF < 225 kN), (226 kN < FF < 350 kN), (351 kN < FF < 500 kN), and (FF > 501 kN)

Today, in 2010, rotary blasthole drills are being made by many companies spread over many countries. A broad idea of their product range is shown by Table 4.1. In the late 1970s the list of manufacturers of rotary blasthole drills included Atlas Copco, Bucyrus Erie, Chicago Pneumatics, Driltech, Gardner Denver, Hausherr, Ingersoll Rand, Joy Robbins, Marion, Reedrill, Rudgormash and Schramm. The manufacture of rotary blasthole drills is not a very lucrative business. Many companies built rotary blasthole drills in the past, but very limited demand for the product has forced them either to close or sell the manufacturing rights and lose their identity. Bucyrus kept their leadership in large rotary blasthole drills virtually till the end of the last millennium, through acquisition of Marion. In the meanwhile in 2003 Atlas Copco, well known all over the world for their top hammer and down-the-hole hammer drills acquired and amalgamated the Ingersoll Rand drilling equipment line with their percussion and DTH drill line. With this move they seem to have overtaken Bucyrus. However, Bucyrus have recently

Table 4.1 Rotary blasthole drill manufacturers and their range. Manufacturing range

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Manufacturer

Country of mfg.

Small

Medium

Large

E. large

Atlas Copco Bucyrus Intl. Changsha Keda CMCL Emico Elecon Globe Drill Hausherr Heavy Eng. Corp. KLR LMP Precision NKMZ P&H Mining Revathi Equipment Rudgormash Sandvik Driltech

USA USA China India Australia Germany India India India USSR USA India USSR USA

Yes Yes Yes Yes

Yes Yes Yes Yes

Yes Yes Yes

Yes Yes Yes

Yes

Yes Yes Yes Yes Yes

Yes

Yes Yes

Yes Yes

Yes Yes Yes

Yes Yes Yes Yes Yes

Yes

Yes

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countered this move by acquiring Terex Mining who, about three years ago had acquired and amalgamated Reedrill. Whether Bucyrus regain their leadership is yet to be seen. Other major players in large rotary blasthole drills are P&H Mining with inheritance from Robbins and Gardner Denver, and Sandvik with inheritance from Driltech. Very heavy machines, weighing more than 150,000 kg and capable of exerting bit loads more than 550 kN are now being manufactured by Bucyrus International, P&H Mining and Atlas Copco. Some of them can drill 20 m deep holes of diameters in excess of 406 mm in one single pass. The largest amongst these is the Bucyrus International 59HR that weighs up to 193,000 kg, can exert a bit load up to 734 kN and is rated for drilling blastholes of 445 mm diameter. It is shown in Figure 4.18. Most of the manufacturers mentioned above are of US origin, but the Russians had also independently developed electric blasthole drills for drilling blastholes of 320 mm diameter to 32 m depth some forty years back. They are often found in East European and some other Asian countries. One of the deep-rooted ideas in the minds of blasthole drill designers was a drill with twin mast, meant to drill two blastholes simultaneously. Marion Power Shovel Company developed one such drill and put it to field trials at Squaw Creek Mine. It did work and produced results but not to the level of satisfaction that the customer could order more of its kind. The drill is shown in Figure 4.19. All through the history of rotary blasthole drilling, the objective in improving machinery has been to introduce safety, efficiency, reduced maintenance and automation. The earliest step towards this was pipe handling. Carousel pipe changers were used even in the early 1950s. Single pipe changers, developed in the 1960s, were better at handling long and heavy drill pipes. Drilling recorders were introduced in blasthole drills in the early 1960s. In the early 1980s they became electronic. Automatic fire extinguishing systems and automatic lubricating systems were incorporated into blasthole drills in the late 1960s. In 1971 blasthole drills entered an era of automation, when Bucyrus Erie adapted a primary logic circuit to control some of the drilling operations on their blasthole drill model 60R. Early electric drills featured Ward Leonard controls for accurately controlling the DC electric motors in the rotary head. Some manufacturers adopted AC motors and frequency variation to control the motors in the early 1980s. However, advances in Static or Thyristor Control for DC motors have still retained DC motors on the rotary heads of blasthole drills. In the 1990s integrated circuits revolutionized the blasthole drills to a very great extent. They made it possible to precisely measure and monitor several functions in the drilling operation. Some features offered by such systems were useful in the following manner. 1

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With the drilling recorder the driller could record the values of as many as 6 parameters of the drilling activities. With this facility the operators could moni-

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2

79

tor the drilling process very effectively. Later, when digital systems were introduced the number of parameters went to as much as 12 and their values could be seen on the screen in the operator’s cab. Blasthole drills could be automatically leveled within an accuracy of 0.25°.

Figure 4.18 The largest, Bucyrus International’s 59HR, 2003.

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Figure 4.19 Marion twin mast rotary blasthole drill, 1986.

3

4

5

With specialized drill process software, stored in the computerized system of the blasthole drill, it became possible to control drilling operations in the most appropriate manner by automatically adjusting feed force, feed rate, rotary speed, compressed air supply and so on. Radio control for propelling operations of the blasthole drills made it possible to move the drill to the next blasthole by an operator who stood on the ground near the new hole position rather than the operators cab of the drill at the position of the already drilled hole. Very high positioning accuracy was achieved through such remote control propelling. With the introduction of Global Positioning Systems (GPS) it became possible not only to navigate the blasthole drill even more accurately, but also to integrate the drilling operation with a comprehensive mine operation system to increase productivity.

Technological advances in the new millennium have pushed automation further by introducing GPS. More details on these are given later in a separate chapter that deals with computers in rotary drilling. 4.5 TRUCK MOUNTED ROTARY BLASTHOLE DRILLS For almost a decade after the first world war, self propelled cable tool drilling rigs were being manufactured by manufacturers like Keystone, Star, and Armstrong. However when trucks became easily available in the early 1930s, many drill manufacturers started making their drills on a frame which could be simply mounted at the time of delivery onto the truck chassis purchased from the truck manufacturers.

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Figure 4.20 Sandvik truck mounted rotary blasthole drill, 2008.

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This move was beneficial to both manufacturers and purchasers. Drill manufacturers because it considerably reduced their inventory, and purchasers because it enabled them to get spare parts for half of their machine far more quickly from the truck manufacturers. The choice of truck-mounted rotary blasthole drills for the main mining operations is made very rarely for several reasons, explained in one of the next chapters, but many mining companies often keep one rotary blasthole drill in their fleet of equipment because: 1 2 3

It can be taken from one bench to another in much less time. It can be used for drilling deep holes often required in mine benches. If the main crawler blasthole drill fails and needs a long time for repairs, the truck-mounted blasthole drill can temporarily carry out the blasthole drilling work as a stopgap arrangement.

In the early 1970s many manufacturers made truck-mounted rotary blasthole drills, but currently only a few of them have continued their manufacture. Truck-mounted rotary blasthole drills are now available from Atlas Copco, Sandvik, Schramm. Most of the truck-mounted rotary blasthole drills can handle four to six drill pipes of length 7.6 m and diameters up to 127 mm, and are suitable for drilling 200 mm blastholes. One such drill is shown in Figure 4.20. Their bit load capability is of the order of 133 kN. Larger truck mounted drills, capable of drilling blastholes of diameters up to 311 mm with 219 mm drill pipes of length 12.2 m each, are also available. These drills have bit loading capacity of 311kN.

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Chapter 5

Rotary blasthole drilling bits

5.1

INTRODUCTION

The drill string used in rotary drilling comprises many items, but the most important component in it is the drill bit. The drill bit disintegrates the formation under its teeth or cutting edges, and enables progress of the blasthole in the intended direction. In this process the bit undergoes very heavy wear and the penetration rate eventually drops down below an acceptable level. Thereafter the bit is required to be replaced. As a drill bit is an expensive and critical item, it is thoroughly dealt with in this separate chapter. Other accessories used in a rotary drill string are elaborated in chapter 6. Two types of drilling bits, viz. drag bits and tricone bits are used in blasthole drilling.

5.2

DRAG BITS

In the early days, drag bits made from cast steel were used for drilling blastholes in soft and very soft formations. Ever since tungsten carbide inserted drag bits made entry into the field of blasthole drilling, the use of cast steel bits stands almost discarded. Today, tungsten carbide inserted drag bits are used everywhere but in the near future drag bits, made with thermally stable diamond compacts, may change the whole scenario of rotary blasthole drilling. Two types of tungsten carbide inserted drag bits, viz. blade bits and claw bits, are currently used.

5.2.1

Blade bits

Blade bits have two types of cutting edges viz. Chevron Type and Step Type. A Bit with chevron cutting edge is shown in Figure 5.1. Such a cutting edge is usually found on small diameter bits used for drilling blastholes in underground mines. The step type cutting edge, shown by line drawing in Figure 5.2, is adopted for larger diameter bits that are used for drilling blastholes in surface mining operations. Step type drag bits are less prone to blasthole deviation. Blade bits have two, three or four wings. At their cutting edges they have flat, blade-like tungsten carbide inserts brazed to the main cast steel body.

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Figure 5.1 Chevron blade type drag bit.

Figure 5.2 Step blade type drag bit.

Experiments at various research centers have indicated that the shear strength of the material used for bonding the carbide inserts to the steel body of the blade bit should be more than 240 MPa for the bit to work satisfactorily. If blade type bits are used with air flushing, they attain high temperatures, and the possibility of losing a tungsten carbide insert increases considerably due to weakening of the brazing material that joins the carbide insert to the bit body. Therefore, blade type bits are either used with water circulation or water is injected in the stream of compressed air used for flushing.

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Since most of the soft formations, such as alluvial overburden in coal mines, coal seams, potash deposits etc., that require blasthole drilling are non-abrasive, the body of the bit remains intact for a long time. In such a situation drag bits with replaceable blades, as shown in Figure 5.3, can be used with advantage. Details of some blade type drag bits used in blasthole drilling are given in Table 5.1. More data can be obtained from the manufacturers.

5.2.2

Claw bits

A typical claw bit is shown in Figure 5.4. Depending upon the bit diameter, it has five or more replaceable teeth and a replaceable pilot. All the teeth are fixed in a forged, machined and heat-treated bit body by means of steel pins or circlips. Since the inserts in the claw bits are not brazed, the bit can be used with air circulation along with water injection. The performance of claw type drag bits is better, as penetration of the teeth of the bit in the soft formation is deeper. For this reason, claw bits require higher torque than the blade bits. Claw bits are chosen for drilling in formations having UCS ranging from 35–70 MPa. In any case, while drilling blastholes with drag bits, a large quantity of water is required to be injected with the compressed air for bit cooling. Details of some claw type drag bits used in blasthole drilling are given in Table 5.2.

5.2.3

Feed force and rotation of drag bits

Due to its low importance, the process of drag bit drilling has not been researched or recorded as deeply as have tricone bits. Field practice in using drag bits suggests that blade bits should be used with a bit weight of 100–200 N per mm bit diameter and rotated at 20–50 rpm, whereas claw bits should be subjected to a feed force of 3000–5000 N per tooth of the bit and rotated at 10–40 rpm.

Connector

Reaming Stabilizer Holder

Bit Shank

Blades

Figure 5.3 Replaceable blade drag bit.

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Table 5.1 Details of some blade type drag bits. Nominal bit diameter mm

Inches

Connection

48–64 70 73 70–89 92–120 127 130 133–140 143 152–171 178–203 178–311

1.7/8–2.1/2 2.3/4 2.7/8 2.3/4–3.1/2 3.5/8–4.3/4 5 5.1/8 5.1/4–5.1/2 5.1/4–5.1/2 6–6.3/4 7–8 7–12.1/4

AW or A Rod AW or A Rod AW or A Rod NW or N Rod 2.3/8" API Reg. 2.3/8" API Reg. 2.7/8" API Reg. 2.7/8" API Reg. 3.1/2" API Reg. 3.1/2" API Reg. 3.1/2" API Reg. 3.1/2" API Reg. 4.1/2" API Reg.

Configuration – (chev./step) and number of wings C3 C3 C3 C3, C4, S3 C3, C4, S3 C3, C4, S3 C3, C4, S3 C3, C4, S3 C3, C4, S3 C3, S3 C3 S3

Figure 5.4 Claw type drag bit.

5.3 TRICONE BITS Today, only about 2% of rotary blasthole drilling is carried out with drag bits. The remaining 98% is carried out with tricone bits. Tricone bits with meshing cones were introduced in oilwell drilling practice in 1933. Initially tricone bits were used only for oilwell drilling. When bits with a facility for compressed air circulation through their bearings were developed in 1949,

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Table 5.2 Details of some claw type drag bits. Nominal bit diameter mm

Inches

Connection

Number of teeth

98–101 105–108 114–120 127–130 140–149 152–165 178–190 203–216 250 270 280 311

3 7/8–4 4 1/8–4¼ 4½–4¾ 5–5 1/8 5 ½–5 7/8 6–6½ 7–7½ 8–8½ 9 7/8 10 5/8 11 12

2.3/8" API Reg. 2.3/8" API Reg. 2.3/8" API Reg. 2.3/8" API Reg. 2.7/8" API Reg. 3.1/2" API Reg. 3.1/2" API Reg. 4.1/2" API Reg. 4.1/2" API Reg. 6.5/8" API Reg. 6.5/8" API Reg. 6.5/8" API Reg.

4 4 4 4 5 6 6–7 8 9–10 10–12 12 14

Pilot bit diameter mm 40 40 40 40 40 40 50 50 50 50 50 50

tricone bits were also introduced into blasthole drilling practice. Tungsten carbide inserted bits were initiated in 1951, and sealed bearing bits made their appearance in 1959. Sealed journal bearing bits were launched in 1968.

5.3.1 Types and nomenclature Drag bits, as shown in Figure 5.1 to 5.4, do not have any moving parts. Tricone bits, however, are assembled by using many parts, some of which move while the drilling operation goes on. The reasons for having three cones on a bit are: 1

2

The contact between the cone of a bit and the horizontal bottom of the blasthole is always linear. In two-cone bits these two lines are actually collinear. In threecone bits the three lines are radial. Therefore, a blasthole drilled with two-cone bits is far more prone to deviation. Tricone bits exert the bit weight in three vertical planes in a far better balanced manner and reduce the hole deviation to a great extent. The balanced bit load distribution also gives longer life to a tricone bit as compared to a two-cone bit. Bits with more than three cones can give better weight distribution but in such case the size of the cones has to be reduced and so the bearings within them have to be smaller. This leads to very weak bits that are susceptible to early bearing failure.

Primarily two types or tricone bits are made, viz. milled tooth bits and tungsten carbide insert bits. Milled tooth bits are also called steel tooth bits. In milled tooth bits the cones are completely made from alloy steel and the teeth are formed by milling – a mechanical metal cutting process.

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Tungsten carbide inserted bits have different types of carbide inserts firmly fixed into the holes drilled in the cones by another manufacturing technique called shrink fitting. Since tungsten carbide is a much harder and abrasion-resistant material, the carbide inserted bits give far better performance – often 3 to 9 times – than that of a comparable milled tooth bit operating under similar conditions. Milled tooth and tungsten carbide insert tricone bits are shown in Figure 5.5. Tungsten carbide inserted bits are more expensive than milled tooth bits due to the higher price of tungsten carbide inserts and added manufacturing costs. Some manufacturers make composite bits where the cones have steel teeth on all the portion except the outer edges, where tungsten carbide inserts are fitted. The design principles of a milled tooth bit also apply to a tungsten carbide insert bit. Three cones of a milled tooth tricone bit, when seen from bottom of the bit, look like those shown in left side picture in Figure 5.6. No. 1 cone is the cone which has a spear point i.e. the teeth at the apex of the cones. The next cone, i.e. cone no. 2, has teeth on a circle at a small distance from the apex of the cone. The row of these teeth is called the nose row. The nose row on cone no. 3 is at a slightly larger distance from its apex compared to cone no. 2. The diameter of a tricone bit is equal to the inner diameter of the smallest circular ring through which the bit can pass. This dimension is treated as the nominal hole size of a new bit. Even though used for blasthole drilling, tricone bits are made to the standards laid down by the American Petroleum Institute. Acceptable tolerances in the diameter of tricone bits as per API Standards are given in Table 5.3.

Figure 5.5 Milled tooth and tungsten carbide insert tricone bits.

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Gage Teeth

Bit Leg

Cone No.2

Air Passage Air Exit

Nose Row

Intermediate Row

Ball Bearing

Nose of Cone 2

Spear point

89

Nose of Cone 3 Cone Shell

Intermediate Row Nose Row Gage Row Roller Bearing

Gage Teeth

Gage Teeth Cone No.3

Cone No.1

Gage Insert Bush Bearing

Thrust Button

Figure 5.6 Nomenclature of tricone bits. Table 5.3 Acceptable tolerances in tricone bit diameter. Bit diameter mm

Higher size tolerance

Lower size tolerance

85.7–349.2 mm 355.6–444.5 mm Larger than 445 mm

0.79 mm 1.59 mm 2.38 mm

0 mm 0 mm 0 mm

In the early days, tricone rock bits of diameters as low as 100 mm were made but, since it is now possible to drill holes of diameter up to 150 mm by rotary percussion drilling at faster penetration rates, use of such small size tricone bits has remained restricted to soil investigation and some other purposes, where drilling is to be carried out with least disturbance to the rock mass. In any case, small diameter tricone bits do not perform well in hard formations, because small tricone bits have very small bearings. Further it is also not possible to give adequate feed force on the tricone bits through slender drill rods as they are susceptible to buckling. The cross section of a roller and corresponding leg is shown on the right-hand picture in Figure 5.6. To enable the roller to roll over the formation as the drill string is rotated, bearings are provided between the cones and respective legs. Usually the innermost bearing is a journal type. It involves no moving parts, and therefore can transfer very heavy radial loads. The middle bearing is ball type. It interlocks the cone with the leg. The outermost bearing is roller type. It allows the bit to withstand very heavy downward thrust while being rotated at low frictional resistance. Figure 5.6 shows three rows of teeth on the cone and three rows of bearings. These are applicable to bits of small diameter. Sometimes more than three rows of teeth as well as bearings are provided on large diameter bits. Components of a tricone bit are made separately and are assembled later. Each of the three legs of a tricone bit is made from a forging of an alloy steel. The forged leg is very precisely machined for the dimensions of grooves for roller, ball and

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journal bearings. Once the machining process has finished, the machined portion is heat treated to the desired hardness by carburizing. Figure 5.7 shows a tricone bit leg. The cones are made from their respective forging, then machined and heat treated. After this, if applicable, carbide inserts are shrink-fitted in the holes. Journal and rollers of the bearing are then inserted in the appropriate grooves of the cone. Once this is done, the leg is inserted in the cone in which the roller bearings are held in their final position. The balls of the ball bearings are inserted from outside through a hole provided for the purpose.

Figure 5.7 A tricone bit leg.

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Assembling of one leg is complete once the balls are appropriately inserted and the hole is plugged by a pin that is later welded to the leg. The three legs of a bit are then held together firmly and their sides are welded to form a tricone bit. After the welding and stress-relieving operations are over, the top portion of the bit is machined to form an API Regular pin connection. This connection enables coupling the bit to the stabilizers or bit subs which are the next bottommost components of a drill string. The names of many other portions of a tricone rock bit are shown in Figures 5.8 and 5.9.

5.3.2

Geometrical design aspects

Many factors relating to the geometry of the components of a tricone bit, such as cone size, cone offset, size and shape of teeth etc., play a very important role in its field performance. While designing a bit, all these must be given due consideration. 5.3.2.1

Fluid circulation

Aspects related to removal of cuttings from a blasthole by circulating compressed air have been dealt with in a separate chapter of this book. Apart from the need to bring the drill cuttings to ground surface, the compressed air is also required to dissipate the heat generated by friction between cones and the formation, as well as within the bearings located inside the cones of a tricone bit. If bits with bearings lubricated in the factory and meant to be used with water circulation are used in blasthole drilling, the compressed air proves insufficient in heat dissipation, and consequently the bearings fail soon after the lubricant melts and oozes out or burns and becomes ineffective. In 1949 this problem was solved by providing air passages like those shown in Figure 5.9. Another, more recent way, is to use a bit provided with a composite seal, made of elastomer plus metal, and a reservoir built into the bit leg to store a special lubricant. Such lifetime-lubricated bits can defy temperatures to even 400°C. 5.3.2.1.1

Air cooled tricone bits

Air-cooled bits are of two types viz. regular and jet circulation as shown in Figure 5.10. In a regular circulation tricone bit, three passages to the bearings of the three cones are provided for air circulation through the bearings and the main airflow is through a central hole. In a jet circulation tricone bit, three passages to the bearings of the three cones are also provided for air circulation through the bearings, but the main airflow is also through three jet nozzles provided on the three legs of the bit. As a general norm, it is desired that at least about 30% of the total compressed air supplied to the bit should flow through bit bearings so the bearings cool sufficiently. The size of the air passages to bearings is given appropriate dimensions to achieve this objective. In most of the regular circulation bits, this objective is achieved when compressed air pressure at the bit is about 280 to 300 kPa. In jet circulation bits, three independently replaceable nozzles are provided to achieve this. When nozzles of smaller

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Nozzle Retaining Pin

Bit Type Emboss

Bit Serial Number Emboss

Shank

Bit Size Emboss

Shank

Nozzle

Bit Assembly Number Emboss One of the Three Legs

Shirttail Hardfacing

Gage Teeth

Figure 5.8 Nomenclature of tricone bits.

Cone Gage Surface

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93

Water Separator

Snap Ring

Air Screen Bit Leg

Rubber Flapper Screen Tube Assembly

Main Air Passage Air Passage to Ball Bearing Air Passage to Pilot Bearing Pilot Pin Shirttail Inserts

Cone Gage Inserts

Leg Journal

Cutting Surface Inserts Cone

Roller Bearing Ball Bearing Bush Bearing

Figure 5.9 Air cooling of blasthole bits.

Regular Circulation

Jet Circulation

Air Flow through Central Hole

Air Passage to Bearings

Air Flow through Jet


Jet Nozzle

Figure 5.10 Regular and jet circulation bits.

central bore are chosen, a lower volume of compressed air flows through the nozzles and consequently a higher volume of compressed air is forced to flow through the air passages of the bearings on account of higher pressure built up behind the nozzles. Blasthole drills often have provisions for injecting water into the flow of compressed air. This injected water flows to the drill bits. The purpose of such water circulation is

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to suppress the dust formed during formation fracture. This water, in the form of fog, also passes through the bearings of the bit and helps in cooling the bearings. Many blasthole drills are provided with filter type dust collection systems instead of dust suppression by water injection. In such blasthole drills, if an oil injection system is provided instead of water injection, the oil mixed with compressed air reaches the bearings of the drill bit and lubricates them. This type of bit lubrication proves far more effective and the life of the bit increases by as much as 20%. Air circulation tricone bits are usually provided with a back flow valve at the top, so sudden flow of water in the bit is prevented. 5.3.2.1.2

Sealed bearing bits

Originally these bits were developed for drilling geothermal wells where they had to work in high temperature surroundings. Now they are also used in blasthole drilling applications. The cross section of a leg of a sealed bearing bit looks like the one shown in Figure 5.11. The lubricant reservoir stores an adequate quantity of lubricant and supplies it to the bearings in the cone. The bellow type reservoir keeps the lubricant under pressure,

Sealed journal bearing

Diaphragm Reservoir Cap Grease Reservoir Silver Plates Beryllium Copper Thrust Face Radial Elastomer Seal Journal Hard Metal Gauge Gauge Protection Friction Pin Cone Tungsten Carbide Inserts

Figure 5.11 Cross section of a sealed bearing bit.

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so the lubricant flows to the bearings. The lubricant is prevented from escaping out of the cone by a composite seal of elastomer and metal. Sealed bearing bits have given considerably better performance in terms of drilled meterage and long life as compared to air circulation bits. 5.3.2.2

Size and shapes of teeth

A tricone bit, and the pattern of fractured formation made by one at the bottom of a hole, are shown in Figure 5.12. The indention made by the teeth can easily be observed in the pattern. Actually, the pattern shown is for a hole made in soft rock by mud flush drilling. In blasthole drilling, a bit is exerted with slightly higher feed force. The hole is flushed by compressed air. It removes larger cuttings from the hole bottom without any need for secondary crushing. Thus, in blasthole drilling, even deeper indentions than those shown in Figure 5.12 are formed at the hole bottom. Since in subsequent formation fracture the teeth have to touch at the crest in between two indentions, the number of teeth is often kept odd for the fracture to be more effective. Milled tooth bits have three distinct shapes for a tooth viz. Chisel, H and T. They all are shown in Figure 5.13. All these teeth shapes are formed by milling processes in the rock bit plant. When the angle between the two slanting surfaces of a tooth is small, tooth height is large and vice versa. For drilling in softer formations, the taller teeth give better performance because they can fragment larger cuttings of rock without getting blunt. For drilling in harder formations, the teeth are required to have lesser height since in such cases the tooth angle reduces and the teeth are less susceptible to breakage. The T and H type teeth are found only on the outermost, i.e. gage, rows where they give excellent resistance to higher wear. The teeth of milled tooth bits are usually given a hard facing on their sides for the purpose of extra wear resistance. In carbide inserted bits, the insert tips are made with different shapes during the sintering process. Some insert shapes are as shown in Figure 5.14. The apex angle between the sides of the surfaces near the tip differs considerably. Larger apex angle at the tip gives more strength and higher wear resistance to the insert; however, it reduces the penetration of the insert in the rock. Tricone bits meant for drilling in soft formations have inserts with low apex angle. Inserts at the cone gage and shirttail portion have a dome shape to give very high wear resistance because even in soft formations they merely rub against the formation. In extremely hard formations, carbide insert bits with synthetic diamond caps or diamond coating are chosen. 5.3.2.3

Cone size

The cones of the tricone bit are never true cones but a composite of two cones as shown in Figure 5.15. The main reason for choosing such a shape for the cone is that such a cone is bulged at the center and, therefore, it can accommodate much larger ball and roller bearings within it. This gives higher strength to the cone and the bearings, so much heavier bit loads can be exerted.

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Figure 5.12 Indention made by tricone bit.

When such a composite cone rolls along the formation surface, the teeth of the bit penetrate into the formation and cause indention. Neither the shapes of the teeth, nor the shapes of the indention caused by the teeth, have the shape of a perfect gear tooth. Therefore, slip is caused and side loads are exerted on the bit as well as on the formation. This causes large scale fracture of the formation.

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Top Surface

Side Surface

Side Surface

Chisel Type Tooth

H Type Gage Tooth Shirttail Hardfacing Gage Hardfacing Tooth Hardfacing

97

T Type Gage Tooth

Figure 5.13 Various shapes of tricone bit teeth.

Soft Formations

Soft to Medium Formations Medium to Hard Formations

Hard Formations

Figure 5.14 Various shapes of carbide inserts.

Deeper indention is desirable while drilling soft formations because it creates larger side loads and a larger portion of the formation gets fractured without causing any damage to the tooth. In the case of tricone bits to be used for drilling in hard formations, the outer cone is made bigger and the inner cone smaller, as shown in Figure 5.16. This enables accommodation of even larger bearings. The teeth of such bits are also made smaller so as to reduce the amount of slip when the cone rotates and the teeth penetrate. All this makes the bit suitable for drilling in hard formations where the strength of the cone has higher importance. In hard formations, the slip and consequent side force on the formation

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Envelope of Outer Cone

Cross Section of Actual Cone

Envelope of Inner Cone

Axis of Cone Rotation

Figure 5.15 Two cone composition of the cone of a tricone bit.

Envelope of Outer Cone

Cross Section of Actual Cone

Envelope of Inner Cone

Axis of Cone Rotation

Figure 5.16 Large outer and small inner cone envelopes of the cone of a tricone bit.

does not contribute much to the formation fracturing. Thus, a lower amount of slip is more desirable. A lower amount of slip also causes less wear on the tooth. From the above it can be easily surmised that a longer tooth will give more formation fracture but will be more susceptible to breakage and wear. If a bit meant for drilling soft formations is used for drilling hard formations, premature tooth breakage will occur. 5.3.2.4

Number of teeth

As the height of the teeth of a soft formation bit is large and that of hard formation bit is small, it is possible to accommodate more teeth on the cones of a hard formation bit. This can be observed in Figure 5.17, where cones of the same size have been

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shown. In harder formations, more teeth are necessary because the material removed by each tooth is relatively small, and to make full use of the bit strength in order to allow higher bit loads, more teeth are desirable. For the same reason, there are more inserts on a tungsten carbide inserted bit meant for drilling in hard formations. 5.3.2.5

Cone orientation

If a perfect cone is rolled over a smooth plane surface, it will roll through a perfect circle on that surface. The apex of the cone will lie at the center of such circle. In such rolling of the cone, there will be no slippage between the cone surface and the surface in the circle over which the cone rolls. Even a composite smooth cone (without any teeth) will not slip on a surface formed by two cones. In order to get more formation fracture and faster penetration rate in very soft formations such as coal, it is desirable to have larger amount of slip of the teeth on the cones. Such slip gives a much higher volume of fractured formation and consequently much faster penetration rate. There is always a limit imposed by the geometry on the height of the teeth. Thus, to achieve the higher amount of slip that is desirable in drilling soft formations, the cones of the bit are oriented with an offset as shown in Figure 5.18.

Many small height teeth can be accommodated in a bit meant for drilling in hard formation

Fewer tall height teeth can be accommodated in a bit meant for drilling in soft formation

Figure 5.17 More teeth on hard formation tricone bits.

Direction of Bit Rotation

Bit Rotation Axis Offset Angle

Axes of Cone Rotation

Figure 5.18 Offset of mounting of the cones.

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Offset can be defined as the perpendicular distance between the axis of rotation of the bit and the axis of rotation of the cone. When smooth cones with offset mounting are rotated on any surface profile they invariably slip. If the cones have teeth, the offset adds to the slip of the teeth and the formation is fractured with higher side forces. This results in a faster penetration rate. Even hard formation tricone bits also have some offset. Only the bits meant for drilling in extremely hard formations do not have any offset because normally such formations are very abrasive and providing offset in them creates a large magnitude of friction which results in very rapid tooth wear. 5.3.2.6

Top connections

Most of the developments of tricone bits were made long before the bits appeared on the blasthole drilling horizon. These developments were quickly built into the Standards set forth by the American Petroleum Institute. Air blast tricone bits are used not only by the blasthole drilling industry but also by the oilwell and waterwell drilling industries. It is, therefore, quite natural that tricone bit manufacturers have retained API Standard for air blast bits and have adapted API Regular type for the top-threaded connection of their bits. Tricone bits of different sizes have different API Regular pin connections as detailed in Table 5.4. The table also gives approximate weights of the bits. Some manufacturers do make blasthole bits with BECO connections which are more popular in the mining domain.

5.3.3

Metallurgy

Metallurgy plays a very important role in the manufacture of tricone rock bits. Important parts of a tricone bit, viz. the legs and cones, are made from forgings. After

Table 5.4 Top connections and weights of most commonly used for tricone blasthole bits. Nominal bit diameter

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mm

Inches

Connection

Weight of the bit in kg

143 152 159 171 187 200 229 251 279 311 349 381 406 445

5 5/8 6 6¼ 6¾ 7 3/8 7 7/8 9 9 7/8 11 12¼ 13¾ 15 16 17½

3.1/2″ API Regular Pin 3.1/2″ API Regular Pin 3.1/2″ API Regular Pin 3.1/2″ API Regular Pin 4.1/2″ API Regular Pin 4.1/2″ API Regular Pin 4.1/2″ API Regular Pin 6 5/8″ API Regular Pin 6 5/8″ API Regular Pin 6 5/8″ API Regular Pin 7 5/8″ API Regular Pin 8 5/8″ API Regular Pin 8 5/8″ API Regular Pin 8 5/8″ API Regular Pin

10 13.6 15.9 21.8 22.7 33 43 62 75 100 125 160 180 210

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101

machining, metallurgical process such as case hardening, stress relieving, shrink filling etc., have to be carried out in many stages. Even the tungsten carbide inserts are made by sintering – a metallurgical process. 5.3.3.1

Materials

The first step in the manufacture of a tricone rock bit is to select the right materials for various components of the bit. Since a tricone bit is subjected to very heavy stresses, the tolerance levels in the composition of the materials to be chosen for bit manufacture are very low. The materials usually used for milled tooth tricone bits are given in Table 5.5. The materials mentioned in the table are readily available, but in composition their tolerance limits are generally wider, and hence unacceptable for manufacture of a tricone bit component. Therefore, it is necessary to choose very specific materials from selected suppliers. Choosing appropriate inserts for tungsten carbide insert type bits is equally important. Detailed information on tungsten carbide is given in Appendix 7 at the end of this book. The following rudimentary knowledge in this regard is given for the sake of completeness. A tungsten carbide insert is formed by sintering a mixture of tungsten carbide and cobalt. Tungsten carbide is extremely hard and wear-resistant but at the same time very brittle. Cobalt, which serves as binder between tungsten carbide particles, is necessary to give the desired toughness. The proportion of tungsten carbide in an insert varies between 84% and 96%. A higher proportion of tungsten carbide is better suited for the manufacture of hard formations bits. Modern manufacturing techniques have made it possible to make even better tungsten carbide inserts. Two such techniques are as follows. Tungsten carbide (TC) inserts prior to the 1980s had the same composition throughout their volume. With the new technique it is possible to use a composition rich in tungsten for the outer side of the insert and a composition with larger proportion of cobalt in the central core. Sintering in this manner results in an insert with high wear resistance at the surface and the desired level of toughness throughout its Table 5.5 Materials used for manufacture of various components of tricone bits. Component name

Desired material properties

Material chosen

Material composition

Cones

Abrasion and Impact Resistance Weldability, High Impact Resistance, Surface Fatigue Resistance High Strength, Impact Resistance Wear Resistance Wear Resistance Wear Resistance Extreme Abrasion Resistance

ANSI 4817

C, Si, Mn, Ni, Mo, S, P

ANSI 8720

C, Si, Mn, Ni, Cr, Mo, S, P

ANSI S2

C, Si, Mn, Mo

ANSI 431 ANSI M2 Chrome Cobalt Tungsten Carbide

C, Si, Mn, Ni, Cr C, W, Cr, Mo,Va Co, Cr, C, W, Ni W, C, Co

Legs

Roller and Ball Bearings Journal Bushing Thrust Button Bearing Hardfacing Tooth Hardfacing

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volume. Often materials like silicon carbide, boron nitride etc. are also mixed in the outer layer in small proportions so the insert surface gets added wear resistance. Another manufacturing technique makes it possible to make tungsten carbide inserts with an outer cover of diamond. Because diamond is the hardest and most wear resistant material, diamond coated tungsten carbide inserts have extreme wear resistance. Modern bits usually use flat diamond compacts of 2 to 3 mm thickness on a base of tungsten carbide for gage wear resistance. When the formations to be drilled are extremely hard and abrasive all the inserts are diamond coated. During drilling operations the bearings of the tricone bits – particularly the bush bearings – generate a very high quantum of heat. This heat is released to atmosphere in two ways as follows: 1 2

Convection of heat through compressed air used for bit cooling. Conduction of heat from the bearings to the surfaces of cones and legs, from where it is again transferred by compressed air through convection.

While drilling in very hard formations with extremely heavy bit loads, the quantum of heat generated is so high that compressed air flowing through the bearings is not able to transfer the heat at the desired transfer rate. Three decades ago some manufacturers used silver alloy inlays at the bush bearing portion, as shown in Figure 5.19. Silver being a far better medium for the transfer of heat, the inlays quickly brought the heat to the outer surface of the cones. Current techniques provide small diameter through holes in the cones. These holes serve as passages for compressed air and dissipate heat efficiently. 5.3.3.2

Heat treatment

Heat treatment is the key process that gives the essential properties to a component of a tricone bit. Many intricate heat treatment processes are required to be adapted during the manufacture of a tricone bit component.

Silver Inlays

Figure 5.19 Silver alloy inlays in a cone.

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103

To give surface wear resistance to a cone, the surface is carburized. Carburization is a process where components are packed in a carbonaceous carburizing compound in a closed box and heated in a furnace up to about 925°C. At such temperatures the carburizing compounds liberate carbon monoxide (CO) gas. Carbon in this gas penetrates the surface of the alloy steel component through the process of diffusion, and dissolves. As a result, the surface of the component becomes very hard and wear resistant. The thickness of such a hard surface lies between 1.7 mm to 3.3 mm depending upon the component and the time period over which it is kept in the carbon- based atmosphere. The central portion, where carbon is not absorbed, remains tough. Carbon content of the surface material varies between 0.7% to 1.2%. Figure 5.20 shows a portion of a cone where the darker carburized surface can be easily differentiated. Sometimes instead of using carbonaceous compounds, CO gas is used directly. The hardness level achieved in this manner is 63–65 Rockwell C. Certain portions of the surface are not to be carburized as they have to be machined further. In such cases, a coating is applied to that portion before it is carburized. The coating does not allow absorption of carbon at the surface beneath it. This type of selective carburizing can also done by heating the desired portion of the part by using a high intensity oxyacetylene flame of proper size and shape. Figure 5.21 shows a bit leg where the dark carburized surface is seen only on one portion that was not coated while carburizing. Metallurgical processes described here need very specialized equipment in the manufacturing plant. Quality control is of prime importance in the manufacturing process since performance of the bits is watched by the end users very carefully. Guarantees of bit performance have now become an almost unavoidable feature of bit sale.

5.3.4

Summary of design features

The chart in Figure 5.22 gives the qualitative summary of design features of various bits. The levels of different parameters indicated by bar graphs are variable between the two extreme values for the parameter viz. 0 to 10. Naturally, all the levels are

Hard Carburized Surface Tough Non Carburized Core

Figure 5.20 Cone carburization.

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Carburized Surface

Non Carburized Surface

Figure 5.21 Selective carburization. Bit Type Basic Design

Soft

Medium

Hard

Extra Hard

Offset Journal Angle

Cutting Structure Design

Scraping Action Crushing Action Tooth Depth Tooth Spacing Included Tooth Angle

Strength

Bearing Strength Carburized Case Depth

Metallurgy Tooth Hardfacing Gage Hardfacing

Figure 5.22 Summary of design features.

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Table 5.6 Recommended bit loads and rotary speeds for tricone bits. IADC code

Bit load kN/mm

Rotary speed in RPM

Steel tooth bits 112⇒142 212⇒242 312⇒342

0.175⇒0.525 0.525⇒0.875 0.700⇒1.225

70–120 60–100 50–80

TC insert bits 412⇒442 512⇒542 612⇒642 712⇒742 812⇒842

0.175⇒0.875 0.525⇒1.137 0.700⇒1.225 0.700⇒1.400 1.050⇒1.575

50–150 50–150 50–120 50–90 40–80

relative and not absolute. Similarly the hardness of the formation indicated by ten columns is also relative and not to any specific scale.

5.3.5

Bit load and rotation of tricone bits

As is amply evident from the discussion about the cutting action under the teeth of tricone bits, very high side forces are also exerted on tricone bits during the formation fracturing process. Naturally, tricone bits have to be rotated at low speed and high torque. With increasing feed force the torque requirement rises. To avoid shocks being experienced by the teeth, the rotary speed is required to be reduced. Normal recommended bit loads and rotary speeds to be used in tricone bit drilling are as shown in Table 5.6. For larger diameter bits, lower rotary speeds and higher bit loads should be chosen. An empirical equation put forth by Praillet for calculating optimum weight on the tricone bit is as follows. Wo = σc ∗ D/2 where Wo = Optimum weight on the bit in kg σc = Compressive strength of rock in kg/cm2 D = Diameter of the tricone bit in cm

5.3.6

IADC code

The International Association of Drilling Contractors (IADC) has formulated a standard four digit code for classification of tricone bits. It is explained in Figure 5.23. Almost every tricone bit manufactured by any manufacturer can be classified by choosing appropriate numerals or letters in place of the digits. This code is very suitable for comparing tricone bits and for unambiguous report preparation.

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First Digit

Used for Bit Type and Formation

Second Digit

Used for Further Formation Suitability

Third Digit

Used for Bearing Type

Fourth Digit

Used for Additional Design Features

Figure 5.23 Explanation of Four Digits in IADC Code.

Table 5.7 Key to first digit of IADC code. Number

Description (bit type and formation)

1 2 3 4 5

Steel Tooth Bit for Soft Formations Steel Tooth Bit for Medium Formations Steel Tooth Bit for Hard Formations TC Insert Bit for Soft Formations TC Insert Bit for Medium Soft Formations TC Insert Bit for Medium Formations TC Insert Bit for Medium Hard Formations TC Insert Bit for Hard Formations

6 7 8

The first digit of the code is a numeral which broadly indicates the type of bit and the formation hardness for which the bit is suitable. One of eight numerals can be chosen for this digit. Table 5.7 gives the key for selecting the appropriate numeral for the first digit. The second digit of the code is a numeral that broadly indicates the formation hardness for which the bit is suitable. One of four numerals can be chosen for this digit. Table 5.8 gives the key for selecting the appropriate numeral for this basis. The third digit of the code is a numeral that indicates the type of bearing used in the bit. One of seven numerals can be chosen for this digit. Table 5.9 gives the key for selecting the appropriate numeral according to this criterion. The fourth digit of the code is a letter that indicates the additional features provided in the bit. One of nine letters can be chosen for this digit. Table 5.10 gives the key for selecting the appropriate letter according to this criterion. Due to continuous developments in manufacturing techniques of tricone bits, most of the manufacturers restrict to first three digits of IADC codes for their bits and use the fourth digit as per a table made by them for their bits which can contain more letters than those presented in Table 5.10.

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Table 5.8 Key to second digit of IADC code. Number

Description (further formation suitability)

1 2 3 4

Soft Formations Medium Formations Hard Formations Extra Hard Formations

Table 5.9 Key to third digit of IADC Code. Number

Description (bearing type)

1 2 3

Standard Open Bearing Standard Open Bearing for Air Flushing Standard Open Bearing with TC Hardfacing or TC Insert Protection at the Cone Heel Roller Sealed Bearing Roller Sealed Bearing with TC Inserts at the Cone Heel Journal Sealed Bearing Journal Sealed Bearing with TC Inserts at the Cone Heel

4 5 6 7

Table 5.10 Key to fourth digit of IADC code.

5.4

Letter

Description (additional design feature)

A R C S D E G Z J

Air Flush Application Reinforced Welds Single Central Jet Standard Steel Tooth Model Deviation Control Feature Extended Jet Nozzle Extra Gage Protection Other Shapes of TC Inserts Jet Deflection

DULL BIT ANALYSIS

Observations about the conditions of a bit are made frequently during drilling. Eventually, like everything else in this world, a tricone bit comes to the end of its useful life. Such bit can be scrapped or sent for salvage, if economically feasible. Before doing so, it is very essential to carefully observe the condition of the failed bit and analyze the reasons for failure. Reasons for failure established by such analysis give good guidance to the user for achieving more efficient operations and better drilling economy in future. If such data are provided to bit manufacturers it can be very helpful in their research towards improved products.

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5.4.1

Bit observation

In any observation ‘when’ and ‘what’ are the most important criteria. As far as ‘when’ is concerned, there are three stages of bit observations viz. 1 2 3

Observation During Drilling Operation Observation After Bit Withdrawal Observation At Final Bit Rejection

5.4.1.1

Observation during drilling operation

During blasthole drilling operations the driller does not see the bit because it is inside the hole. He cannot even see the cuttings ejected out of the blasthole as the area is covered by a large dust hood. However, the driller is able to see the rotating drill pipe and can judge the penetration rate from axial movement of the drill pipe. On most modern blasthole drills a drilling recorder is provided in the operators cab. In addition to the information on other drilling parameters, the recorder almost invariably gives the current penetration rate in numerical units. From such indirect observations an experienced operator can judge if the bit is performing as per expectations or whether there is something wrong with the drilling bit. 5.4.1.2

Observation after bit withdrawal

It is a good practice to inspect a tricone bit after drilling a certain number of blastholes. As the bit usually remains within the dust hood, the hood has to be lifted and a torch has to be used for inspection of the bit. In many sophisticated and large blasthole drills, arrangements are provided for mechanical raising of the dust hood. The dust hood must be lifted before the drill moves to the location of a new blasthole so that the hood itself does not push the cuttings back into the blasthole. These curtain lifting arrangements come in very handy while inspecting the drill bit, cuttings and other associated factors. The best time to inspect the bit is when the drill has moved to the new blasthole position and is being leveled by using its jacks. In this type of inspection, the bit can be inspected for general wear condition of the cones, loss of gage i.e. reduction of the bit diameter, free rotation of the cones and broken teeth. 5.4.1.3

Observation at final bit rejection

When a bit starts giving a penetration rate below the acceptable level, it is replaced by a new bit. Observation of the rejected bit can be done far more thoroughly by considering the following conditions: 1 2 3 4

Tooth or Insert Conditions Cone Shell Conditions Bearing Conditions Bit Body Conditions

For the purpose of keeping records, the abbreviations shown in Table 5.11 are often used. As most of the abbreviations are acronyms of the condition to be observed,

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Table 5.11 Key for recording observations of a rejected bit. Tooth or insert conditions BT Broken Teeth or Inserts CD Cone Dragged CR Cored RG Rounded Gage WG Worn Out of Gage BG Broken Gage Insert Cone shell conditions BA Broken Axially BC Broken on Circumference BS Broken Spear point IH Cone Interference – Heel IM Cone Interference – Middle IN Cone Interference – Nose CS Cracked Axially CC Cracked on Circumference EC Eroded Cone shell

Rotation or bearing conditions BF Bearing Failure BP Broken Bearing Pin BR Broken Rollers CL Cone Locked IT Inward Thrust Wear LC Lost Cone LR Lost Rollers NF Nose Bearing Failure NL Nose Bearing Loose OT Outward Thrust Bit body conditions BL Bent Legs DB Damaged Bit SB Shirttail Broken SW Shirttail Worn

they are easy to remember. The table also gives a very clear idea as to what is required to be observed under each of the above conditions. Besides the reasons for bit rejection, as given in Table 5.11, a remark regarding how well the bit has performed, must also be recorded. For this purpose letters G+, G, A+, A, A−, P and P− are used to indicate very good, good, above average, average, below average, poor and very poor.

5.5

BIT RECORDS

It is very important to keep records of bit performance. For this purpose, a sample form duly filled is shown in Table 5.12. In the form, anything that has to appear on the original blank form, is printed on gray background. Whatever appears on white background is the data to be filled by the user. Sample data is filled up in these cells for guidance. All the data mentioned in the form, including the name of the company, drill models bit manufacturers etc., is hypothetical. Since the data have been chosen arbitrarily, two values of otherwise dependent factors may not even have any correlation. Therefore, the data should not be used for drawing any conclusions. Instructions for filling the form are as follows.

5.5.1

General data

All the details to be filled in are self explanatory. They can be obtained from geographical and meteorological records of the mine site.

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Table 5.12 Sample bit record form. Indian Copper Ltd.

Head Office - 1732, Uttung Towers, 9th Road, Andheri, Mumbai 400053. Mines - Anypore, Bihar; Malanjkhand, Madhya Pradesh.

Tricone Bit Record

Date :- April 14, 1997

General Data Mine Location

Malanjkhand

Block Name

17H

Altitude in m

1000

Approximate Temp. in °C

24

Waste Type

Granite

Ore Name

Chalcopyrite

UCS of Waste in MPa

220

UCS of Ore in MPa

100

DRI of Waste

38

DRI of Ore

68

Special Details of the Block

The block has very low percentage of intermixed ore and therefore all the blasted and excavated material will be treated as waste

Formation Data

Blasthole Drill And Accessories Details Make and Model

Bucyrus Intl. 39R

Drill Recorder and GPS

Drill Recorder Yes, GPa None

Maximum Bit Weight (kN)

540

Max. Rotary Speed (RPM)

160

Air Discharge (m3/min)

85.5

Air Pressure (kPa)

448

Drill Pipe Diameter (mm)

273

Single Pass Length (m)

13.7

Drill Bit Diameter (mm)

349

Stabilizer Type

Roller Stabilizer

Bit Performance Details Bit Make Bit Model

Serial No.

Meters Drilled

Assay. No.

Ore

Waste

Total

Hours in Use

Bit Weight kN

ROP in m/h

RPM

Perform. Class

Hughes

HH55

AA123

722K

-

2615

2615

38.5

67.92

360 - 380

50 - 60

G+

Smith

Q5JL

BQ117

573H

1320

1257

2577

48.25

53.41

350 - 370

60 - 70

A+

Reed

MCM

AJ338

119P

837

1570

2407

53

45.42

360 - 380

45 - 60

A

Security

M8M

ZZ101

357S

889

1805

2694

55.5

48.54

350 - 390

50 - 60

G

Bit Discard Details Serial Number

AA123

BQ117

AJ338

Discard Reasons

BF, RG

NF

SW, LR

ZZ101 BT

Cone Number

1

2

3

1

2

3

1

2

3

1

2

3

Bearing Condition

7

6

7

7

7

8

7

7

7

6

6

5

Spear point Condition

5

0

0

7

0

0

2B

C

1M

7

6

7

Inner Teeth Condition

5

5

5

6

6

6

1B1M

3B

C

5

6

6

6B1M

1B

8B2M

5

5

4

1B

1M

2B

1B

2B

2M

Gage Condition NOTES -

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5.5.2

111

Formation data

UCS means Unconfined Compressive Strength. DRI means Drilling Rate Index. Both these values are determined in a rock testing laboratory. In most of the mining operations the type of rock in which blastholes are drilled does not change very frequently. Hence the tests for UCS and DRI are carried out only when a substantial change is experienced in penetration rate, bit wear etc. In the item ‘Specific Details of the Block’, as many relevant details of the rock mass as possible are to be given.

5.5.3

Blasthole drill and accessories details

All the details to be filled in, are self-explanatory. These can be copied from the manufacturers’ details of the blasthole drill and accessories.

5.5.4

Bit performance details

Bit details, viz. bit type, serial no., assembly no., etc. are embossed on the pin as shown in Figure 5.8 from where they should be copied. ROP means the net rate of penetration which equals meters drilled divided by hours spent on actual drilling operation. Time for moving the drill to the blasthole location, leveling the drill, adding or removing the drill pipe etc. should not be included in the actual drilling operation hours. Bit weight should be mentioned in terms of kN i.e. if the pulldown ranged between 550 kN and 580 kN the figure to be entered in the column should be 550–580. RPM means rotary speed in revolutions per minute. The range should be mentioned by choosing the fastest and slowest speed. For entering proper values in the bit weight and RPM columns, data may be obtained from the drilling recorder provided in the blasthole drill. Abbreviations meant for indicating the performance class have been mentioned in earlier sections.

5.5.5

Bit discard details

Discard reasons should be mentioned by using the abbreviations given in Table 5.11. Bearing, spear point and inner teeth conditions should be mentioned for each of the three cones of the bit by estimating the amount of wear in terms of levels 0 to 8 (Table 5.12). Here 0 means virtually no wear and 8 means full wear. If some inserts of the bit have broken or are missing, the number of such inserts followed by B (broken) and M (missing) should be mentioned. Photographs of rejected bits may be taken in cases where the failure is abnormal and they should be stored systematically by modern methods of electronic data storage. For any additional details separate blank pages should be used. If a warranty claim for very poor performance of the bit is to be raised, the bit along with copies of the bit report form, corresponding drill recorder chart and the bit photograph must be sent to the manufacturer. The manufacturer is not likely to consider replacement without such information.

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Depending upon the manner in which bits have failed, it may be possible to repair and reuse some tricone bits. Many manufacturers do such types of refurbishing. Lastly, a discarded tricone rock bit may still have tungsten carbide inserts, silver alloy etc. and can fetch some salvage value. Therefore, it should be preserved for a meaningful salvage. 5.6

BIT FAILURE ANALYSIS AND REMEDY

Table 5.13 gives many types of bit failures along with their photographs and discusses the reasons of failure as well as some suggested remedial actions for achieving better performance. Table 5.13 Bit failure analysis and remedial measures. The teeth of this bit are worn out only at the surface that touched the bottom of the blasthole. Teeth on the inner portion of the cone are intact. Spiral marks can be easily observed at the worn out surface. This means that the bit failed prematurely because the rollers, somehow, got locked and drilling operation was carried out by friction scraping of the formation at the bottom of the hole much like that of a diamond bit. Tricone bits are primarily designed for crushing action with little scraping associated with it due to offset. Such failure can be easily noticed by an attentive operator because the penetration rate drops suddenly and the torque required for rotating the drill string increases suddenly. Wear on the gage teeth of this bit is much more than that on the inner teeth. Such failure pattern can be encountered when medium soft but very abrasive formations, such as sandstone, are drilled with a bit meant for drilling non abrasive formations. Obviously the selection of this bit must have been incorrect. A bit with better gage protection such as H or T gage with gage hardfacing or a milled tooth bit with TC inserts on the gage will give better performance in the formation. Excessive wear of the teeth can be seen on all the three cones. Further the wear is uniform from gage to the center of the cone as well as all around the cone. If the bit has failed prematurely the reason of failure must have been excessive weight on the bit. Excessive weight can result in quick wear of the tip of the teeth and once the carburized portion at the tip of the teeth is worn out the inner portion wears very easily as it is not hardened. Operating drill with lesser feed force may give better performance. Such a wear can be caused when the bit is rotated in hard and abrasive formations at a high speed without adequate feed force. Rotary speed in such conditions causes wear on the bit in a manner similar to grinding action. Performance of the bit will improve by reducing the rotary bit or alternatively by choosing a harder formation bit which has more number of teeth.

(Continued)

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Table 5.13 (Continued) In this bit very heavy wear on the shirttail and the gage inserts of left hand side cone is clearly visible. Such failure can be caused when the bit is coupled to a bent drill pipe. Rotary motion given to such a pipe always cause very heavy side thrust through the bit. In particular, one or two of the cones have to bear the load to a much greater extent than the others. Such failure can also occur if the carbide inserts of the bit brake loose and remain in a position just above the cone due to insufficient flushing. Remedial action will be in replacing the damaged drill pipe. In this bit the wear on all the three cones of the bit is very similar. The inserts on the gage of all the three cones have broken but comparatively much less wear can be seen on the inner cones. Such wear pattern can occur when a bit meant for drilling in the soft formation is used for drilling very hard formations. In such conditions the force to which the gage inserts are subjected is very high. Remedial action will be to choose a proper hard formation bit.

In this picture a vertical crack can be seen on the pin of the bit. Besides some damage to the threads of the pin shown by another bunch of arrows can also be seen in the middle of the pin. Such thread damage and cracks are formed when the bit is cross threaded to the drill pipe or the stabilizer and is used for drilling. Remedial measure will be to remove the drill pipe or the stabilizer with damaged threads of the box. Normally the box of the drill pipe or stabilizer. Bearing failure of the bit can be easily observed in upper photograph. The failure is so predominant that the rollers of the bearing have been lost. The bearing failure can be seen only on one of the three cones. For further investigations when the three legs of the bit were separated, it could be seen that the thrust button below the cone of which the bearing failed, had bent as shown in the lower photograph.This must have resulted in very little or no air circulation through the bearing. Consequently the bearings and the surrounding surface got hot and the carburized case as well as the bearings must have lost hardness.The journals of the bearings of other two cones were intact. A centrally located crack, all along the seam of the two legs, can be observed on the bit. Such a failure can occur if the bit was accidentally dropped to the bottom of the hole.The impact experienced by the bit was so high that the tensile stresses developed along the seam were higher than the strength of the weld. Alternatively such failure can occur if the stress relieving operations during the process of manufacturing the bit were not done or were not properly carried out. A centrally located crack, all along the seam of the two legs, can be observed on the bit. Such a failure can occur if the bit was accidentally dropped to the bottom of the hole. The impact experienced by the bit was so high that the tensile stresses developed along the seam were higher than the strength of the weld. Alternatively such failure can occur if the stress relieving operations were not done or were not properly carried out during the process of manufacturing the bit. (Continued)

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Table 5.13 (Continued) The cone of this bit has broken circumferentially. The nose and inner portion of the cone has been lost. The bearings under the cone have been exposed and some have been lost. When bearing of a bit fails the cones may interfere with each other and only one of the three cones gets subjected to full weight on the bit. In such circumstances this type of failure occurs. The cone of the bit has cracked near the shirt tail of the bit. This type of failure can take place when the bit is dropped in the blasthole with a heavy drill string weight on it. Such failure can also be caused by loading the bit with excessive bit weight in extremely hard formation. One of the three legs of the bit is completely broken and lost. This type of failure takes place when a bit is dropped or when it hits a ledge in the blasthole or when it hits. If the leg has a crack and that has remain undetected in the quality control checks the failure of this type can take place very easily. One cone of the bit is completely lost. The other cone is also broken circumferentially. This failure can be considered as advanced stage of cone breakage because a cone can be completely lost only when one cone breaks circumferentially. The root cause of this type of failure is the bearing failure which in turn is caused by when one of the three cones gets subjected to the full weight exerted on the bit.

5.7

DISCARDING DRILL BITS

Bits wear out when used for drilling. In some cases the loss of gage is intolerable or in other cases the penetration rate drops down to an unacceptable level. A driller must know what is meant by ‘intolerable’ or unacceptable. Drilling with a bit that has lost its gage will give lower blasthole volume. The quantity of explosive accommodated in the blasthole will thus be lower, and insufficient for achieving the desired level of fragmentation. Generally the tolerance level of a quantity of explosive in a blasthole is about 8%. Therefore, a drill bit should be discarded when the cross sectional area of the hole reduces to 92% of the original. This happens when the diameter of the bit reduces to about 0.96 times the original nominal diameter. It is very uneconomical to continue drilling with a bit that is giving low penetration rate because the blasthole drill that costs several hundred times more than the cost of a drill bit, is being utilized wastefully for merely saving the cost of a bit. Solutions to these situations are in sharpening the bit by grinding to give a faster penetration rate, or in a very few cases welding hardmetal by special electrodes. Blade type drag bits can easily be reconditioned by blade grinding. Hardmetal side welding to restore gage is a bit more difficult but can be tried. As such, blade bits can be used almost to complete destruction.

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115

Percentage of Penetration Rate of the New Bit

The teeth of a claw type drag bit can be changed when penetration rate is unacceptable. It is possible to grind these tips and reuse them, but in most cases this will result in a serious loss of gage so replacing the peripheral inserts is very essential. Steel tooth tricone bits are not subjected to grinding because in grinding the hardened surface is seriously weakened and cannot be restored. If a steel tooth bit is reused after grinding, it may give satisfactory penetration rate for a very short duration but very soon will become worth discarding. Thus, the reuse will not be worth the cost incurred in grinding and keeping the drill functionally idle while changing the bit. TC (tungsten carbide) insert tricone bits can be subjected to grinding if the wear on the inserts is very high and the bit is otherwise, i.e. from the viewpoint of bearing condition etc. in good condition. Figure 5.24 shows the manner in which the penetration rate of a tricone bit reduces with the increase in bit wear. The actual curve for a bit lies within the shaded area and the curvature is very similar to the upper or lower limiting curve. Averages of the data indicate that in the case of tricone bits, a completely worn out bit gives only 20% penetration rate as compared to the new bit. Whether to continue drilling till the bit is completely worn out, is more of an economical decision rather than a technical one. From the economic angle this topic has been discussed in the chapter on costing. It is not out of place to mention here that many bit manufacturers conduct short training courses about use of their drill bits in the field. Such training courses can also be arranged at the worksite office. Attempts should be made to arrange such a training course so that many drillers can take advantage of it.

100 80

60 40

20 0

0

20

40

60

80

100

Percentage of Bit Wear Figure 5.24 Reduction in penetration rate with bit wear.

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Chapter 6

Rotary blasthole drilling accessories

6.1

INTRODUCTION

Besides a drill bit, the other components in a rotary blasthole drill string are a stabilizer, drill pipes, a shock absorber and crossover subs. Apart from their specific individual functions, all these components have three common functions viz: 1 2 3

Transmitting the feed force exerted by the feed mechanism of the blasthole drill to the drill bit. Transmitting the rotary motion generated at the rotary head of the blasthole drill to the drill bit. Allowing a continuous flow of compressed air to the next lower component in the drill string without significant pressure loss.

During the transmission of feed force and torque, the components are subjected to very heavy vibrations and must be very strong in their fatigue life. Blasthole drilling operations also require some more items that are not part of drill string but are frequently needed for some specific purpose. Specific information on these accessories is given in the following elaboration.

6.2

DRILL PIPES

The drill pipe is a very important component of a drill string. In all the rotary blasthole drills, the drill pipes are stored in a pipe changer fitted in the mast and are mechanically added to the drill string as the drilling advances, or removed and stored back in the pipe changer when the drill string is being withdrawn after completion of the blasthole. Drill pipes used in rotary blasthole drilling are flush type – meaning that their outside diameter is the same through out their length. They have a box connection (i.e. female threads) at one end and a pin connection (i.e. male threads) at the other end. They have wrench flats near both their ends to facilitate easy coupling or uncoupling by the wrenches provided in the blasthole drill. Drill pipes are of two types – fabricated or integral. Both are illustrated in Figure 6.1.

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Upper Tool Joint Fabricated Drill Pipe

Lower Tool Joint

Lower Connection

Wrench Flat

Weld

Seamless Tube

Weld

Wrench Flat

Upper Connection Integral DrillPipe Upper Tool Joint

Figure 6.1 Fabricated and integral drill pipes.

6.2.1

Integral drill pipes

Integral drill pipes are made from solid alloy steel bars and as such they are in one piece. They have a central bore of uniform diameter through which compressed air flows to the bit. As the wall thickness of integral drill pipes is large, they are very heavy and sturdy. Due to large wall thickness they are stiff, and their susceptibility to buckling under heavy feed force is very low. Integral drill pipes are made from alloy steel, such as ANSI 4145, and are surface hardened by heat treatment. Compared to fabricated drill pipes, integral drill pipes are expensive but have much longer life. Single piece integral drill pipes are available in lengths up to about 12 m.

6.2.2

Fabricated drill pipes

Fabricated drill pipes have three pieces, viz. upper tool joint, seamless tube and lower tool joint, welded together. The upper and lower tool joints are made from alloy steels such as ANSI 4140 or 4145. The seamless tube is made from standard carbon steels such as A-106B or J55. Some manufacturers use an alloy steel tube of the same material as that of the tool joint so as to enable uniform heat treatment to be given to the complete drill pipe. Due to lesser weight and manufacturing cost, the price of fabricated drill pipes is relatively low. Fabricated drill pipes have a large diameter bore. Therefore air pressure loss in these pipes is slightly less compared to integral drill pipes. For large diameter drill pipes, tubes with larger wall thickness are chosen. Such thick tubes enable the drill pipes to withstand very heavy compressive and shear stresses generated by heavy feed force and torque exerted by the blasthole drill. Details of standard fabricated and integral drill pipes are given in Appendix 11 at the end of this book. The length of tubes used for fabricating drill pipes is normally limited to about 12 m. Longer drill pipes, of length ranging between 13.71 to 21.3 m length are also required for some blasthole drills. In such case many manufacturers use special welding techniques to weld two tubes end to end, and then add the tool joints to the long

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tube. This alternative is better than using two shorter drill pipes permanently coupled together. Tool joints, i.e. box or pin type end connections of the drill pipes, are either API Regular or BECO type. API Regular tool joints originate from the oilwell drilling industry. For a long time they were successfully used for drill pipes in the long rotary drill string used for oilwell drilling. In the beginning of rotary blasthole drilling, they were adopted for the short drill strings of rotary blasthole drilling also. In the late 1950s Bucyrus Erie Company developed BECO tool joint with a coarser BECO thread form. Profiles of the thread forms used for API and BECO tool joints are shown in Figure 6.2 and their dimensional details are given in Table 6.1 and 6.2 respectively. Profile and dimensional details of the API Regular tool joints are in Figure 6.3 and Table 6.3. Profile and dimensional details of the BECO tool joints are in Figure 6.4 and Table 6.4. API threads were made from the viewpoint of deep drilling with fluid circulation. They were fine threads, so the cross section of the spiral tunnel formed between apex API Threads

P Box

Srn rcn H

h

r

Position of Spiral Tunnel Fcs

Fcn 60°

r 30°

rcs

fcn Pin Axis of Thread

90°

BECO Threads Box

T

P H

Position of Spiral Tunnel

30° 30° r

Pin 90°

Figure 6.2 Thread forms used for API and BECO tool joints.

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Table 6.1 Dimensional details of thread forms used for API regular tool joints. Thread form

Taper %

H in mm

hn = hs in mm

Srn = Srs in mm

fcn = fcs in mm

Fcn = Fcs in mm

rcn = rcs in mm

r in mm

V.0.040 V.0.050

25 25

4.38 5.47

2.99 3.74

0.51 0.63

0.88 1.09

1.02 1.27

0.51 0.63

0.38 0.38

Table 6.2 Dimensional details of thread form used for BECO tool joints. Thread form

Taper %

H in mm

P in mm

T ± 0.0508 in mm

r in mm

BECO

25

7.366

12.7

2.7686

1.1938

A

F 5/ 8” 30°

G

D

C

H – Threads Per Inch

B

E G

K – Inclination of the Taper

Figure 6.3 Profile of API regular tool joint. Table 6.3 Dimensional details of the API regular tool joint and thread forms used (All dimensions in mm). API pin size

A

B

C

D

E

F

G

H

K

Thread form

2–3/8 2–7/8 3–1/2 4–1/2 5–1/2 6–5/8 7–5/8 8–5/8

3 3–1/2 3–3/4 4–1/4 4–3/4 5 5–1/4 5–3/8

3–1/8 3–3/4 4–1/4 5–1/2 6–3/4 7–3/4 8–7/8 10

2–5/8 3 3–1/2 4–5/8 5–33/64 6 7 7–61/64

1–7/8 2–1/8 2–9/16 3–9/16 4–21/64 5–5/32 5–11/16 6–19/64

2–11/16 3–1/16 3–9/16 4–11/16 5–37/64 6–1/16 7–1/16 8–1/64

3–3/8 3–7/8 4–1/8 4–5/8 5–1/8 5–3/8 5–5/8 5–3/4

1 1–1/4 1–1/2 2–1/4 2–3/4 3–1/2 4 4–3/4

5 5 5 5 4 4 4 4

3 3 3 3 3 2 3 3

V-.040 V-.040 V-.040 V-.040 V-.050 V-.050 V-.050 V-.050

of male threads and trough of female threads was very small and did not allow any significant leak of the circulating fluid that was flowing under very high pressure. As against this, BECO threads were made for blasthole drilling, where the pressure of compressed air was low. Thus, despite the larger cross section, the leakage in the spiral tunnel was sufficiently low.

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A

121

F 5/8” 30°

G

D

H – Threads Per Inch

B E

C

G

K – Inclination of the Taper

Figure 6.4 Profile of BECO tool joint.

Table 6.4 Dimensional details of the BECO tool joint and thread forms used (Dimensions are in mm unless specified otherwise). BECO tool size

A

B

C

D

E

F

G

H K Thread form

3″ BECO 3½″ BECO 4″ BECO 4½″ BECO 5¼″ BECO 6″ BECO 8″ BECO 10″ BECO 11″ BECO (11½″)

80 83 90 93 109 112 115 166 179

4½″ 5″ 5½″ 6¼″ 7½″ 8 5/8″ 10¾″ 13 3/8″ 15″

85.70 98.43 111.00 123.80 142.80 162.00 212.70 263.50 289.00

51.00 60.33 73.00 84.00 98.50 117.50 167.00 205.00 255.50

92.00 104.80 117.50 130.20 149.00 168.30 219.00 270.00 295.00

111.00 114.00 120.00 124.00 140.00 140.00 146.00 194.00 206.50

38.00 38.00 57.00 57.00 76.00 102.00 120.00 120.00 120.00

2 2 2 2 2 2 2 2 2

3 3 3 3 3 3 3 3 3

BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard BECO Standard

Comparatively very coarse BECO threads enable the tool joint to have the following advantages: 1 2 3 4 5

Ease in frequent coupling and uncoupling of the drill pipes. Rigidity in the tool joint to reduce vibration. Leakproof when they are lubricated with thick grease that fills the spiral tunnel. Excellent ability in withstanding rough use. Much higher maximum torque rating as can be seen from Table 6.5.

Since all the above aspects are of greater importance in rotary blasthole drilling, BECO type connections are becoming more popular in rotary blasthole drilling practice. On many occasions purchasers are unaware of the difference between the drill pipes used in the oil or waterwell drilling industry and those used in the blasthole drilling industry. Purchase of inexpensive drill pipes meant for the waterwell drilling industry and their use for rotary blasthole drilling work can prove disastrous. Hence to avoid such instances, the differences are elaborated in Table 6.6.

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Table 6.5 Max. torque ratings of BECO and API reg. tool joints. BECO tool joint

API regular tool joint Maximum rated torque

Maximum rated torque

Drill pipe diameter in

mm

Tool loint name

lbft

Nm

Tool joint name

lbft

Nm

3.500 4.000 4.500 5.000 5.500 6.000 6.250 6.500 6.625 7.000 7.625 8.625 9.250 9.625 10.750 11.750 12.750 13.375 15.000

88.9 101.6 114.3 127 139.7 152.4 158.75 165.1 168.275 177.8 193.675 219.075 234.95 244.475 273.05 298.45 323.85 339.72 381

None None None 3½ BECO 3½ BECO 3½ BECO 4½ BECO 4½ BECO 4½ BECO 5¼ BECO 5¼ BECO 6 BECO 6 BECO 7 BECO 8 BECO 8 BECO 8 BECO 10 BECO 12 BECO

Nap Nap Nap 18000 18000 18000 24300 24300 24300 36300 36300 55600 55600 88000 120000 120000 120000 120000 NS

Nap Nap Nap 24405 24405 24405 32946 32946 32946 49216 49216 75383 75383 119312 162698 162698 162698 222100 NS

2 3/8 API R 2 7/8 API R 3½ API R 3½ API R 4½ API R 4½ API R 4½ API R 4½ API R 5½ API R 5½ API R 5½ API R 6 5/8 API R 6 5/8 API R 6 5/8 API R 6 5/8 API R 6 5/8 API R 6 5/8 API R 7 5/8 API R 8 5/8 API R

NS 7000 9500 9500 16000 16000 16000 16000 32000 32000 32000 44000 44000 44000 44000 44000 44000 70000 98000

NS 9491 12880 12880 21693 21693 21693 21693 43386 43386 43386 59656 59656 59656 59656 59656 59656 94907 132870

6.2.3

Choice of a drill pipe

For efficient blasthole drilling at fastest penetration rates, choosing the most appropriate drill pipe is of extreme importance. The choice is mainly related to the diameter of blasthole. The diameter of a blasthole is decided on the basis of many factors related to production requirements, blasting method, parameters pertaining to loading and hauling equipment used in the mine, etc. They are elaborated elsewhere in this book. Factors to be considered for choosing the appropriate drill pipe are: 1 2 3 4 5 6

Drill Pipe Weight Drill Pipe Dimensional Parameters Drill Pipe Surface Treatment Size and Shape of Drill Cuttings Bailing Velocity in the Blasthole Drill Pipe Wall Thickness

A logical method to be adopted for choosing the most appropriate drill pipe is presented in Chapter 10 of this book.

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Table 6.6 Comparison of drill pipes used for blasthole drilling with those used for oil well/water well drilling. Aspect

Oil well or water well drilling

Blasthole drilling

Dimensional precision

Need not be so exact as the pipes are not handled by a mechanical pipe changer but are stacked loosely near the mast. Length is medium i.e. often of the order of 9 m. Often 3 drill pipes are stacked without being separated.

Has to be very precise as they are handled by a mechanical pipe changer. Length is long. Mostly more than 9 m. Often as much as 18 to 21 m. They are stored singularly in a pipe changer. Since the pipes are subjected to high compressive forces during drilling hence they are susceptible to buckling. For this reason the thickness of drill pipe has to be large so the moment of inertia increases. The weight of the drill pipe is very important as they are stored in a pipe rack and affect the stability of drill. Surface hardening treatment is very essential because the pipes are continuously subjected to sand blasting from the large rock cuttings that travel upwards through the annulus at a very high uphole velocity.

Length

Thickness

Are relatively thin as they are mainly subjected to high tensile stresses while drilling or very low compressive stresses while being stacked. There is no question of buckling of these drill pipes.

Weight

The weight of drill pipe does not matter as the drill pipes are stored on ground.

Wear Properties and surface hardening

The velocity of flushing water is low. The size of particles is also small. Hence these pipes are not subjected to very high external abrasion. Surface hardening treatment of the drill pipes is not very essential when used for water well drilling. In oil well drilling it may be necessary because when holes are deviated intentionally the pipes rub against the walls of the hole. Sealing properties are very important as the pressure inside the pipe bore and that in the annulus differ very greatly. Small pitch threads or special threads that have a very small cross section of the spiral tunnel at the apexes of the threads, are used for tool joints.

Threads

6.3

Sealing properties are not very important as the pressure inside the pipe bore and that in the annulus is nearly the same. BECO threads are preferred for tool joints.

STABILIZERS

Even though the depth of the blasthole is shallow, the characteristics of the formations lying on its alignment are not truly uniform. They change in terms of voids, cracks, compressive strength etc. If drilling is continued in a blasthole having formation conditions as shown in Figure 6.5, the direction of progress of the blasthole tends to deviate from the intended direction. Reasons for such deviation are: 1 2

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Male and female threads of the tool joint of the drill pipe are not perfect. Due to the bend caused, the feed force exerted on the drill bits is not truly axial. Drill pipe is relatively slender.

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Drill Pipe Upward Air Flow of Compressed Air Flow Used For Flushing Blasthole Wall

Drill Bit Soft Rock

Hard Rock Deviated Path of the Blasthole Intended Path of the Blasthole

Figure 6.5 Blasthole deviation.

3

4 5

A large gap between outer surface of the drill pipe and the walls of the hole may allow the drill pipe to buckle under heavy feed force exerted by the feed mechanism of the drill. Vibrations are caused by imperfect material balance in the drill string. The length that effectively resists hole deviation is only the short length of about 100 to 200 mm of the drill bit shoulder.

Besides deviation, the hole may also not be truly cylindrical as it assumes a spiral shape. Disadvantages of having deviated blastholes are: 1

2

3

The periphery i.e. the shirttail of the tricone bit undergoes non-uniform wear. The gap between the bit leg and the cone, meant for circulation of air through the bearings, may increase and small cuttings may enter into the gap. This causes bearings to wear. Withdrawal of the drill string and the tricone bit from such a drill hole requires more retract force. This results in heavy wear on the shirttail i.e. outer periphery of the drill bit leg. When the crooked or randomly deviated blastholes are filled with explosive, the distribution of the explosive in the rock mass is not uniform. This causes ineffective and non-uniform blast that results in non-uniform fragmentation and secondary blasting may have to be carried out frequently.

Stabilizers are used in the rotary blasthole drill string to lessen blasthole deviation and crookedness.

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A stabilizer is fixed just above the drill bit. Since the outside diameter of the stabilizer is only slightly less than that of a drill bit, the assembly of bit and stabilizer becomes more rigid. With such a close fit stabilizer, the length of drill string that resists deviation becomes much longer. This forces the progress of the blasthole much more closely in the intended direction than when the drill bit is used without a stabilizer. Since the outer periphery of the stabilizer is made with embedded tungsten carbide inserts, during rotation of the drill string these inserts abrade the walls of the hole and reduce hole crookedness. The stabilizer body is usually manufactured from ANSI 4145H Series Alloy or non–magnetic steel. Rotary blasthole drilling stabilizers are of four types: 1 2 3 4

Replaceable Sleeve Stabilizers Welded Blade Stabilizers Integral Blade Stabilizers Roller Stabilizers Details of these are given below.

6.3.1

Replaceable sleeve stabilizers

This type of stabilizer comprises a mandrel and a replaceable sleeve. The sleeve is threaded from inside and is coupled to the mandrel. The mandrel has a box connection to enable its coupling to a tricone bit as shown in Figure 6.6.

Top Connection

Mandrel

Sleeve Body in Elevation Tungsten Carbide Inserts on Sleeve Body Sleeve Body Cross Section

Thread Connection Between Mandrel and Sleeve

Sleeve

Tricone Drill Bit

Figure 6.6 Replaceable sleeve stabilizer stabilizers.

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The sleeve has four ribs as shown in the cross section of the sleeve in Figure 6.3. These ribs have large diameter tungsten carbide inserts. These tungsten carbide inserts come in contact with the wall of the blasthole and prevent other portions of the drill string from getting abraded. The sleeve can be replaced after it has worn out. Replaceable sleeve stabilizers have fallen out of favor with users for the following reasons: 1 2 3

Due to greater tooling and materials required for manufacture, their prices are higher. Their resistance to deviation is very limited. Assembling the two parts in the field is a time-consuming job.

6.3.2 Welded blade stabilizers Welded blade stabilizers are easier to manufacture and hence are relatively inexpensive. They have straight or spiral ribs as shown in Figure 6.7. These ribs have studded tungsten carbide inserts or carbide hardfacing on their outer periphery. Technologically, welded blade stabilizers are better than the replaceable sleeve stabilizers because they prevent deflection of a drill string over a longer length. It is also possible to rework and reuse them with inexpensive welding and grinding facilities.

6.3.3

Integral blade stabilizers

These are very similar in their shape and their look to the welded blade stabilizers, except that their ribs are integral with the main body and not welded later. Straight as well as spiral integral blade stabilizers are shown in Figure 6.8.

Figure 6.7 Welded blade stabilizers.

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Figure 6.8 Integral blade stabilizers.

Figure 6.9 Air flow in roller stabilizers.

6.3.4

Roller stabilizers

Roller stabilizers have three or more rotating elements called rollers that have studded tungsten carbide inserts. The rollers are cooled by compressed air that is made to flow through the central bore of the pin on which the rollers are mounted as illustrated in Figure 6.9. The moisture-laden air flowing upward through the annular space also cools the rollers. Technologically, roller stabilizers are better than the other types as they require lesser torque than the others.

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Use of rollers stabilizers is particularly beneficial in angle hole drilling, where other types of stabilizers give very high rotary resistance caused by drill string weight. If roller stabilizers are used in their place the resistance experienced is much less, as the rollers can rotate around their bearings. Stabilizers with three rollers are commonly used, but in extremely hard and abrasive formations stabilizers with even six rollers have been used. The advantages of roller stabilizers stated above become more prominent as the diameter is larger, hence they are available in large diameters. Roller stabilizers are more expensive than the other types. Details of some standard roller stabilizers are given in Appendix 12 at the end of this book. If the diameter of the blasthole is very large, say more than 381 mm, and if vertical holes are being drilled with drill pipes with large wall thickness, a stabilizer may not be required at all because the likelihood of hole deviation without the stabilizers is very low.

6.4

CROSSOVER SUBS

Crossover subs, also called ‘Threadsaver’ subs, are shown in Figure 6.10. The main reason behind the use of these subs is that when they are coupled at both ends of the drill pipe and used in drilling, the frequent coupling and uncoupling takes place at the threaded connection of the subs and not the drill pipes. This results in long life to the drill pipes that are far more expensive items. Another feature of these subs is that they can have two different types of threads at their two ends and so enable inclusion of tools with different threaded connection in the drill string. In blasthole drilling, a crossover sub is often located between the tricone roller bit that has API Regular Pin connection and the drill pipe or stabilizer that is often provided with the popular BECO threads.

6.5

SHOCK ABSORBERS

When a drill bit is fracturing the formation at the bottom of a blasthole, it is continuously subjected to erratic movements in the vertical as well as radial directions. Shock waves are generated in the drill string due to such movements, and also due to heavy feed force and high torque. Such shock waves are particularly predominant in rotary drilling with tricone bits. In blasthole drilling the rotary and feed

Figure 6.10 Crossover subs.

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force mechanisms are somewhat firmly attached to the blasthole drill. Hence the shock waves are transmitted from the rotary head to the other parts of the blasthole drill. Basically, shocks experienced by the drill string are of four types as shown in 1 2 3 4

In the axial direction of the drill string, In the direction of the blasthole drill length In the direction of the blasthole drill width In the direction of the drill string rotation

If the drilling process is considered minutely, then the reasons for generation of these shocks can be thought of as follows: A

B

C D

When the teeth of a drill bit crush the formation, the cuttings get displaced, and the teeth also move very slightly but suddenly in the direction of pulldown force and in the direction of rotation. This results in sudden reduction and rebuilding of compressive force, and sudden decrease and increase of rotary torque transmitted through the drill string. This creates vertical and circular shock waves that travel from the bit to the drill head through the drill string components. However perfectly manufactured they may be, all the drill string components have some acceptable tolerances in their axis, threads, materials etc. Thus, each and every item of the drill string deviates from a perfect axis either in direction or offset or both. Further, material distribution around this axis also changes in terms of inertia. Hence the drill string generates radial vibrations in a plane perpendicular to the drill string axis. By nature, all rocks are formed of small particles bound together. They are not ideally homogeneous. This creates shock waves of all the four types mentioned above. Often, large size rock fragments formed in the process of drilling do not get lifted with the available air drag. They greatly resist rotary motion momentarily when stuck in the annulus and release the resistance suddenly when they are further fragmented. This also creates a shock wave in the radial and circular directions of the drill string axis. Axis of Drillstring

Lateral Vibrations Rotary Vibrations Longitudinal Vibrations

Direction of Length of the Blasthole Drill Direction of Width of the Blasthole Drill Axial Vibrations

Figure 6.11 Modes of vibration in a drill string.

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All the four types of shocks reduce the fatigue life of the blasthole drill to a very great extent. Shocks in the axial direction of the drill string are particularly damaging. Shock absorbers are the devices which reduce the transmission of shocks to the rotary head and other components of blasthole drills. Advantages of installing a shock absorber in the drill string can be summarized as follows: 1 2 3 4 5 6

Increased life of the drill head components due to reduced shock waves. Slight increase in penetration rate. Reduced maintenance. Increased life of the drill pipes and bit due to jerkless feed force and rotary motion. Possibility of exerting more bit weight and rotational speed to give increase in the penetration rate. Reduced noise level due to reduction of metal-to-metal contact between vibrating elements in the blasthole drill components. There are two types of shock absorbers as follows:

1 2

External Shock Absorber In-the-Hole Shock Absorbers Details of these shock absorbers are as under.

6.5.1

External shock absorber

External shock absorbers are fitted in the drill string just below the rotary head, as shown in Figure 2.6. Two basic designs of shock absorbers have evolved. Both use synthetic rubber for shock absorption. Figure 6.12 illustrates the one made by Mining Technologies International (MTI). The complete assembly, shown on the left, consists of elements like the one shown on right. In this type of shock absorber, the axial shocks are absorbed by the shear stresses and strains developed in the rubber elements that are sandwiched between two drive lugs as shown in the figure. For absorbing the shocks in a plane perpendicular to the drill string axis the rubber element moves in compressive stress and strain. The construction of the MTI shock absorber is such that, if one or even two rubber elements fail, the other elements bear the load and the shock absorber continues to transmit the feed force and rotary motion to the drill pipes. The elements can then be replaced after the blasthole is completed. The effectiveness of the MTI shock absorber can be seen from Figure 6.13 that shows a vibrograph of the torque absorption. Another shock absorber, shown in Figure 6.14, is from Duraquest. It features a rubber element sandwiched between two plates. These plates have lugs that pass through the rubber element and holds all the components together. The rubber element absorbs all the four types of shocks with the help of

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Single Element

Vertical Shocks

Drive Lug

Horizontal Shocks

Drive Lug Rubber Element Shock Absorbing Elements

Vertical Shocks

Figure 6.12 MTI shock absorber.

Vibrations Recorded Before Installing a Shock Absorber

Vibrations Recorded After Installing a Shock Absorber

Figure 6.13 Effectiveness of MTI shock absorber in reducing shocks in rotary motion. Lugs

Complete Assembly

Top Plate Lugs

Rubber Element

Bottom Plate

Figure 6.14 Duraquest shock absorber.

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the lugs. The effectiveness of Duraquest shock absorbers is very evident from the vibrographs shown in Figure 6.15, which reveal that these shock absorbers can reduce vibrations to an extent of 79% in the axial direction, 10% in the radial direction parallel to the length of drill and 39% in the radial direction parallel to the width of the drill. Details of shock absorbers manufactured by MTI and Duraquest are given in Appendix 13 at the end of this book.

6.5.2

In the hole shock absorbers

These shock absorbers are attached at the bottom of the drill string just above the bit. They thus go in the hole, and hence the name. One shock absorber of this type is shown in Figure 6.16. These shock absorbers have a mechanical spring between two elements to absorb the shocks. The rotary motion and torque are transmitted through the splines of the two elements that intermesh with each other. In-the-hole shock absorbers are meant primarily for the oilwell drilling industry and are designed for absorbing shocks in the axial direction only. They are also used in horizontal directional drilling but very rarely in blasthole drilling. The blasthole drilling industry almost exclusively uses external shock absorbers for the following reasons: 1 2

3

4

External shock absorbers can absorb all types of shocks, whereas an in-the-hole shock absorber absorbs only the axial shocks. In blasthole drilling the drill head is very firmly fixed to the blasthole drill. Further, the drill string is very short and in heavy compression. Hence shocks transmitted to the blasthole drill components are of very high intensity. In oilwell drilling, the drill string is very long and much of it is in tension. Further, in oilwell drilling practice the drill string is hanging on wire ropes, hence the question of transmission of shocks to the drill components (except the rotary table) does not arise at all. So, in-the-hole shock absorbers are more suitable in oilwell drilling but not in blasthole drilling. An in-the-hole shock absorber goes inside the hole. In blasthole drilling it is continuously met with harsh sandblasting treatment from the drill cuttings flowing with the compressed air. Therefore an internal shock absorber wears very fast. The external shock absorber does not meet with any sandblasting. In blasthole drilling, failure of external shock absorber and subsequent remedial measures may be time-wasting but are not disastrous like the failure of an inthe-hole shock absorber. Shock absorbers have many fixtures to match with the fixtures of the rotary drill head. They should be selected for maximum rated bit weight of the drill because maximum bit weight is occasionally exerted in the process of drilling.

Many drill manufacturers provide shock absorbers of their own make on their drills. Most of these are very similar to those made by Atlas Copco. Eventually upon failure they can be replaced by shock absorbers made by specialist companies.

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Vibrographs on the right side were taken under identical drilling conditions but with and without Duraquest Smoothdrive”. Pulldown was 85000 lb. rotary speed was 90 rpm and average penetration rate was 0.80 ft/min. The upper vibrograph represents the vibrations at the drill head without Duraquest Smoothdrive, whereas the lower graph represents the vibrations at the drill head with Duraquest Smoothdrive. Comparison of the two vibrographs reveals following.

Aspect

With Smoothdrive

X Axis Average Acceleration

9g

5.5g

X Axis Peak Acceleration

12g

9g

Y Axis Average Acceleration

5g

4.5g

Y Axis Peak Acceleration

9g

11g

Z Axis Average Acceleration

12g

2.5g

Z Axis Peak Acceleration

22g

11g

Reduction of Vibrations in X Axis

(9g-5.5g)/9g = 39%

Reduction of Vibrations in Y Axis

(5g-4.5g)/5g = 10%

Reduction of Vibrations in Z Axis

(12g-2.5g)/12g = 79%

Figure 6.15 Effectiveness of duraquest shock absorber in reducing shocks in XYZ directions.


Without Smoothdrive

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Figure 6.16 In the hole shock absorber.

6.6

MISCELLANEOUS ACCESSORIES

Apart from the items that go into the drill string, many other items are used in rotary blasthole drilling. The following information about such items is given to bring completeness to this book.

6.6.1 Thread protectors Thread protectors are made up of nylon or similar stiff material or even cast steel. Many varieties exist. Their function is to protect the threads of the drill pipes or other threaded drill string components during transportation and handling. Male and female type thread protectors are shown in Figure 6.17.

6.6.2

Lifting bails and hoisting plugs

The main function of lifting bails is to enable easy handling of drill pipes and other heavy threaded drill string items from their horizontal position on the ground. To ensure that the lifted items are not unduly stressed they have a box or pin joint as shown. Lifting bails are made of cast steel. Hoisting plugs are meant to lift the drill string from the hole. They have bearings that enable rotation of the drill string components without rotating the wire rope used for lifting. This facilitates easy coupling and uncoupling of the drill string components. Hoisting plugs are seldom used in rotary blasthole drilling. Lifting bails as well as hoisting plugs for male and female threads are illustrated in Figure 6.18.

6.6.3

Blasthole plugs

Blastholes, particularly those with a large diameter, are very dangerous if left open. A serious injury or even fatal accident can result when some person or animal inadvertently steps on the cutting surrounding the blasthole that slides down to the hole. In very large blastholes someone going inside the hole to bottom cannot be unimaginable. Such incidents have happened in large diameter holes drilled for waterwells and left unattended at night. Besides, a blasthole may get filled up partially when the loosely heaped cuttings on the side slide into them. To prevent such incidents the best way is to plug the blastholes at the top. For this purpose, plastic plugs as shown in Figure 6.19, are used. They are made of plastic and have various sizes to match the diameters of blastholes.

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Figure 6.17 Thread protectors.

Figure 6.18 Lifting bails and hoisting plugs.

Figure 6.19 Blasthole plugs.

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6.6.4

Rotary deck bushing

The main purpose of rotary deck bushing is to provide horizontal stabilization to the drill string at the deck level of the blasthole drill. Rotary deck bushings are usually provided on the drill platform floor by the drill manufacturers. Their design and construction can, however, be a little less sophisticated in the sense that they may merely be non-rotating bushings. Such a simple friction bearing requires higher torque for drill string rotation and extra power, and they undergo rapid wear-out. They are also required to be lubricated from time to time. Besides, some blasthole drill manufacturers offer this as an optional item with their drills. Many manufacturers of rotary drilling accessories have developed more sophisticated rotary deck bushings that consist of an outer casing which fits on to the drill platform and the inner sleeve through which the drill pipe is passed down. Two or more rows of bearings allow the inner sleeve to rotate with drill pipe. The cross section of such rotary deck bushing is shown in Figure 6.20. When a drill pipe rotates and touches the inner sleeve due to radial vibratory movement, the inner sleeve also rotates as required. This reduces the torque requirement and wear. As less torque is spent in overcoming the friction, the drill bit gets more torque and penetration rate increases. Due to the soft inner sleeves of rotary deck bushings, the wear of the drill pipe reduces. These sleeves are relatively inexpensive and are replaceable.

6.6.5

Drill stem wrench

Early rotary blasthole drills used to have fork type spanners for uncoupling the drill pipes. They engaged at the wrench flats provided on the drill pipes. Due to several reasons these devices proved unsatisfactory. Most modern blasthole drills have one spanner to hold the drill pipe lying below the tool joint at its wrench flat and one wrench that grips the drill pipe lying above the tool joint at most of the cylindrical periphery. This wrench is called the drill stem wrench.

Top Flange to Suit the Drill

Outer Housing Adapter Bushing Inner Rotating Sleeve

Ball Bearings

Figure 6.20 Rotary deck bushing.

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Drill stem wrenches operate on hydraulic cylinders and are available in different levels of sophistication. They have jaws for gripping the drill pipes or accessories. One such wrench is shown in Figure 6.21. In the first part of the hydraulic cylinder stroke the wrench grips the drill pipe with jaws and in the second part of the stroke it loosens the drill pipe. Once the drill pipe is loosened the rest of the uncoupling can be done by rotating the drill head.

6.6.6

Bit breaker

Drill pipes usually have slots and a uniform diameter, hence while coupling or uncoupling they can be restrained from rotation by using a spanner or drill stem wrench. When a drill bit is required to be detached from the drill string, the spanner or drill stem wrench is of no use as the bit does not have slots or uniform diameter. For uncoupling the drill bit, a bit breaker as shown in Figure 6.22 is used. The bit breaker fits into the top frame of the dust hood where it is restrained from rotating. The drill bit in turn fits into the bit breaker and is also restrained from rotating. Thus, the bit can be uncoupled by counter-rotating the crossover sub or the stabilizer with the drill stem wrench.

6.6.7

Blasthole inclinometer

In early days, blasthole drills used to be equipped with simple pendulum type inclinometers. With these it was possible to align the mast for drilling inclined blastholes. The least count of such an inclinometer was about 1°.

Figure 6.21 Drill stem wrench.

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Figure 6.22 Bit Braker.

Later more sophisticated blasthole inclinometers were developed by some manufacturers. The precise angle measuring mechanism in such an inclinometer could measure the mast inclination angle up to an accuracy of 0.5° and show it by means of a digital display that could be read easily by the operator while sitting in the cab. One such inclinometer to be appropriately fitted on the mast is shown in Figure 6.23. An even more sophisticated inclinometer is shown in Figure 6.24. It comprises a vertical and two horizontal angle sensors. The horizontal sensors, fixed on the main frame of the blasthole drill, measure the angle of inclination of the main frame in the direction of drill length and drill width. The vertical sensor fixed on the mast shows the angle of the mast from the vertical direction. The digital display, central unit and control unit of such a system is kept in the operator’s cab. Such systems have even better accuracy – of the order of 0.2°. It is sometimes necessary to measure alignment and inclination of a blasthole after it has been drilled and the drill has moved away. For such measurements a hand-held inclinometer as shown in Figure 6.25 can be used. It can also be used for checking the correctness of mast alignment. In such inclinometers the magnetic compass, fitted on the top tilt arm of the instrument, can read angles with respect to north within an accuracy of 2°. For measuring blasthole inclination a torch, provided with the inclinometer and equipped with necessary fixtures, is lowered to the blasthole bottom. The torch light is then seen through the cylinder on the inclinometer. Pressing a button on the holding arm at the time when the torch light is fully visible, gives the angle of inclination with an accuracy of 0.1°.

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Figure 6.23 Inclinometer to be fitted on the mast.

Reading Unit Vertical Sensor Central Unit

Control Unit Horizontal Sensor

Power Supply Cable

Figure 6.24 Schematic of inclinometer system.

Even more advanced inclinometers, that record angle of inclination continuously while they are being lowered in the blasthole, are available.

6.6.8

Blasthole camera

Occasions when a blasthole camera needs to be used are rare, but it is essential for logging a blasthole. As it is an expensive item, it is usually hired. Most of the cameras are video type. While being lowered they store an image of the inner side of the blasthole continuously and show it on the screen.

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Figure 6.25 Hand held inclinometer.

These cameras have a cylindrical probe which consists of a light-emitting device and a camera that takes images through all 360° rotation. For taking images of the blasthole walls from the same distance, they are equipped with fixtures that align them centrally in the blasthole. Some cameras also have facilities of taking images on the front side so they can be of great help in locating a valuable item fallen into the blasthole. The image captured by the camera it is sent through a fiber optic cable to the main equipment on the surface of the ground. In this equipment, facilities for viewing the image through a television terminal as well as recording all the images on video cassette or a compact disk are available. One such camera is shown in Figure 6.26. In most cameras the images taken by the camera are immediately converted into electric signals and these signals are sent to the ground equipment. The signals themselves can be analog or digital. Analog or digital recording of these images can be done by using video cassette recorders or computer and hard disk through specialized computer software. Some advanced cameras have arrangements for being lowered into the blasthole in a controlled manner. This enables them to sense the depth, inclination, direction of inclination etc. and after analyzing the signals in a computer, a continuous log of the complete blasthole can be generated. The main control unit and the probe of such a camera, and photo log generated by it, are shown in Figure 6.27. The blasthole photo log shown in the figure is actually unwrapped in the sense that it is as if the inner surface of the blasthole is cut along a line parallel to the blasthole axis and then flattened on to a plane vertical surface. The probe containing the camera and the light-emitting diodes is shown in the left picture in Figure 6.27 while the right picture shows a photo log of a blasthole.

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Control Unit Casing and Guide

Reel

Probe

Camera

Figure 6.26 Blasthole camera system with fiber optic cable.

Figure 6.27 Digital blasthole camera and blasthole photo log.

6.6.9

Feed force measuring kit

This kit is needed for calibrating the feed force mechanism of a blasthole drill after some major repairs, or installations of additional items, or alterations in the drill. The kit is shown in Figure 6.28. The pressure unit of the kit is like a hydraulic jack and contains hydraulic oil in the enclosed space between the lower container and upper ram. For actual measurement of the feed force an adapter is fitted at the bottom of the first drill pipe instead of the drill bit. The pressure unit is placed on a firm surface below the adapter in such a way that the spherical surface of the adapter sits properly in the hemispherical groove on the top of the pressure unit. When feed force is exerted from the feed mechanism, pressure is developed in the hydraulic fluid. As the ram diameter is precisely known, the gage attached to the pressure unit gives direct reading of the force exerted. Unless the feed force mechanism on the blasthole drill is properly calibrated, the feed force measured by this kit, and indicated by the feed force reading in the blasthole, differ as shown in Figure 6.29. The chart is based on actual measurements on a well-known blasthole drill.

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Figure 6.28 Feed force measuring kit. 300

Bti Loading in kN

250 200 150 100

999.74

913.56

861.84

775.66

689.48

603.29

517.11

344.74

430.92

258.75

127.37

0.00

0

86.18

50

Pressure in kPa Indicated on the Pressure Gage in the Drill

Figure 6.29 Comparison of feed force, externally measured and indicated in a blasthole drill.

The reasons for differences can be due to one or more of the factors mentioned below. 1

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In any feed force mechanism the weight of the drill string and rotary head starts acting on the bit even before the feed force mechanism comes into action.

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2

3

143

A manufacturer may have calibrated the bit loading on the basis of a certain weight of drill head and drill pipes. Even if the weight of the rotary head is constant, the weight of the drill pipe changes when drill pipes are added. A manufacturer may use the weight of drill pipes with certain diameter and certain wall thickness for the calibration but the drill pipes used in actual drilling are different than those used for calibration.

It is very difficult to precisely calibrate the bit load indication system in the blasthole drills that do not have any microprocessor arrangements. When the drill is equipped with a microprocessor-based system all the reasons for inaccurate calibration can be wiped out and nearly perfect calibration can be achieved.

6.6.10

Blasthole dewatering pump

Conventional suction and discharge pumps have two limitations in dewatering blastholes as follows: 1 2

Atmospheric pressure limits their suction capabilities to 3 to 8 m. Such pumps do not throw out drill cuttings easily.

A compressed air-operated positive pressure lift pump, such as the one shown in Figure 6.30, can prove ideal for dewatering blastholes. The pump does not have any mechanical moving parts but is made up of a bellows and hose pipes. When the bellows are in a deflected state, the pump including the bellows is lowered into the blasthole in such a way that the hole below the bellows is immersed as deep in the water as possible while the bellows itself is above the water level. Compressed Air Ejection Hose for Water and Cuttings

Hose for Supplying Compressed Air to the Bellow and Blasthole Springs Meant For Deflating the Bellow

Bellow Compressed Air Pushing the Water Downward to the Bottom of the Central Hose Blasthole Wall

Water in Blasthole Under Compressed Air Pressure

Central Hose Through Which Water and Drill Cuttings are taken out

Figure 6.30 Schematic of blasthole dewatering pump.

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When the compressed air supply to the pump is started the bellows inflates and grips the walls of the hole. Thus, the blasthole below the bellows becomes airtight. As soon as this happens the compressed air is diverted to the lower side of the blasthole and compressed air pressure starts acting on the water surface in the airtight blasthole. Naturally, the water starts flowing up through the central hose till such time as the top level of the blasthole drops below the lowermost level of the central hose. After the blasthole is completely dewatered the compressed air supply is stopped. Due to the springs acting on the bellows the bellows deflates and can be withdrawn from the hole for dewatering the next blasthole or can be lowered further for dewatering the same blasthole to greater depth. This type of positive pressure lift pump can be used for dewatering very deep blastholes by using compressed air with adequate pressure. Since there are no obstacles in the central hose, the pump can lift cuttings along with the flowing water. Such a pump, however, is size-specific and can be used for holes of certain diameter range only. Advantages of positive pressure lift pumps are that they act very fast, lift cuttings of large size with the flowing water and dewater blastholes far more completely than other types of pumps. Sometimes diaphragm-type suction discharge pumps or submersible pumps are also used for dewatering blastholes.

6.6.11

Recovery tools

In very rare instances the drill pipe can break while rotary drilling is going on. In such an event a piece of drill pipe along with the drill bit remains inside the blasthole. Instances where a complete cone or part thereof has broken and remained in the blasthole, are also known. Whether the part of the drill string or the piece of broken bit is to be taken out and drilling is to be continued in the same blasthole, or whether the blasthole is to be discarded and another one drilled in the near vicinity, is a decision that changes from case to case and can be based on a little financial and technical common sense. In cases where the drill bit is almost at the end of its life and only a small part of drill pipe has remained in the blasthole, it may be worthwhile to abandon the hole, drill another blasthole at a small distance and use it for charging and blasting instead of the abandoned blasthole. If a blasthole is abandoned and a new one is chosen for charging, the subsequent blast can give rise to two possibilities as follows: 1 2

The possibility of very ineffective fragmentation in the vicinity of the abandoned blasthole and need for subsequent redrilling. The possibility of the steel pieces left in the abandoned blasthole flying to distances and damaging objects in a manner similar to a flyrock.

In many cases it is desirable to recover the broken part of the drill string. For this purpose many different types of tool, often referred to as fishing tools, are used. These tools are attached to the smaller size drill rods for ease in handling

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and are lowered into the blasthole to go inside the drill pipe bore or to cover it from outside. When pushed down and rotated, their sharp tap-like threads actually make threads in the drill pipe and the drill pipe gets attached to the fishing tool. Thereupon the rod string can be lifted so the drill pipe can be retrieved with the bit. One of the recovery tools is a recovery tap, meant to go within the central bore of the drill pipe. It is shown in left side picture of Figure 6.31. The other, meant to cover the drill pipe from outside, is called a die collar overshot. Fishing tools are case-hardened to have extremely high hardness to enable them to make threads in the drill pipes. Therefore they are sharp as well as brittle. They must be handled very carefully so as to avoid any impact with other objects.

6.6.12

Blasthole sampler

In the last three decades there has been virtually an explosion in the improvements in techniques of rotary blasthole drilling. Blasthole drills can now be precisely leveled, accurately trammed to the position of the next blasthole, and with automatic capture of all the drilling parameters such as load and torque exerted on the drill bit, penetration rate and vibration level experienced while drilling, flushing air flow etc. How this is achieved is the subject of one of the forthcoming chapters. However, such automation does not indicate the exact mineral composition of the material drilled by the drill bit. The mineral composition can only be known by visually observing the sample and later analyzing the sample in the laboratory. Therefore in many cases it becomes essential to obtain a sample of the material drilled. Collecting the samples manually is not possible because for such collection it is necessary to take pause drilling operations, open the hood and expose the collector to dust. To avoid all these hazardous and time-consuming operations a blasthole drill sampler has been developed. It is an attachment that can be fixed on a rotary blasthole drill, as shown in Figure 6.32.

Figure 6.31 Recovery taps (Left) and Die collar overshots (Right).

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Figure 6.32 Blasthole sampler.

In a recent patent filed in 2003 in the USA for the device, Harrison, the inventor of the device, has given its brief description as: “A sample collection system for collecting a drill cuttings sample from a hole being drilled during its drilling includes a stem collector surrounding the drill pipe. Pressurized air injected into the hole through the drill pipe forms an air-entrained drill cuttings stream which travels up the drill hole during drilling and which is directed by the stem collector into a conduit which directs stream to a sampling device. Sampling device samples drill cuttings stream and creates a stream of sample cuttings which is directed to a diffuser which separates the cuttings sample from the air stream. The cuttings sample passes into a sample collection chute from which it or portions thereof may be transferred into a sample container or bag. Stem collector seals against the ground surrounding the hole being drilled to direct substantially all drill cuttings to sampling device, thus substantially eliminating loss of fines prior to sampling and thus helps form a substantially representative sample. The sampling system can be mounted under the drill deck of a drilling rig.”

In prototype trials of the blasthole sampling device, the following things were noticed: 1

2 3

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The blasthole sampler may do the job of sample collection but it cannot be guaranteed that the sample is truly representative of the formation at a certain depth in the blasthole. This is because large cuttings from the bottom of the blasthole do erode the walls of the blasthole at upper parts of the blasthole and the eroded material mixes with the sample at the particular depth. In certain formations where voids exist in the ground, the flushing air is partly or totally lost. In such instance no sample is collected. In deep blastholes drilling has to be paused for short intervals.

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4

147

Most importantly, the blasthole sampler interferes with the blasthole flushing process and reduces the penetration rate that might be achievable without the blasthole sampler. Therefore, production department personnel do not like the idea of such a sampler being fitted on a production tool i.e. blasthole drill.

In view of all the above a blasthole sampler is not likely to be a popular attachment on a rotary blasthole drill.

6.6.13

Laser measuring instruments

In simple words these gadgets are a measuring instrument with a dedicated computer built into them. In this modern era, this instrument has put several other measuring instruments like old styled dumpy level, theodolite etc. in obsolescence. As far as blasthole drilling is concerned, a special version of a laser measuring instrument, called a burden finder has been made by MDL Systems Ltd. This handheld instrument looks like the one shown in Figure 6.33. The laser rays emitted from the instrument measure the distance of any target on which they strike. To actually measure the burden the surveyor has to stand on the bench floor in such a way that the vertical plane, in which he stands and takes all the measurements, is perpendicular to the row of blastholes. At first the surveyor aims at a point C that lies on the true crest line and stores the reading of distance OC and angle VOC. In the next step he aims at point T on the toe line and stores the readings of distance OT and angle VOT. All these points are shown in Figure 6.34. With these, angle COT is known by the instrument and it trigonometrically calculates the distance CT and angle OCT. Similarly, with the known bench height i.e. the distance CH, it calculates toe distance and face slope angle. Now referring to Figure 6.35, the distance between the true crest line (point C) and the blasthole row, which actually equals to the intended burden, is known and the inclination of the blasthole is also known. When all the above information is stored in the computer that is built into the instrument, the coordinates of the vital points are defined and the process of measuring actual burden can be started. When the laser is aimed at any point on the bench face as

Figure 6.33 Burden finder (Laser measuring instrument).

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C Crest

Face Slo

V

pe Len gth

Face Slope Angle

O

T

H

Toe Toe Distance

Figure 6.34 Measurement for basic data input.

Intended Burden

C

Actual burden

Angle of Inclination

Figure 6.35 Measurement for basic data input.

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shown in Figure 6.35 it is immediately defined in terms of its coordinates with reference to the operator. This enables the calculation of actual burden. Burden Finder does the calculations and shows the results on the screen provided on the side of the instrument. Some instruments have a communication port that enables printing the data on paper or transferring the data to a computer. Sometimes it is necessary to map the whole face of the bench. In such cases use of a handheld Burden Finder is not advisable as it is meant to take reading only in 2D. For such measurements a 3D instrument named Quarryman Pro is available. It is mounted on a tripod as shown in Figure 6.36. As Quarryman Pro is mounted on a tripod, it can turn in accurate angles in both vertical and horizontal directions. It is motorized and is turned in the horizontal and vertical axes by the motors in the instruments. With this facility and the built in computer, the Quarryman Pro can take measurements of 250 points per second

Figure 6.36 Quarryman Pro.

Figure 6.37 Profiling of a bench face with Quarryman Pro.

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on the face of the bench. The distance between the scan lines can be defined in the instrument. With large distance between scan lines, a large bench face can be mapped with somewhat less accuracy. For some part of the bench face where a high resolution scan is needed, the distance between scan lines can be kept very low. This is clear from Figure 6.37.

6.7

MISCELLANEOUS SAFETY ITEMS

Several miscellaneous items are required by the driller, the assistant, or the blaster who carries out blasting operations subsequently. Such items are expected to be provided for them by their employer i.e. the owner of the drill. Some of these items and their functions are as follows. Some more items are described in chapter 22 that pertain to blasting consumables and accessories

6.7.1

Close fitting shirt and long pants

If the clothes worn by the operators while on the drill are loose, there is a chance that they may get entangled with a lever and operate it inadvertently. Such clothes can easily get gripped within moving parts and pull the operator to cause him injury. To avoid these possibility close fitting shirts and long pants must be used by the operators.

6.7.2

Safety glasses

To avoid tiny flying objects, such as very small rock pieces, dust formed in grinding, tiny oil drops etc., hitting the eye, safety glasses to cover both the eyes from all sides are essential.

6.7.3

Safety toed shoes

These are needed by the operators to protect their feet and particularly toes from a hand tool falling on the feet.

6.7.4

Ear plugs or ear muffs

One of these two items is needed by the operators to reduce the nuisance of noise. Ear muffs cover the ear whereas the earplugs go inside the ear and close them. These items reduce noise intensity by as much as 22 db. They are shown in Figure 6.38.

6.7.5

Safety vest

These vests not only offer protection to the operator’s chest, back and abdomen but because of the orange fluorescent color the operator becomes easily visible in low light.

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Figure 6.38 Ear plugs and ear muffs.

6.7.6

Respirator

A respirator is occasionally needed by the operator to filter and allow only clean air to his nose for breathing.

6.7.7

Rain coat

These are needed when the operator or helper have to go out of the cabin to protect from rain, snow stormy wind etc.

6.7.8

Face shield

On a few occasions the driller or the assistant has to do some welding. For such activity the safety glasses are not sufficient and a face shield is required for safety of the eyes as well as the face.

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Chapter 7

Rotary blasthole drills

7.1

INTRODUCTION

The era of churn drills in the mining and quarrying industry started declining with the introduction of the rotary blasthole drill in the late 1940s. Soon after the introduction of the model 50R top drive rotary drill by Bucyrus Erie Company in 1952, most of the mines shifted to rotary drilling and the era of churn drills ended by about 1960. In the early days, rotary blasthole drills were used in surface mines and in large size quarries for drilling soft and medium-hard formations. In the late 1950s and early 1960s many equipment manufacturers started producing bigger surface mining equipment such as shovels, draglines, dumpers etc. To match with such equipment, larger blasthole drills were developed by Bucyrus Erie, Marion, Joy and Gardner Denver. Blastholes of diameters up to 381 mm could be satisfactorily drilled by such drills even in hard formations. Today gigantic blasthole drills, e.g. Bucyrus 59R or 49R, P&H 120A, Atlas Copco PV351 etc., are rated for blastholes of diameter 406 mm or larger. Their tall mast enables drilling a hole to a depth of about 20 m with the first drill pipe itself. Price of such blasthole drills can reach US$ 6,000,000. This chapter is devoted to explaining various layouts and components used on rotary blasthole drills of various types and sizes.

7.2 ASSEMBLIES IN ROTARY BLASTHOLE DRILLS A rotary blasthole drill is made by putting together many assemblies. These assemblies are very costly, have many small parts and can attain very long life. They are carefully chosen by the design engineers together with the manufacturers of the blasthole drill. Most of these assemblies can be repaired by changing a few parts. Assemblies mounted on a rotary blasthole drill have specific purposes as listed in Table 7.1. Accessories, elaborated in chapter 6, are not affixed to the blasthole drill. They are comparatively inexpensive, have no or few parts, and a relatively short life. Accessories are chosen by the user. In most cases, they cannot be repaired satisfactorily.

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Table 7.1 Common assemblies on rotary blasthole drills and purpose behind their inclusion. Assembly

Purpose

Undercarriage

Undercarriage enables the movement of the machine from one hole to the other or in some cases from one worksite to the other.

Main Frame

Excepting the undercarriage all the other assemblies of a rotary blasthole drill are mounted on the top of a very sturdy frame called main frame.

Leveling Jacks

Leveling jacks, attached to the main frame, are meant for leveling the machine after it moves to the location of the hole to be drilled.

Prime Mover

A prime mover is the main source of power. All the driven components in the machine are driven by use of this power source so that they generate desired movements of the components. Diesel engines are used as prime movers in many drills. In some very large rotary blasthole drills electric power supplied to the mine is directly supplied to the drill from where it is distributed to various components.

Air Compressor

Air compressor is meant for compressing the atmospheric air to a preset pressure and circulating it to the bottom of the hole through the drillstring components for flushing the blasthole and removing cuttings formed in the process of formation fracture.

Operator Cab Operator’s cab, located at the rear of a rotary blasthole drill, has all the necessary devices for observing and controlling all the necessary operations required to be performed during the process of drilling blastholes.

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Driver Cab

In truck mounted rotary blasthole drills the operator’s cab is located at the rear end. It is, therefore, not suitable for moving the drill on the road over long distance at higher speeds. In such case a separate driver’s cab on the front – just like that of a heavy truck – becomes essential.

Mast

Mast accommodates the rotary head assembly and the feed force mechanism that enables linear movement of the drill head with necessary feed force. By suitably aligning the mast in vertical or angled position, drill string can be moved and blasthole can be drilled in the intended blasthole alignment.

Auxiliary Winch

Most of the rotary blasthole drills are equipped with a winch. Its wire rope enables handling the heavy accessories from the top of the mast.

Rotary Head

Rotary head rotates the drill string at desired rotary speeds by exerting the required torque.

Pipe Rack

Pipe Rack – also known by many different names – is built into the mast. It enables storing and adding or removing the drill pipes to the drill string mechanically. This gives speed and accuracy in the operations.

Hydraulic System

Only a hydraulic system is capable of generating linear motion with very high force required for such operations as mast raising/lowering, leveling the machine by hydraulic jacks etc. Every blasthole drill, therefore, consists of hydraulic systems. Besides, in many rotary blasthole drills, hydraulic system also supplies high pressure hydraulic oil to many hydraulic motors to generate rotary motion.

Dust Control Equipment

During drilling operation, huge quantity of very fine dust is generated at the bottom of the hole. It is ejected out of the hole with the circulating compressed air. To prevent such dust from mixing with the atmosphere and polluting it, a dust hood is provided on the drill. However, due to many reasons the dust hood proves insufficient in preventing mixing of the dust with atmosphere. This makes it essential to incorporate dust control equipment in a blasthole drill. It is of two types viz. Water injection or Dry dust control.

Machinery House

Machinery house is an enclosure that shelters many components and assemblies within it so as to protect them from heat, cold, rain, dust and flying rocks. It also creates pleasant surrounding to the personnel to safely repair or replace the components.

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7.2.1 Assembly placement Three major groups of assemblies, viz. lower, upper, and mast, are generally recognized in a blasthole drill and are shown in Figure 7.1. Lower refers to the lowermost part of the blasthole drill. It comprises of ground support components, viz. a frame to hold the components together, two crawler tracks and the components that impart movement to the blasthole drill through the frame. Upper refers to a main frame that is placed on the top of the lower frame and contains many important assemblies viz. leveling jacks, prime mover, air compressor, operator’s cab, catwalks, hydraulic system components, dust control equipment, machinery house (if provided) etc. Figure 7.2 shows the configuration of the components on the upper. The mast is the tall, tower-like structure of a blasthole drill that is supported by the main frame. A mast consists of a rotary head, feed force mechanism, pipe changer, wrenches for coupling and uncoupling the drill pipes, auxiliary winch etc. Details of the layout of a blasthole drill and placement of assemblies are given later in this chapter.

Mast

Upper

Lower

Figure 7.1 Assembly groups in a rotary blasthole drill.

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Operator’s Cab

Engine

Air Compressor

Hydraulic Oil Tank

Dust Collector

Engine Exhaust Radiator Catwalks

Railings of Catwalks Hydraulic Oil Cooler

Figure 7.2 Configuration of assemblies on upper frame.

7.3 TYPES OF ROTARY BLASTHOLE DRILLS Rotary blasthole drills are differentiated into various types on the basis of diverse criteria. If the suitability of a rotary blasthole drill for drilling a particular class of blastholes is used as a criterion, it can be classified as conventional or special purpose. If the type of undercarriage of the rotary blasthole drill that enables its movement on the ground is taken as a criterion, it can be distinguished as either crawler-mounted or carrier-mounted. When the power source and distribution in a rotary blasthole drill is the basis of a class, diesel-hydraulic, diesel-electric, electric-hydraulic or electric-electric are the four types. Of the two words in the class names, the first refers to source and the second to the mode of power distribution. A generalized classification of rotary blasthole drills that takes into account all the four classification criteria is shown in Figure 7.3. The capability of a rotary blasthole drill in exerting feed force (FF) on the drill bit through its rotary head and the drill string components is often taken as a criterion. On this basis a blasthole drill can be placed in one of the following types: Small (100 kN < FF < 225 kN), Medium (226 kN < FF < 350 kN), Large (351 kN < FF < 500 kN), and Extra Large (FF > 501 kN)

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157

Blasthole Drills

Special Purpose

Conventional

These drills are diesel hydraulic type and of small size

Crawler Mounted

Diesel Hydraulic These drills can be small, medium, large or even extra large

Diesel Electric Nowadays these drills are almost extinct

Carrier Mounted

Electric Hydraulic

ElectricElectric

These drills can be small, medium, large or even extra large

These drills are usually extra large

Figure 7.3 Generalized classification of rotary blasthole drills.

7.3.1

Special purpose rotary blasthole drills

In very soft formations, e.g. a thick coal seam in a surface coal mine, a potash bed etc., blastholes of diameters ranging between 152 to 229 mm are required to be drilled at close spacing so the blast will give very good fragmentation and the blasted material can be easily loaded by wheel loaders and transported in coal haulers or dumpers. A few manufacturers have designed blasthole drills specially for drilling such blastholes. Such rotary blasthole drills are often referred to as ‘Single Pass Drills’ because they have only a single long Kelly instead of a drill pipe. A tricone or a drag bit is attached to the bottom of the Kelly through a sub. The Kelly is rotated by means of a rotary table placed on the main frame of the drill. Load on the bit is exerted through the Kelly which is pushed down through the feed force mechanism built into the mast. The use of such blasthole drills proves economical because the conventional blasthole drills, meant for attaining an equal hole depth in a single pass, are too expensive. A typical single-pass blasthole drill, the Atlas Copco model DMLSP, is illustrated in Figure 7.4. Different models of single pass drills are available to suit different hole depths and diameters. It can be easily observed that the single pass drill shown in Figure 7.4 looks far smaller but taller than the conventional drill capable of drilling similar hole diameters shown in Figure 7.1.

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Figure 7.4 Special single pass blasthole drill.

The mast of a special single pass rotary blasthole drill is very tall to accommodate the long kelly. The cross section of the mast is very small for the following reasons: 1 2

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The drill does not have any pipe changer and the mast is not required to support it. The rotary torque on the drill bit is exerted by the rotary table fixed on the main frame instead of a moving rotary head in the mast. Thus, the mast need not be designed for heavy twisting reaction to the torque exerted by the moving rotary heads as in conventional drills.

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3 4

5

159

As the Kelly drills the blasthole in a single pass, no coupling and uncoupling wrench is involved. As the drill is basically meant for drilling in soft formations, limited feed force is to be exerted. A 250 mm blasthole can be drilled with about 225 kN feed force. If the same size blasthole is to be drilled in very hard formations the rotary blasthole drill must be capable of exerting a feed force of about 350 kN. Feed force is exerted by hydraulic motors and chains which weigh far less as compared with the hydraulic cylinder and chain or rope mechanism in conventional drills.

The mast designed by taking into consideration all the above factors is lightweight, and the drill can be made stable without having any need for heavy main frame members, ballast, or a wide and long crawler base for the purpose of stability of the drill. Further, as a special single-pass rotary blasthole drill is not meant for drilling holes over a wide range of diameters, there is no choice of many alternative engines or compressors. The designer can therefore place the assemblies on the main frame very closely. Some time ago, it was thought that blasting of horizontal blastholes at the bench toe in addition to normal vertical blastholes would result in better fragmentation, smoother bench surface, etc. To cope with such requirement of drilling horizontal blastholes, a special twin mast drill, as shown in Figure 7.5, was developed by Joy Manufacturing Company during

Figure 7.5 Twin mast rotary blasthole drill for drilling horizontal blastholes.

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early 1970s. This drill could simultaneously drill two blastholes of size up to 270 mm in hard formation. It was very heavy as it had to withstand high horizontal reaction. Such drills, however, never became popular because: 1 2

Small horizontal blastholes of 100 to 127 mm diameter were found to be better for the purpose than the 250 to 270 mm blastholes drilled by using the twin mast drill. Horizontal holes of diameter to 127 mm could be drilled at a much faster speed by small sized, comparatively inexpensive rotary percussion drills rather than the bulky and heavy twin mast rotary blasthole drill.

Some designers also thought of a rotary blasthole drill with two masts on one frame so two blastholes can be drilled simultaneously. This concept, however, proved impractical for large rotary blasthole drills. One rotary blasthole drill with two masts meant for drilling two small diameter holes simultaneously is manufactured in South Africa by Smith Capital Equipment.

7.3.2

Conventional rotary blasthole drills

Conventional rotary blasthole drills are designed to be useful for drilling holes over a wide range of drilling parameters viz. hole diameter, hole depth, hole inclinations, formation hardness etc. Almost all the manufacturers have a range of models of such drills. They can be equipped with many options available to suit specific drilling conditions. 7.3.2.1

Crawler mounted rotary blasthole drills

Crawler-mounted blasthole drills are very popular. More than 98% of the rotary blasthole drills purchased all over the world are crawler type. These drills are very sturdy and rugged. In most of the cases they are used for round-the-clock operation throughout the year. Primarily they are used in surface mining operations. 7.3.2.2

Carrier mounted rotary blasthole drills

In the 1970s and 80s a few carrier-mounted blasthole drills were chosen by some mines for surface drilling operations. For many reasons they proved unsatisfactory. Today hardly anybody ventures in using carrier-mounted blasthole drills for such mining operations. Carrier-mounted rotary drills are very suitable for drilling waterwells that are usually located at long distances from each other. Therefore, on-the-road mobility of the drill is of prime importance. Although termed as rotary, these drills are more often used for rotary percussion drilling with down-the-hole hammers. Many manufacturers publish brochures of carrier-mounted rotary blasthole drills but they are essentially waterwell drills that can also be used for drilling blastholes. In any case, the size of these so-called blasthole drills is small. They can be used for blastholes of diameters up to about 170 mm. An extra large carrier-mounted rotary blasthole drill, Atlas Copco Model T4BH, that is sometimes used for blasthole drilling, is shown in Figure 7.6.

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161

Figure 7.6 Carrier mounted rotary blasthole drill.

7.3.2.3

Comparison of crawler vs. carrier

Why almost all the users select a crawler-mounted rotary blasthole drill in preference to carrier mounted blasthole drill can be understood from the detailed comparisons of the two types made in Table 7.2.

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Table 7.2 Comparison between crawler mounted and truck mounted rotary blasthole drills. Aspect

Crawler mounted

Road Worthiness Virtually nil due to large size, huge weight and slow speed of the crawler undercarriage.

Carrier mounted Designed to travel on most of the paved roads by taking into account proper weight distribution on the axles and other essential features.

Ability of Long Travel

Very poor due to very low travel speed Very good due to high travel speed that is often lower than 4 km/h that ranges between 50 to 70 km/h

Power Source

Diesel or electric power sources are easily possible

Can only be diesel driven. Electric power can not be used.

Width

Wide, usually between 3 m for small drills to 8.5 m for extra large drills

Narrow, often less than 2.4 m or even in extreme cases less than 3 m as the drills are required to travel on regular paved roads.

Weight Range

Ranges between 25000 kg for small Limited to about 60000 kg even in drills to 184000 kg for extra large drills. case of largest drills. The drill must adhere to 20000 kg per axle as per road worthiness for usual roads.

Rough Terrain Ability

Excellent due to crawler undercarriage that gives low ground pressure of the order of 100 kN/m2 and widely spaced long tracks.

Very poor due to wheeled undercarriage that gives high ground pressure of the order of 600 kN/m2 and narrow width.

Feed Force Ability Can exert very high feed force due to their heavy weights. The usual range is between 110 kN for small to 600 kN for extra large drills

Can not exert high feed force due to limited weights. The usual range is between 80 kN for small to 300 kN for extra large drills.

Mast Height

Masts can not be tall. Heights are usually limited to 12 m.

Masts can be very tall. Heights can reach to height in excess of 27 m in case of extra large drills.

Longest Drill Pipe Long drill pipes of about 18 m length Drill pipes are short. Limited to about are possible in many medium size drills. 7.5 m in almost all the drills. Can be even 21 m in case of extra large drills. Dust Suppression Both dust collector and water injectors Water injection is easily possible. are easily possible. Dust collectors are difficult to mount. Stability

Excellent because heavy weight in the lower and upper. Can travel with raised mast on the rough surface at the top of mine bench.

Maintenance Cost Relatively low maintenance cost due to crawler under carriage. Most of the machines can be equipped with automatic lubrication system.

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Poor because of limited weight and limited width. Traveling with raised mast on the rough surface of a mine bench is dangerous. Relatively high maintenance cost due to tired under carriage. Machines usually can not be equipped with auto lubrication system

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7.4

163

LAYOUTS OF ROTARY BLASTHOLE DRILLS

The layout of assemblies and components of a rotary blasthole drill differs from manufacturer to manufacturer as well as model to model. The objectives behind an ideal layout are given below. 1 2 3

4

There should be adequate space for repair or maintenance of assembly/component. Removal or replacement of the assembly or component should be possible without removing other assemblies or components. The length of pipes, hoses and cables used for carrying the hydraulic oils, compressed air or electric current should be minimum so the replacement cost becomes low and pressure or voltage losses are minimized. The position of the combined center of gravity of all the assemblies/components should be such that the blasthole drill will remain sufficiently stable in different positions of the mast that prevail while carrying out drilling operations, or while the blasthole drill moves from one location of hole to another.

Layouts of typical diesel engine-driven carrier-mounted and crawler-mounted blasthole drills are shown Figure 7.7. Figure 7.8 shows layout of a typical all-electric rotary blasthole drill.

7.5

DETAILS OF ASSEMBLIES

There is no specific standard that stipulates what assemblies or devices should be provided as standard and what should be optional in a blasthole drill. In some rotary blasthole drills some of the assemblies mentioned in the table are not provided as standard but are available as options at extra cost. Here is some additional elaboration.

7.5.1 Assemblies in the lower group This group is also called an undercarriage. It includes the items required to support the upper and the mast and enable movement of the blasthole drill. As said earlier, crawler and carrier are the two types of undercarriage used in rotary blasthole drills. 7.5.1.1

Carrier undercarriage

Carrier undercarriages are not very common in rotary blasthole drills. They are described briefly. The prime objective behind choosing a carrier undercarriage to a drill is to make it roadworthy and capable of traveling at fast speed on paved roads. A crawler undercarriage is never considered as roadworthy as the point loads between track and the surface of a paved road destroy the surface of the paved road. Further, for many reasons, they cannot be easily made to travel at fast speed. A carrier undercarriage, since it has to travel on paved roads, must comply to all the regulations stipulated for being roadworthy. These regulations change from country to country. Some common regulations are given in Table 7.3.

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Radiator

Air Receiver

Hydraulic Pump Drive

Operator’s Cab

Operator’s Cab Cab Air Cond., Heating and Pressurizing Unit

Leveling Jack Cylinder

Valve Bank

Hydraulic Tank Diesel Hydraulic Engine Pumps

Air Compressor

T Tank

Leveling Jack Cylinder

Air Cleaners

Hydraulic Oil Cooler

Diesel Engine Hydraulic Oil Tank Dust Collector Front Hydraulic Leveling Jack Engine Oil Filter Mast Raising Lube PLC Compressor Oil Cooler Tank Control Diverter Cooler Leveling Jack Cylinder Cylinders Engine Drill Deck Leveling Jack Cylinder Muffler

Air compressor

The Layouts in this Figure are not to the Same Scale

Driver’s Cab

Front Hydraulic Leveling Jack Fuel Tank

164 Rotary drilling and blasting in large surface mines

Figure 7.7 Layout of a carrier mounted (top) and crawler mounted ( bottom) rotary blasthole drill.

Carrier undercarriages used for rotary blasthole drills look like a common three or four axle truck. The normal heavy duty trucks, made by heavy vehicle manufacturers, are generally not suitable as a carrier to a blasthole drill. This is because the frame of these

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Figure 7.8 Layout of an all electric crawler mounted rotary blasthole drill.

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Bit Viewing Hatch – Hydraulically Operated

Hydraulic Breakout Wrench

Dry Dust Control

Leveling Jack – Hydraulically Operated

Service Platform

1

2

3

4

5

Item Description

Transformers Dry Type Compressor Oil Cooler

8 9

Heated and Insulated Cooler Compartment

Hydraulic Oil Cooler and Guard

7

10

Water Injection System for Dust Control

6

Item Description

Main Air Compressor Compressor Drive Motor Hydraulic Pumps and Gearcase Pipe Racks Operator’s Cab Hydraulic Oil Reservoir Electrohydraulic Valves

13 14 15 16 17 18 19

25

24

23

22

21

Automatic Lubrication System

Hoist /Pulldown Drive Control Cabinet

Rotary Drive Control Cabinet

Programmable Controller

Low Voltage Motor Control Center

Air Compressor Motor Control

Boarding Stairs

12

20

Cable Reel

11

Item Description

166


Table 7.3 Restrictions imposed on carrier undercarriages for being worthy of traveling on paved roads. Aspect

Restriction on carrier undercarriage

Road Support

Road support must essentially be rubber tired wheel type. Steel crawlers are not allowed as they damage the road surface. Maximum permitted axle weight is usually of the order of 20000 kg. For many reasons the front axles of the carrier are restricted to an axle loading of about 12000 kg. With these restrictions a rotary blasthole drill carrier is required to have minimum of three or four axles. Usually the maximum travel speed of the vehicle is desired to be in excess of 80 km/h. Lower speeds may be allowed but are undesirable from the viewpoint of travel time. Commonly height of the vehicle in excess of 4.5 m is not allowed The width of the vehicle should to be 2.5 m. Width of a maximum of 3 m is allowed as special case and if vehicles with 3 m width are to travel on roads special arrangements may be necessary. Maximum permitted length for a vehicle is of the order of 10 m. Longer vehicles may be permitted travel but only with special travel arrangements. All the regulations regarding front lights, turn signals, side marker lights and special marker lights The drivers cab has to be in such a position that he can see the road ahead in the best possible manner to drive the vehicle efficiently. In a rotary blasthole drill drivers cab is invariably separate because the operators cab is located at the rear end from where the road view is almost totally invisible due to several assemblies mounted on the deck. Drivers controls such as steering, brakes, accelerator, gear selector etc., along with all the indicators have to be in the drivers cab. Drilling controls are required to be in the operators cab located on the rear.

Axle Weights

Travel Speed

Height Width

Length Lighting Drivers Cab

Controls

trucks is not very rigid and too weak to withstand the stresses developed during drilling operations. Hence drill manufacturers themselves manufacture the carrier or get it manufactured as per their design. A few more points about carrier undercarriages are presented in a subsequent chapter on design criteria of blasthole drills. 7.5.1.2

Crawler undercarriage

A crawler undercarriage comprises two crawler tracks. A photograph of one such track is shown at the top of Figure 7.9. Each of these tracks contains many parts shown schematically in the lower drawing in Figure 7.9. These crawler tracks are joined together by means of a rigid cross frame in the case of a slewing undercarriage, or a frame having a pivot type and a fixed type cross beam to make the undercarriage pivot type. 7.5.1.2.1

Slewing crawler undercarriage

A typical slewing undercarriage is shown in Figure 7.10 where the rigid cross frame is clearly visible.

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Drive Sprocket

Track Pad Chain

Lower Load Bearing Rollers

Idler Roller

Track Beam

167

Idler Sprocket

Slack Adjuster Mechanism

Figure 7.9 Side frame of a crawler track.

Figure 7.10 A slewing crawler type undercarriage.

In the center of the cross frame is a large diameter circular roller bearing. The upper frame is mounted on this bearing. With such an arrangement the upper can be rotated in a 360° angle with respect to the lower crawler tracks. In certain cases such maneuverability enables the drilling of two blastholes without moving the drill. Besides this, a slewing crawler undercarriage offers good advantage in terms of placement of the drill on a narrow surface mine bench that has few rows and many blastholes in each of the rows.

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This can be explained by means of Figure 7.11. The left side illustration shows placement of the drill having a slewing undercarriage. The arrow shown in the figure is the direction of drill travel and also represents the movement of the center of gravity of the drill. This arrow lies at a distance D from the edge of the bench. It is also possible to make a drill with pivot type undercarriage to travel in this manner as shown by position B in the right side illustration in the figure. However, in this case the center of gravity of the drill, which is at a distance d from the bench edge, happens to be very near the edge. This greatly increases the danger to the drill on account of the slope stability. Hence, a drill with pivot type undercarriage has to be placed like the one shown by position C. The advantage of having a slewing crawler undercarriage, however, is not very great as it is only for drilling in the first row. For drilling blastholes in the second row, the drill with pivot type undercarriage can also be placed with crawler tracks as in position B, because in such case the distance D happens to be sufficiently great. Despite the above advantage, almost all the rotary blasthole drills being manufactured now are with pivot type undercarriages. These undercarriages have many advantages over the slewing undercarriages, as below.

Drill With Slewing Undercarriage

Drill With Pivot Beam Undercarriage

Elevation View

Elevation View

Position A

Position C

Plan View

Plan View

D

Blasthole Positions Drill Travel Direction

Position B

d Drill Travel Direction

Figure 7.11 Advantage in placement of a drill equipped with a slewing crawler undercarriage.

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1 2 3

4 5

169

A slewing undercarriage must be able to revolve within the space between hydraulic jacks. This imposes limitations on the crawler track size and blasthole drill stability. Design of the blasthole drill with slewing undercarriage is more difficult as the designer has to take into account different slewing positions while evaluating drill stability. A special hydraulic swivel has to be used between the junction of upper and lower to facilitate the flow of hydraulic oil to the propel motors in the undercarriage in all the slewing positions. Such swivel greatly obstructs the space available on the upper frame. Maintenance efforts required in the drill with slewing undercarriage are more than those required in case of pivot type undercarriage. In slewing undercarriages the crawler tracks are rigid and do not ensure uniform ground contact pressures as in the pivot type undercarriage.

In fact some manufacturers used to have slewing undercarriage for their rotary blasthole drills but later shifted to pivot type undercarriages. 7.5.1.2.2

Pivoted crawler undercarriage

In a pivot mounting, the main frame of a blasthole drill is supported on two crawlers through one fixed cross beam and one pivot type cross beam as shown in Figure 7.12. Such arrangement enables the crawler track to remain in contact with the ground even if the ground is not level because the crawler tracks are free to rotate around the fixed beam within the limits imposed by the pivot beam that is supported on the two crawlers through pins as shown in the figure. 7.5.1.2.3

Crawler track components

As the need for crawler-mounted equipment has grown, a few manufacturers have started making crawler tracks that can be used for a variety of equipment. They are available for supporting gross weights up to even 120000 kg. These ready-made crawler tracks are used for most blasthole drills. When, in case of extra large blasthole drills, the weight of the drill exceeds the load limits, manufacturers choose even sturdier tracks specially made for the purpose. Most of these manufacturers also make

Track Beam

Pin Support

Pivot Beam Pivot Pin

Rigid Beam Crawler Tracks

Figure 7.12 A pivot crawler type undercarriage.

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large crawler equipment e.g. shovels, and have the necessary manufacturing and heat treatment facilities for making suitable components. All the components in a crawler track are made from special alloy steels that are deep case hardened. An undercarriage may be slewing type or pivot type, but the crawler tracks comprise the same components as schematically shown in Figure 7.9. Features and functions of these components are as follows. TRACK BEAM

A track beam is usually made up of a welded box section. For the purpose of additional weight and rigidity the box can be filled with a special type of concrete that contains scrap steel pieces treated to avoid corrosion. A track beam supports the rigid frame from the side as shown in Figure 7.10 or fixed and pivot type cross beams from the underside as shown in Figure 7.12. IDLER AND LOAD ROLLERS

The upper side of the track beam supports a few idler rollers. These idler rollers in turn support the crawler pads from underneath. On the lower side of the track beam many load rollers are fixed. These rollers lift the beam on the crawler pads. They also ensure smooth and straight line movement of the crawler pads. SPROCKET

On one side of the track beam a sprocket is mounted. The sprocket is also called a drive tumbler. A sprocket has many teeth. These teeth interlock with the track pads. When the sprocket is rotated the track pads are forced to move. As these track pads are offered resistance from the ground, in effect the blasthole drill mounted on the undercarriage moves. In modern blasthole drills the rotary motion to the sprockets is imparted by means of a high torque hydraulic motor through a gear train like the one shown in Figure 7.13 or through an epicyclic gear box that is more compact. IDLER TUMBLER

On the other end of the track beam an idler tumbler is mounted. This idler tumbler also supports the track beam like the load rollers and sprocket. In addition, the tumbler is also equipped with a spring mechanism that keeps the tumbler under a pushing force so the crawler track does not lose its desired tightness. One type of spring arrangement is shown in Figure 7.14. TRACK PADS

A chain of track pads linked together forms a crawler track. When the crawler track is made from shovel type track pads as shown in Figure 7.12, the track pads are directly linked together by means of pins. The large bearing area between the pads and pins results in a very sturdy crawler track.

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171

Hydraulic Motor

Gear Train

Track Pad

Figure 7.13 Hydraulic motor and gear train for track drive.

Springs

Idler Tumbler

Figure 7.14 Spring arrangement for tightening the crawler track.

Track manufacturers usually make grouser type track pads made from cold rolled steel that is cut and heat treated. These track pads are bolted onto the track links that are joined together by means of springs to make a crawler track. Three types of grouser pads are available viz. single grouser, double grouser and triple grouser, as shown in Figure 7.15. Single grouser pads have excellent grip with the ground due to larger grouser height but lower bending strength. They are meant

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One Large Grouser

Holes for Bolting the Track Pad on Track Link

Two Medium Grousers

Three Small Grousers

Figure 7.15 Three types of track pads.

for dozing applications in softer ground. For rotary blasthole drills the high bending strength of the grousers is more important than the grip as the drills rarely work in soft grounds. Obviously, most of the rotary blasthole drills have triple grouser track pads. The number of pads in each of the track is so chosen that the area of contact of all the track pads on both the crawler tracks exerts pressure less than 50 to 100 kPa on the ground. For reduced maintenance, most of the ball or roller bearings used in crawler tracks are lifetime lubricated and sealed. Travel speed of most of the crawler blasthole drills ranges within 1 to 4 km/h. Higher travel speeds enable quicker placement of blasthole drills as well as bench-to bench-travel movement. However, speeds more than about 3 km/h do not offer much advantage as the travel distances are limited to about 6 to 15 m on average. Making the drill travel at high speed is very dangerous from the viewpoint of the stability of the drill. Larger blasthole drills with tall masts have lower travel speeds. Maximum grade ability of a crawler-mounted blasthole drill is about 30% with lowered mast.

7.5.2 Assemblies in the upper group All the assemblies in the upper group are mounted on a frame called the main frame. Even the mast that contains all the assemblies in the mast group, is also mounted on the main frame.

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173

In the case of a diesel engine-driven blasthole drill, the important assemblies in the upper group are: 1 2 3 4 5 6 7 8 9 10 11 12

Main Frame Leveling Jacks Prime Mover Compressor Hydraulic System Automatic Lubrication System Radiators and Oil Coolers Machinery House Dust Control System Operator’s Cab Driver’s Cab Fuel Tanks

If the blasthole drill is electrically operated the diesel engine is replaced by an electric AC motor that is suited to run directly on the high voltage electric supply in the mine. For all electric drills, some functions are better performed by AC electric supply and others need DC electric supply. Low power motors need lower supply voltage. Thus, a transformer and a converter are needed. These are also mounted on the drill. All electrically operated blasthole drills usually have transformers to reduce the high voltage of mains to an appropriate voltage for various motors, and also a converter to convert some electricity from AC to DC. They also have a provision for a cable reel mounted on front side of the drill. Details of the above assemblies/components are given hereunder individually. 7.5.2.1

Main frame

During blasthole drilling operations the weight of all the components in the three groups is borne by the main frame and transmitted to the ground through hydraulic jacks. The main frame is also subjected to heavy vibrations generated in drilling. Therefore, the frame not only has to be strong but very rigid as deflections in the frame give rise to high bit wear and blasthole deviations that are undesirable for drilling and subsequent blasting. A typical main frame is made by using two extra thick I/C sections welded together by means of cross beams. For many large drills I/C sections of appropriate size are not readily available. They have to be made by welding three pieces thereof. To ensure adequate support to the many assemblies that are to be placed outside the frame, many cantilever beams are also welded to the frame. Vertical members that support the mast, hydraulic jacks, machinery house, operators cabin etc. are also directly welded to the main frame. In most cases the steel chosen for manufacture of the main frame is ASTM A36 or equivalent. Often main frame is provided with catwalks, also called walkways, that are required for easy access to the assemblies/components mounted on the main frame. Anti-skid plates or thick wire gages are used for the floor of catwalks and mainframe.

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174


The main frame in turn is fixed onto the crawler undercarriage and looks as shown in Figure 7.16. 7.5.2.2

Leveling jacks

Rotary blasthole drills exert very heavy feed force on the drill string. A crawler track is not ideally suited to resisting the heavy reactions generated in drilling. Further, a blasthole drill resting on uneven ground through the crawler track rarely gives a firm horizontal position to the main frame. As all the mast alignments are designed for a perfectly horizontal position of the main frame, it is essential to ensure that while drilling, the main frame is in a firm horizontal position. Leveling jacks facilitate horizontal positioning of the main frame. A blasthole drill is equipped with two jacks on the rear side i.e. the mast side and either one or two jacks on the front side. In electrically-operated drills two jacks on the front side are often essential to accommodate the cable reel. Leveling jacks are operated hydraulically. For guiding the operator in accurate leveling, two tube type, and one circular, bubble levels are provided in the operators cab. The operator controls the flow of hydraulic oil to the jacks and extends/retracts them as required to level the drill. Hydraulic valves used for the hydraulic jacks are an automatically locking type so even if the tube supplying hydraulic oil to the jack leaks, the hydraulic oil confined within the cylinder of the jack does not flow out and the jack remains in the same extended position. Usually the length of hydraulic jacks is so designed that they enable leveling the drill even if the ground is sloping at an angle of 10%. To prevent the penetration of the jack rod into the ground a thick circular steel pad is provided at the bottom of the jack pad. The joint between jack pad and jack rod is knuckle-type so the pads can easily align with the ground surface without inducing stresses on the pad or rod. The diameter of the jack pads is so designed that the contact pressure between jack pad and ground is between 200 to 350 kPa.

Operator’s Cab Mast Support

Transverse Beam

Crawler Undercarriage Cross Beam Support for Leveling Jack Figure 7.16 Main frame of a blasthole drill mounted on crawler undercarriage.

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7.5.2.3

175

Prime mover

Depending upon its type and size, the power requirement of a rotary blasthole drill ranges between 200 kW to 1300 kW. Diesel electric drills are rarely made nowadays. For diesel-hydraulic, electrichydraulic and electric-electric crawler drills, typical schematic power source and distribution diagrams are shown in Figures 7.17, 7.18 and 7.19.

Air Compressor

Batteries

Diesel Engine

Starter

Alternator

Controls

Miscellaneous Operations Fans, Lighting and Air Conditioning in Cab and Machinery House.

Hyd. Pump 1

Hyd. Pump 2

Drill String Rotation, Propel Left Track, Leveling Jacks

Drill String Feed Propel Right Track

Hyd. Pump 3

Work Lighting Auxiliary Operations Mast Raise/Lower, Mast Brace, Carousel Rotation, Hyd. Spanner in/out, Breakout Wrench in/out, Breakout Wrench Lock, Breakout Wrench Rotation, Hydraulic Oil Cooling, Water Injection Pump, Oil Injection Pump, Auxiliary Winch, Radiator Cooling Fan, Other Auxiliary Operations

Figure 7.17 Power source and distribution schematic for diesel – hydraulic blasthole drill. High Voltage Supply SD Transformer Batteries

Electric Motor

Starter

Controls

Miscellaneous Operations Fans, Lighting and Air Conditioning in Cab and Machinery House.

Air Compressor

Hyd. Pump 1

Hyd. Pump 2

Drill String Rotation, Propel Left Track, Leveling Jacks

Drill String Feed Propel Right Track

Hyd. Pump 3

Work Lighting Auxiliary Operations Mast Raise/Lower, Mast Brace, Carousel Rotation, Hyd. Spanner in/out, Breakout Wrench in/out, Breakout Wrench Lock, Breakout Wrench Rotation, Hydraulic Oil Cooling, Water Injection Pump, Oil Injection Pump, Auxiliary Winch, Radiator Cooling Fan, Other Auxiliary Operations, Cable reel Winding

Figure 7.18 Power source and distribution schematic for electric – hydraulic blasthole drill.

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High Voltage Supply SD Transformer

Starter

AC/DC Converter

High Voltage Motor (AC) Rotation Motor (DC)

Hoist and Pulldown Motor (DC) Propel Left Track

Low Voltage Motors (AC)

Air Compressor

Hyd. Pump 1 Propel Right Track

Hyd. Pump 2 Leveling Jacks

Water Injection Pump, Oil Injection Pump, Radiator Cooling Fan, Cab and Machinery House Pressurization, Cable Reel Winding Converter

Batteries

Controls

Fans, Lighting and Air Conditioning in Cab and Machinery House, Work Lighting Auxiliary Operations Mast Raise/Lower, Mast Brace, Pipe Rack Movements, Hyd. Spanner in/out, Breakout Wrench in/out, Breakout Wrench Lock, Breakout Wrench Rotation, Auxiliary Winch, Other Auxiliary Operations

Figure 7.19 Power source and distribution schematic for electric – electric blasthole drill.

How much power is required for propelling and the different drilling operations is a topic discussed in a later chapter on design criteria for blasthole drills. In carrier-mounted rotary blasthole drills, the diesel engine used for driving the carrier is used for actuating all the devices used in drilling. Power from the carrier engine is taken through a power take off, also called the transfer case. A typical power train diagram for a carrier-mounted rotary blasthole drill is shown in Figure 7.20. In many such drills the power available from the carrier engine is sufficient for actuating all other assemblies except the compressor which is driven by a separate engine. 7.5.2.4

Compressor

Compressors currently used on rotary blasthole drills are oil-flooded rotary screw type or in a few cases rotary vane type. Whichever the type may be, a compressor consumes more power than any other assembly or component on a rotary blasthole drill except perhaps the rotary head. A compressor is directly driven by the diesel engine or a high voltage high power electric motor. As the two components are very heavy they are transversely mounted at the center of the main frame for stability. Other components associated with an oil-cooled compressor are air oil separator and oil cooler. They are separately mounted on the main frame. More details in respect of the working of the compressors are given in chapter 8 in this book.

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Mast Rotary Head Pipe Changer Rotary Head Drive Assembly Twin Shaft Transmission Clutch Assembly Triple Gear Pump Air Compressor Transfer Case i.e. Power Take Off Main Transmission Driver’s Cab

Operator’s Canopy Rear Steering Wheels

Front Steering Wheels

Figure 7.20 Power train in a carrier mounted blasthole drill.

7.5.2.5

Hydraulic system

Every blasthole drill has a hydraulic system. Major items in the hydraulic system are hydraulic reservoirs, filters, pumps, tubing, valves, motors, cylinders etc. Pumps and motors used in the hydraulic system are often variable displacement type. Hydraulic pumps are directly driven from the other end of the diesel engine or the main electric motor. In many cases special pump drives, like those shown in Figure 7.21 are used rather than double or triple pumps mounted on the same central shaft, so they can be serviced separately. Wherever possible metal tubing is used as it give less pressure loss and seldom requires replacement. Most of the valves are electro-hydraulic type. This eliminates the need for linkages that are costly and difficult to repair or maintain. A typical circuit diagram of hydraulic system of a electric-electric blasthole drill is given in Figure 7.22. Blasthole drills with hydraulic power distribution requires many more components. 7.5.2.6

Bit lubricating system

Tricone bits used for rotary blasthole drilling are of two types viz. air cooled or sealed bearing. In air cooled bits, even if the bearings are cooled by compressed air flowing through the bits, the requirement for lubrication of the bearings remains unfulfilled. For this reason the provision of a bit lubrication system on a blasthole drill becomes essential. A typical bit lubrication system consists of a tank and a timer-controlled adjustable discharge pump for injecting lubricant oil into the compressed air piping through a nozzle.

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For Mounting Four Hydraulic Motors

For Mounting Two Hydraulic Motors

Clutch Assembly

Figure 7.21 Hydraulic pump drives.

Figure 7.22 Hydraulic circuit diagram for a typical blasthole drill.

When the pump develops very high pressure in the delivery line, the nozzle discharges the oil in the air stream in the form of very fine particles. This oil mist travels along with the compressed air to the tricone bit bearings. Many parameters such as hole diameter, compressor discharge, type and viscosity of lubricant etc. are required to be considered while designing a good bit lubrication system. 7.5.2.7

Automatic lubrication system

Despite use of lifetime-lubricated sealed bearings in most of the locations, many components of rotary blasthole drills, such as the gear assembly in the rotary head, slack adjusting assembly, sprockets in the crawler tracks etc., need to be externally lubricated.

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It is a very tedious job to reach near the lubrication points and lubricate all these parts by using a grease gun. Many points, particularly those on the mast and components therein are accessible only with great difficulty. Further, during such manual lubrication the drilling operation is required to be stopped for the purpose of safety. To overcome this difficulty many blasthole drills have an automatic lubricating system as shown schematically in Figure 7.23. The following is the general idea of the working of such a system. Primarily the system comprises a central pump and control station located on the main frame of the drill. This station has a lubricating oil reservoir from where the oil is fed to a lube line by a feed pump. The control unit ensures that the pressure of lubricant in the line is kept at a certain predetermined level so the lubricating oil remains without any voids or bubbles. The presence of bubbles in the oil creates an air lock and

Motorized Pump High Pressure (3.5 MPa) High Volume (4 L/hr) 24 VDC Battery Operated Inbuilt Filter, Pressure Gage, Relief Valve Excess Pressure Switch Electrical Vent Valve Injector Adjustable Output Visual or Electrical Indicator Independent and Serviceable

Progressive Distributor Output Adjustable in Multiples Contamination insensitivity Special Relief Valve to Identify Blocked Line Choice of Output Connection Electric Cable Lubricant Piping

Figure 7.23 Central pump and control station of the lubricating system.

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stops the flow of oil in the line. A timer in the station operates the pump and increases the pressure in the manifold to certain higher levels only at predetermined time intervals. The tubing attached to the manifold carries oil up to the point where lubrication is required to the parts. Near every point of lubrication, a spring loaded injector is placed. This injector opens its nozzle only when the pressure in the manifold rises to preset higher levels. Lubricant released by the injector in this manner is taken up to the point of lubrication by means of a small tube and released. The pressure setting of the injectors can be adjusted as shown in Figure 7.24 so that the right quantity of lubricant is fed to the point from the manifold by means of small tubes as shown in Figure 7.25. A system monitor lights located in the operator’s cab indicates the status of the system by sensing the pressure in the manifold. If sufficient pressure is not being developed, the operator is warned through an audible alarm as well as a blinking red light signal. In some blasthole drills two or more pumping and control stations are provided, so the time intervals between lubrications can be kept flexible. The number of points which an automatic lubrication system lubricates automatically is dependent upon the design of the rotary blasthole drill. In many large blasthole drills the number of points is in excess of 100 as listed in Table 7.4.

Lubricant Tube from Injector

Lubricant Tube to Injector

Injectors

Figure 7.24 Lubricant injectors.

Figure 7.25 Lubricant tubes on a track beam.

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Table 7.4 Points of lubrication on a rotary blasthole drill. Assembly of the drill

Lub. Points

Details

Crawler Undercarriage

35

Drive tumbler, Fixed axle bushings, Lower rollers, Slack adjuster, Equalizer beam, Propel chains

Hydraulic Jacks

16

Upper and lower guides

Propel Machinery

4

Mast and Platform

33

Tool wrench cylinder guides and bearing blocks, Sprocket pins, Wire ropes, Pulldown chains, Hoist cylinder pins, Curtain rope pulleys, Angle drilling pipe positioner, Idler sprockets, Shipper shaft bearings, Rack and pinions, Mast guides

Rotary Head

11

Main Machinery

20

Air seal, Adjusting collar, Gear case, Seal, Rotary coupling, Pipe joint Propel gear case, Propel shaft, Propel sprockets, Propel brakes, Hoist/Pulldown gear case, Hoist sprocket, Planetary unit, Sprockets, Auxiliary winch drum, Auxiliary brake lever shafts

7.5.2.8

Intermediate propel shaft inner and outer bearings

Radiators and oil coolers

Almost all the diesel engines used on blasthole drills are the water-cooled type. Water circulated for engine cooling needs a large radiator to ensure adequate cooling of the engine. Besides this, the compressor oil also needs a cooler. To ensure that the radiators and coolers transfer heat to atmospheric air, they are placed on the edge of the main frame. 7.5.2.9

Machinery house

The importance of having a machinery house on a rotary blasthole drill for protection of the assemblies and components on the drill, is emphasized in chapter 10 which deals with operation of the drill at high altitudes and in severe weather conditions. The interior of a machinery house looks like the cut view shown in Figure 7.26. The roof the machinery house is either easily removable or can be tilted to near vertical position to allow lifting of the heavy assemblies or components by a crane. The walls, doors and even rooftop of the machinery house can be made up of a double wall where two sheets – an inner and outer – are provided. The air between the two sheets acts as a thermal insulator. In many cases the machinery house is pressurized so the outside hot or cool air does not enter the machinery house. The machinery house can also be equipped with air conditioner or heater or both. 7.5.2.10

Dust control system

Rotary blasthole drills flush the hole by compressed air. Along with large chips, very tiny microscopic particles are also formed during the process of rock crushing under a drill bit. The chip- and dust-laden air emerges out of the blasthole during blasthole flushing. If such dust-laden air is allowed to mix with the atmosphere, the atmosphere gets polluted. Inhaling such polluted air gives discomfort to everyone. The harmfulness

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1 - Machinery House 2 - Rotary Motor 3 - Air Compressor 4 - Full Height Door 5 - Electric Console 6 - Machinery House Filter 7 - Mast 8 - Rotary Head 9 - Operator's Cab

Figure 7.26 Cut view of machinery house and operator’s cab.

of polluted atmospheric air depends upon the composition and concentration of particles mixed with the air. Medical research has proved that crystalline varieties of silicon dioxide or silicates in the form of asbestos are dangerous and cause silicosis or lung cancer respectively. Both these diseases are considered to be incurable, but some oxygen therapy practitioners have claimed that oxygen therapies can reverse them. Other ingredients such as fines of feldspar, olivine, amorphous varieties of quartz glass etc. are relatively harmless. To minimize pollution, dust control arrangements are essential on blasthole drills, particularly the rotary type as they create a large quantity of fine particles. Dust control arrangements are of two types viz. wet dust control and dry dust control. 7.5.2.10.1 Wet dust control This is also called water or foam injection. This simple method of dust control is quite effective as can be seen from Figure 7.27. The wet dust collection mechanism consists of a water pump and a water tank mounted on the main frame of the blasthole drill. The water or foam from the tank is sprayed into the compressed air line by the water pump. A mist of water or foam formed in the compressed air line travels to the bottom of the hole and escapes out into the annular space through bit nozzles and rises up. The fine cuttings in the compressed

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Without Water Injection

With Water Injection

Figure 7.27 Effectiveness of wet dust control.

air get coagulated with the moisture or foam. When such aggregates of fine particles come to the top of the hole, being heavier they fall to the ground within the dust hood enclosure. Thus, clean air escapes to the atmosphere. Mixing foam with the water proves very effective because the higher surface tension of the foam mixture coagulates the particles with higher binding forces. Bailing velocity decreases by use of foam but can achieve effective blasthole flushing. The quantity of water to be injected by a water injection system should be such that it forms a mist and remains in gaseous form in the compressed air. The quantity of water to be injected can be determined on the basis of Table 7.5. If foam is to be used the quantity of water must be reduced depending upon the properties of the foaming agent. In wet dust collection the dust hood must have adequate openings within it so the clean air escapes to atmosphere without building up pressure within the dust hood itself. The advantage of water or foam injection is that it is inexpensive to construct, maintain and operate. The disadvantages of water injection system are: 1

2

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Since the compressed air is mixed and the mist travels to the bearings of the drill bit, the life of bit bearings and, therefore the bit, decreases considerably – in many cases as much as 15% in terms of meterage. In conditions where atmospheric temperatures are sub-zero, water in the water tank and particularly the water injection pump and tubes can easily freeze and crack the system. Use of antifreeze liquids can be harmful to the bit and stabilizer bearings and other components of the drill string.

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Table 7.5 Quantity of water to be injected for dust suppurations. Atmospheric temperature in °C

Quantity of water to be injected in cc per m3 of compressed air

45 40 35 30 25 20 15 10 5

40–70 30–60 25–40 18–30 14–22 10–17 8–13 6–9 4–7

Figure 7.28 A dust hood at mast bottom.

Some countries do not accept water injection dust control since the dust deposited on the surface of the ground becomes dry in very short time and gets easily mixed up with the atmospheric air thereafter. 7.5.2.10.2

Dry dust control

In dry dust control systems, when compressed air laden with cuttings, dust and dry clay moves up into the dust hood, most of the sand and larger size particles settle down within the dust hood itself due to decrease in air velocity. But still a large quantity of dust and clay-laden air remains in the dust hood. Dust hoods provided on dry dust collection systems have to be airtight. Further they must keep in contact with the ground even when the mast is tilted for angle hole drilling. One such dust hood is shown in Figure 7.28. The alignment, size and shapes of dust hoods are often controlled by mechanical or hydraulic means. The dust and clay-laden air in the dust hood is further cleaned by a dust collector. Dry dust collector units provided on rotary blasthole drills are usually the filter type. These units suck the dust and clay-laden air from the dust hood and bring it into the filtration unit. Dust collector operation can be explained by means of Figure 7.29.

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Blower Motor

Vibrators Reverse Blower

Blower Fan Reverse Blower

Dust Hood Middle Chamber Filter Elements Hopper

Dust Collection Bag

Figure 7.29 Dust collector schematics.

Drill Pipe

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The blower fan at the top of the dust collector creates a vacuum in upper chamber. This results in a flow of dust-laden air through the conduits to the middle and upper chamber of the dust collector. As the dust-laden air passes into the upper chamber from the middle, it gets filtered by the filter elements in the middle chamber. Thus, the air that passes to the upper chamber and released to the atmosphere is very clean. Dust collected in the middle chamber side of the filter element is made to fall down in the dust bag by a vibrator that vibrates the filter elements at predefined time intervals. A reverse blower also blows compressed air on the filter elements in the reverse direction at some predefined time intervals. This cleans the filter elements by forcing the dust to fall into the dust bag. When the dust bag becomes full of dust, it is replaced with another empty bag. The best way to dispose of the dust in the dust bag is to mix it with some cement and make a concrete block. Some sophisticated dust collectors have electronic circuits that automatically sense the air pressure differential between the upper chamber and middle chamber and operate the vibrator or reverse blower if the differential exceeds a predefined value. It may be a good idea to have another filtering unit to continue the filtering operation even when the dust in the first unit is being cleaned. The filter element fabric is such that it arrests almost all the particles of size larger than 5 microns. Particles of size smaller than this are considered harmless and, therefore, can be ejected with air blowing through the blower fan into atmosphere. The air blowing capacity of the blower fan has to be far in excess of the capacity of the air compressor on the blasthole drill. Only this can ensure that no pressure is built up in the dust hood and that the blasthole flushing remains very effective. Dry dust collectors are more expensive but are more effective in pollution control as compared to wet dust control. They require more maintenance in changing the filter fabric elements. 7.5.2.11

Operator’s cab

In a blasthole drill the operator’s cab is located on the rear end, from where the mast, drill pipe, rotary deck bushing, drill stem wrench, pipe changer and many other items that take part in drilling operations are easily visible. For this purpose one side of the cab is equipped with large glass windows right from the floor to the roof. For an all-round view the other walls of the cab have windows in the upper half of the wall. In most cases the roof also has a glass window so the operator can see the rotary head and upper portions of the mast. The glass used is shatter resistant type, tinted glass. Roof window glass is extra thick and is provided with a steel guard. A typical view of the mast through the roof window is shown in Figure 7.30. The cab is constructed from thick steel sheets and is of FOPS (Falling Object Preventive Structure) type. For noise and vibration reduction, the cab is insulated from the main frame by synthetic rubber mounts. Facilities provided in an operator’s cab include operator’s adjustable seat, helper’s seat, windshield wipers and washers, cab heater and air conditioner, cab pressurization unit etc. Nowadays, in most of the drills electro-hydraulic valves are used. Thus, the controls are electrically operated. Apart from on-off and rotary switches, joystick controls

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Figure 7.30 Mast view through roof window.

Figure 7.31 Operator’s console of a blasthole drill.

are also used. In computer-controlled drills a large color LCD display in front of the operator shows all the important drilling parameters as shown in Figure 7.31. The console in front of the operator contains many indicators and controls. Indicators of two types viz. gauges and lights are provided in the operator’s cab of the drill. A rotary blasthole drill usually has some or all of the indicators mentioned in Table 7.6. Some drills also have some additional indicators depending upon the assemblies incorporated in it. Controls are also of two types viz. levers or switch buttons. Controls normally provided in a rotary blasthole drill are mentioned in Table 7.7. Some drills also have some additional controls.

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Table 7.6 Indicators usually provided on rotary blasthole drill. Indicator list

Indicator list

Drill Level Indicator Rotary Speed Gauge Water Injection Tank Level Gauge Machinery House Heater or Air conditioner On Light Engine Water Temperature Gauge Voltmeter or Ammeter Engine Hour Meter Fuel Level Gauge

Compressor Temperature Gauge Pulldown Force Gauge Rotary Torque Gauge Water Injection System On Light Cab Heater or Air conditioner On Light Engine Oil Pressure Gauge Engine Speed Gauge

Table 7.7 Controls usually provided on rotary blasthole drill. Control list

Control list

Engine Start Button Ether Injection Control Button Leveling Jack Levers Pulldown Force Control Lever Breakout Wrench Control Levers Pipe Rack Control Lever Mast Locking Lever Water Injection Start Button Dust Curtain Lifting Control Levers

Tool Handling Winch Lever Engine Stop Button Compressed Air Shut Off Lever Mast Hoisting Lever Propel Control Levers Rotary Speed Control Lever Pipe Rack Indexing Control Pipe Rack Locking Lever

7.5.2.12

Driver’s cab

When the blasthole drill is carrier-mounted, it is meant to travel on highways. In such case, the drill itself and the driver’s cab must have all the facilities to conform to road regulations. Every manufacturer takes care of providing all the required facilities to comply with such regulations. Commonly essential facilities are: 1

2

3

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Travel lights consisting of side marker lights, red light on the front top and rear top of the horizontal mast, revolving dome light, two position headlights, direction signal lights, parking lights etc. Facilities required in driver’s are driver’s and rider’s seats, seat belts, rear view mirrors, light with switch, steering wheel, accelerator, gear shift lever, clutch, air brakes, parking brakes etc. The front panel requires air pressure gage, voltmeter, fuel level gage, engine oil pressure gage, engine coolant temperature gage, low air pressure indicator, speedometer, tachometer with totalizer, audiovisual warning, high beam indicator light, parking brakes on indicator light, key type starting, engine start button, engine stop button, hand throttle control, transfer case lock-up control with indicator light etc.

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Optional items available in the driver’s cab may include air conditioner, heater, AM/FM radio etc. 7.5.2.13

Fuel tanks

Every diesel-operated rotary blasthole drill is equipped with a fuel tank, usually located below the upper floor on the main deck. In most of the cases the capacity of the tank is such that fuel in the completely filled tank is sufficient for continuous engine operation of 24 hr. An option to enable filling the fuel in the tank from a fuel tanker very rapidly is usually available. A few drills have the option of having an additional fuel tank so as to reduce the frequency of fuel fill. 7.5.2.14

Transformer and cable reel

All electric drills have electric motors working on different voltages and different types of current i.e. AC and DC. The main electric supply available in a mine is AC type and has voltage range between 4160 to 6600 V. The DC motors used in the drill typically work on 475 V, whereas the other AC motors work on 440 to 460 V. Low voltage AC power used for air conditioning and lighting is typically 120 or 240 V. Therefore all electric drills are equipped with one or two transformers. Oil-cooled transformers of the past have now been replaced by dry transformers. A cable reel, mounted on the front end of a blasthole drill, as shown in Figure 7.32, is usually an optional item, but is very essential in electric drills. A cable reel through an electric or hydraulic motor automatically ensures tidily and tightly wound position of the power cable that supplies power from the mine power supply to the blasthole drill even as the drill moves from one blasthole to another.

7.5.3 Assemblies in the mast group Assemblies in the mast group are usually mounted on the mast itself and include: 1 Mast 2 Mast Raising Cylinders and Mast Braces

Figure 7.32 Cable reel.

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3 4 5 6 7 8 9 10


Mast Ladder Pipe Changer Angle Hole Attachments Rotary Head Feed Mechanism Auxiliary Winch Tool Handling Jib Centralizer Details of these are given hereunder individually.

7.5.3.1

Mast

The mast of a rotary blasthole drill is an item of predominant importance. It is subjected to many reactions in drilling which result in complex stresses in the mast members. When modern concepts of design of rotary blasthole drills were first introduced, the most radical change was the mast and moving rotary head accommodated in the mast. The mast in itself is a column structure fabricated from square or rectangular steel tubing. Unlike the circular tubing that requires contour cutting, square or rectangular tubing needs plane cuts and the welds can be easier as well as more reliable. The cross section of the mast is rectangular in most cases but some drills have trapezoidal or triangular masts. Three or four main members of the mast are joined together by means of lattice members to form a open-faced hollow structure as shown in Figure 7.33. The rotary head is visible and accessible from the open face and slides in an up and down direction within the members. Main Members Rotary Head Containing Gear Reduction Mechanism

Lattice Members

Hydraulic or Electric Motor Position of single pipe racks

Rotary Swivel and Spindle

Hydraulic or Electric Motor

Main Members

Figure 7.33 Cross section through a mast.

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7.5.3.2

191

Mast raising cylinders and mast braces

The mast is raised and lowered by means of two hydraulic cylinders. For long distance travel such as from one bench to the other, the mast of the blasthole drill is kept in the lowered position so that the center of gravity of the drill attains its lowermost position and travel is safer. While moving from one blasthole to the other, the mast is kept in the raised position to save time. Some blasthole drills have a very tall mast in order to accommodate long drill pipes to facilitate drilling deep holes in a single pass. In such structures the stresses and vibrations induced in the mast by the drill head are amplified. Thus, a tall mast needs additional support that is provided by two mast braces. Mast braces are often hydraulically extended and mechanically locked so the longer fixed length gives necessary rigidity to the mast. This reduces the unsupported mast length and results in low vibrations which otherwise would have been higher by relying only on the hydraulic mast raising cylinders. For alignment of the mast in prefixed angles, mast braces have additional fixing points. Mast-raising cylinders and mast braces can be easily seen in Figure 4.12. 7.5.3.3

Mast ladder

Occasionally it becomes necessary to climb up to the top of the mast. For this purpose a mast is equipped with a ladder. For safety purposes the ladder is surrounded by circular rings or a steel grill. With this grill climbing becomes very easy and safe. 7.5.3.4

Pipe changer

A pipe changer is also known by many names viz. pipe rack, rod changer or carousel. In blasthole drilling, the addition of a drill pipe to the drill string and its removal after drilling a blasthole to the desired depth is avoided by selecting a drill that can drill to full depth with one drill pipe. However, for drilling deep holes, extra drill pipes are to be added. For this reason, almost every blasthole drill has an arrangement for storing additional drill pipes in the mast and mechanically joining them to the drill string. This mechanism is called a pipe changer. Pipe changers differ from drill to drill but mostly they are either carousel type and single pipe type. 7.5.3.4.1

Carousel pipe changers

Small, medium, and some of the large, blasthole drills are equipped with a carousel pipe changer. It is often fixed to the main member of the mast on the opposite side of the operator’s cab or within the mast behind the drill head like the one shown in Figure 7.34. The number of drill pipes stored in a carousel pipe changer depends upon the pipe diameters but in most cases it is between 3 to 6. In a carousel pipe changer the drill pipes are stored in a rack that consists of a central spindle with one bottom and two top plates. As shown in Figure 7.35 the bottom plate has sockets on the upper side to accommodate the lower ends of the drill pipe. The upper ends of the drill pipe are contained within a slot in the lower of the two top

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Figure 7.34 Pipe changer within a mast.

Upper of the Top Plates

Drill Pipe Free to Move Out from the Carousel

Lower of the Top Plates Drill Pipe Locked in the Pipe Rack Slot

Socket Bottom Plate

Figure 7.35 Pipe locking in the pipe changer.

plates. When the pipe is in lowered position it is restricted in its upper movement by the upper of the top plate. It cannot come out of the slot because the slot opening is of smaller size than the drill pipe diameter. Thus, the drill pipe in its lower position is locked. The upper of the top plate has an opening from the top. The drill pipe can be

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lifted through this opening by the drill head. In the lifted position the drill pipe comes out of the socket on the bottom plate and the small diameter groove on the drill pipe aligns to the level of the slot opening as shown in Figure 7.35. Thus, in this position the drill pipe is unlocked and is free to come out of the carousel. The steps taken in adding a drill pipe to the drill string are as follows: 1 When drilling with one drill pipe is finished and a drill pipe needs to be added, a spanner at the bottom of the mast is forced out and holds the upper end of the drill pipe through its flat. This is shown in Figure 7.36. 2 The rotary head rotates in the reverse direction so the drill pipe is uncoupled from it. In most drills such uncoupling is started by a hydraulically operated tool wrench that grips the saver sub at the lower end of the drill head and slightly rotates it. After this loosening the drill head uncouples the saver sub. 3 The rotary head moves to the top position in the mast along with the saver sub. 4 The complete carousel assembly is moved from its normal position to the center of the mast by means of a hydraulic cylinder in such a way that the opening of the upper of the top plates is just below the saver sub. 5 An indexing mechanism, also operated hydraulically, rotates the lower of the top plates and the bottom plate in such a way that one drill pipe is in alignment with the saver sub and the drill pipe in the blasthole is held at the mast bottom by the spanner.

Drill Head

Spanner Saver Sub

Drill Pipe

Figure 7.36 Bottom spanner.

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6 The rotary head moves down and rotates the saver sub in the right direction so that the drill pipe in the carousel is coupled to the saver sub. 7 The rotary head is then moved up slightly so the drill pipe in the carousel moves up and attains a position in which it can come out of the socket at the lower end and the narrow groove of the drill pipe comes in the level of the slot opening. 8 When the pipe is in such position the carousel is moved back to the original position on the left of the mast. Thus, the newly added drill pipe remains in the drill string alignment and the pipe changer is away from it. In case of inclined blastholes it may be necessary to hold the newly added drill pipe in proper alignment by yet another arm. 9. The rotary head with newly coupled drill pipe is then moved down. When the drill head rotates in the right direction the newly added drill pipe gets coupled with the drill pipe in the blasthole held in position by the spanner. 10. The spanner is then retracted back. This once again allows free rotation of the drill string and drilling is continued further. The above steps are repeated for adding one more drill pipe whenever required. This mechanism appears quite simple but to have it working precisely, it has to be designed very carefully because even small deflections can cause problems. Such problems increase considerably in coupling drill pipes in an inclined direction while drilling angle holes. The advantage of the carousel type pipe changer is that it can allow the use of many drill pipes. 7.5.3.4.2

Single pipe changers

Single pipe changers are used in some large and most of the extra large blasthole drills. This is because single pipe changers are sturdier without being very heavy, and are better suited to handle very long drill pipes of large diameter. In modern drills there can be as many as four single pipe changers as shown in Figure 7.37. Single pipe changers are totally independent of each other and are operated separately. After completing drilling with the first drill pipe, it is held in position by a spanner as explained earlier for a carousel changer. The rotary head detaches from the pipe and moves up in the same manner as explained in step 1 and 2 for a carousel pipe changer. The drill pipe in a single pipe changer is held in locked position by means of a socket shown in the in Figure 7.37 at the bottom and a gate at the top. Besides this, other means of locking the drill pipe are also used for the purpose of safety. By extending the hydraulic cylinder at the bottom of the pipe rack, as shown Figure 7.38, the drill pipe gets positioned in line with the drill pipe held by the spanner at the bottom of the mast. The gate is then opened to allow coupling of the drill pipe to the rotary head as described in step 5 and 6. Once the drill pipe is coupled to the rotary head, a centralizer swings into position and holds the drill pipe at some distance from the lower end of the mast. The centralizer is particularly essential while coupling drill pipes in angle hole drilling. At this juncture, since in this position the drill pipe is safely and securely held, other locking arrangements are released by retracting the hydraulic cylinder. Thus the

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Hydraulic Wrench

195

Four Single Pipe Changers

Figure 7.37 Four single pipe changers and hydraulic wrench.

Upper Arm

Gate

Latch

Drill Pipe

Main Section of the Mast

Hydraulic Cylinder for Lowering the Drill Pipe

Lower Arm

Figure 7.38 Operation of single pipe changer.

single pipe changer moves back to its original position. As the pipe is free to slide and rotate in the centralizer, it is rotated by the drill head and coupled to the drill pipe held by the jaw at the bottom of the mast. Drilling can be continued after withdrawing the jaw and releasing the pipe from the centralizer.

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7.5.3.5

Bit changer

Pipe changers have been used on rotary blasthole drills for the past sixty years. In fact it was one of the first automation features of the blasthole drills. Recently Bucyrus International have come up with a new bit carousel and forked pipe wrench, as shown in Figure 7.39 on their 49 series blasthole drill. The bit changer accommodates up to four tricone bits. In the bit changing operation the forked pipe wrench engages the drill pipe flats just above the bit. Then the wrench rotates while simultaneously transmitting impacts on the pipe. In the process the tool joint loosens. The drill bit is then maneuvered into the bit carousel and by reversing the sequence a new bit is attached to the drill pipe. All these operations are automatic and are carried out by the driller while sitting in the cab. Even the assistant need not go to the drilling platform. All this means significant safety measures are introduced into the bit changing operations. More importantly the time required for changing the drill bit reduces to about 10 minutes from about 60 minutes that are required for the same operation without the tricone bit carousel. 7.5.3.6

Angle hole attachments

For reasons elaborated elsewhere in this book, blastholes are often drilled at an angle with the vertical. In most instances this angle is within limits of 0 to 30° off vertical. When the mast of a blasthole drill is raised from the horizontal to the vertical position by using mast raising cylinders it does attain such angles, but mast raising cylinders are meant for raising the mast and not for firmly holding it at an angle with the vertical. These cylinders are rather short and keep the very long length of the mast unsupported. This causes very heavy vibrations in the mast and if resonance occurs it can fail. Thus, to align the mast at the desired angle and keep the unsupported length to a bare minimum, some additional arrangements are essential. These are termed angle hole attachments.

Figure 7.39 Bit carousel (upper) and forked wrench (lower).

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Most of the modern blasthole drills have angle hole attachments that enable mast inclinations at angles up to 30° at increments of 5°. Some have a maximum inclination angle of only 20°. In some drills the angle can be infinitely variable within their range from vertical to maximum inclination. The only rotary blasthole drill that can give reverse inclination is the Bucyrus 39R. It can give reverse inclination of 15° in addition to the normal inclination of 30° in an infinitely variable manner, as illustrated in Figure 7.40. There are three types of angle hole attachments. The first type contains a yoke provided on the main frame at the bottom of the mast as shown in Figure 7.41. The mast is raised by means of mast raising cylinders up to the desired angle and then it is firmly fixed on the yoke by means of pins or nut bolts. This alternative is used in blasthole drills where the unsupported distance of the mast above the mast raising cylinder support is not great, or in other words the mast is short.

Figure 7.40 Negative inclination angle of bucyrus 39R.

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Figure 7.41 Bottom yoke arrangement for angle locking.

The second type, usually adapted for large blasthole drills that have tall and heavy masts, consists of telescoping braces with locking fixtures at desired distances as can be seen in Figure 7.42. The mast is raised in the desired angle by using the mast raising cylinders and then the telescoping braces are locked at the holes provided within them. The third type is very similar to the second type. It consists of telescoping braces that extend by means of hydraulic cylinders. The main hydraulic mast raising cylinders tilt the mast up to the vertical and then the hydraulic cylinders provided on the braces lower the mast to the desired inclination. This type gives infinitely variable mast angles within the two limits. Attractive and advantageous as it may appear, this arrangement does not offer a great advantage over the telescoping braces because once the blasthole drill mast is set for a particular inclination, it is not required to be changed frequently since all the blastholes are to be drilled at the same inclination. When a blasthole drill frame is positioned horizontally the angle of inclination is automatically achieved. 7.5.3.7

Rotary head

The rotary head of a blasthole drill rotates the drill string and exerts feed force on the drill bit through the drill string. In rotary drills the rotary head is fabricated in the form of a box from steel plates. Power to the rotary head is given by hydraulic or electric motors. Trains of gears drive the main spindle at highly reduced speed and increase torque. The drill string is attached at the lower end of the spindle. The upper end of the spindle is attached to the air hose through a built-in swivel. An inner view of a rotary head in Figure 7.43 clearly shows all the gears used for transmitting power. In large and extra large blasthole drills the rotary speeds of the spindle on rotary blasthole drills are infinitely variable from 0 to about 125 rpm. In small blasthole drills the rotary speed range is often 0 to about 200 rpm.

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Figure 7.42 Angle hole drilling attachment through mast braces.

Figure 7.43 Inner view of a rotary head.

Some drills have dual speed hydraulic motors so the slow speed high torque or high speed low torque options can be chosen as required. Hydraulic motors enable finer control of the drill string rotation. In extra large electric drills the motors are DC type. Electric motors have high efficiency, low operating cost and low maintenance requirements compared to the hydraulic motors. A shock absorber, fixed at the lower end of the spindle, has now become almost a standard item. Shock absorbers have been described in detail in the chapter on drilling accessories.

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7.5.3.8

Feed mechanism

The feed mechanism is meant for moving the rotary head in an up and down direction and to exert necessary feed force on the drill head. As the drill head moves down, the drill string exerts force on the bit. There are different types of feed mechanisms: 1 2 3 4 5 6

Double Drum and Wire Rope Hydraulic Cylinder and Wire Rope Hydraulic Cylinder and Chain Fixed Sprockets and Chain Rack and Pinion with Chain Chainless Rack and Pinion Details of these mechanisms are given below.

7.5.3.8.1

Double drum and wire rope

This mechanism was used on blasthole drills made in the early 1950s but has now been discarded, as the other mechanisms are far superior in functi0nal efficiency. The following description is for the record. The mechanism consisted of one drum, called a shipper shaft drum, on the drill head. The other drum was that of a powered bull reel placed on the main frame. Two pieces of rope were wound on both the drums in opposite directions. When the bull reel rotated, one of the ropes was wound on it and the other unwound from it. These ropes gave up and down movement to the shipper shaft drum on the drill head and consequently the drill head itself. Possible free fall of the drill head was prevented by the brake on the shipper shaft drum. This mechanism was used for long feed travels. 7.5.3.8.2

Hydraulic cylinder and wire rope

Schematic illustration of this mechanism is in Figure 7.44. In this mechanism, when the yoke mounted on the rod of the hydraulic cylinder moves through a distance X the rotary head moves through a distance 2X in the opposite direction. In this process, the speed at which the ram moves is doubled at the drill head, but the force produced by the hydraulic cylinder is reduced to half at the drill head. As constant-pressure variable displacement hydraulic pumps are used to supply hydraulic oil to the cylinder, the feed rate gets automatically adjusted to the penetration rate. The merit of this mechanism is that it allows precise control over the feed rate by finely adjusting the hydraulic oil outflow from the variable displacement pump. The drawbacks of this mechanism are: 1 2

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Large sheaves are needed, to match the desired ratio of sheave to wire rope diameter. Thus, the cross section of the mast becomes large. Since it is not easily possible to manufacture hydraulic cylinders of length in excess of about 6.5 m, the maximum travel of the rotary head is limited to about 13 m.

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Upper Sheave or Sprocket

201

Upper Rope or chain Fixing Block Wire Rope or chain

Front Member of the Mast Through Which the Drill Head Moves Up or Down

Yoke

Rod of Feed Cylinder Drill Head Mast Shock Absorber

Drill Pipe Lower Sheave or Sprocket

Hydraulic Feed Cylinder Lower Rope or chain Fixing Block

Figure 7.44 Hyd. cylinder and wire rope feed mechanism.

3 4

The wire rope is required to be tightened and lubricated very often. The yoke needs a guide for its movement.

Occasionally, in this mechanism the hydraulic cylinder is fixed to the yoke rather than the mast, and the piston rod of the cylinder is fixed in the mast i.e. exactly opposite placement of the cylinder to the one shown in Figure 7.44. This gives higher retract force and lesser pulldown. 7.5.3.8.3

Hydraulic cylinder and chain

This mechanism is identical to the hydraulic cylinder and wire rope mechanism described above and shown in Figure 7.44, except that the wire rope is replaced by a chain. It has all the limitations cited earlier but the use of a chain allows use of small diameter idler sprockets in place of large diameter sheaves. This, in turn, allows the mast cross section to be reduced. The maintenance required in tightening and lubrication of chain is also less as compared to that of wire rope. This mechanism is adopted on many of the current small and medium blasthole drills. 7.5.3.8.4

Fixed sprockets and chain

The fixed sprocket and chain mechanism comprises two pairs of components. One pair is positioned on the left of the mast and other on the right of the mast. Each pair

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consists of two upper and two lower sprockets, two idler sprockets, one drive sprockets and a chain as shown in Figure 7.45. Two ends of each of the chains are attached to the drill head – one from the bottom and one from the top. When the shaft common to both the drive sprockets is rotated by an electric or hydraulic motor, through a reduction gear box if necessary, the sprockets pull the chain up or down and the rotary head also moves up or down. In most of the cases a single strand chain is used for sturdiness and ease in maintenance, but in a very few cases a multi-strand chain may have to be used. The assembly comprising the motor, reduction gear box, drive shaft and driving sprockets is kept near the lower end of the mast for ease of maintenance and lower center of gravity. This mechanism can be used for very tall masts meant for drilling with long drill pipes. 7.5.3.8.5

Rack and pinion with chain

A schematic illustration of rack and pinion with chain mechanism is shown in Figure 7.46. In this mechanism the drill head slides through the two front main members of the mast. Each of these two main members has a gear rack from top to bottom. The pinions on the drill head intermesh with this rack. The pinions are driven through a gear box by chains and sprockets located on the drill head. The chains are moved up

Upper Rear Sprocket

Upper Front Sprocket

Mast Front Member of the Mast Through Which the Drill Head Moves Up or Down

Drill Head

Chain

Idler Sprocket Drive Sprocket

Shock Absorber

Idler Sprocket

Drill Pipe Lower Rear Sprocket

Lower Front Sprocket

Figure 7.45 Fixed sprocket and chain feed mechanism.

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Upper Rear Sprocket

203

Upper Front Sprocket

Mast Front Member of the Mast Through Which the Drill Head Moves Up or Down

Rack Teeth Driven Sprocket

Idler Sprocket Pinion

Idler Sprocket

Drill Head

Drive Sprocket

Idler Sprocket

Idler Sprocket

Shock Absorber Drill Pipe Lower Rear Sprocket

Lower Front Sprocket

Figure 7.46 Rack and pinion with chain.

and down by drive sprockets positioned on the mast which are driven by an electric or hydraulic motor. As the pinions rotate the drill head moves up or down. The racks are made from forgings that are cut and deep hardened. The pinions with a little less hardness may have to be changed, but the rack gears are extremely durable and last almost the lifetime of the drill. This mechanism can be used for very tall masts meant for drilling with long drill pipes. Nowadays this mechanism is rarely used as it requires many more components, particularly the chain, that require considerable maintenance. 7.5.3.8.6

Chainless rack and pinion

Chainless rack and pinion mechanism is shown in Figure 7.47. In this mechanism two gear racks are welded onto two main members of the mast. Each rack extends from the bottom to the top of the mast. The rotary head slides up and down between these two main members. Four pinions are built into the drill head as shown. These pinions are intermeshed with the rack gears. The pinions are driven by a hydraulic/electric motor through a reduction gear box. When the pinions rotate the drill head moves up or down. The rack gears are made from special forgings that are deep hardened after gear cutting.

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Upper Left Pinion

Upper Right Pinion Drill Head

Lower Left Pinion

Lower Right Pinion

Left Rack

Right Rack

Shock Absorber

Figure 7.47 Chainless rack and pinion feed mechanism.

The chainless rack and pinion feed mechanism is reckoned to be the best and very long lasting. The pinions with a little less hardness may have to be changed after a long time, but the rack gears are extremely durable and last almost the lifetime of the drill. This mechanism can be used for drills with a tall mast that accommodates long drill pipes. It is found mostly on the extra large blasthole drills. Maintenance efforts required for this mechanism are considerably low as compared to the other mechanisms. 7.5.3.9

Auxiliary winch

Almost all the accessories used in rotary blasthole drilling are so heavy that they cannot be manually lifted, shifted and handled. Therefore, almost every rotary blasthole drill is provided with a wire rope and an auxiliary winch. The winch is powered by hydraulic or electric motors through a planetary reduction gear box for compactness. The winch can be placed on the main frame or near the lower end of the mast as shown in Figure 7.48. Depending upon the size of the drill, and thus the weights of the accessories used with it, the single line pull of the auxiliary winches range from 15 to 50 kN. The diameter and width of the winch drum depends upon the wire rope diameter and length, which in turn are dependent upon the maximum single line pull of the winch and the height of the mast. The wire rope is taken up to the top of the mast through a few idler sheaves. It comes down from a set of sheaves placed at the top of the mast. 7.5.3.10

Tool handling jib

To enable movement of front sheave at the top of the mast to right or left, a tool handling jib, similar to the one shown in Figure 7.49, is sometimes provided at the top of

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205

Figure 7.48 Auxiliary winch.

the mast. The movement of the jib is controlled by means of two hydraulic cylinders, one of which enables the jib to telescope in and out and the other facilitates lateral shifting of the jib. 7.5.3.11

Centralizer

This item, shown in Figure 7.50, consists of an arm attached to the mast at its middle height. A long drill pipe, when it is being attached to a drill string, tends to bend in the middle because of its slenderness. This is particularly so when the mast is at an angle for drilling inclined blastholes. The curvature of the new drill pipe makes it difficult to attach it to the drill string and the drill head. Cross-thread connection and thread damage can occur easily. The centralizer keeps the alignment of the new drill pipe very close to the alignment of the drill string and reduces the possibility of cross-connection and thread damage.

7.5.4

Special purpose items

Apart from the main items described earlier in this chapter, many more items are built into a blasthole drill for fulfilling some particular needs.

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Figure 7.49 Tool handling jib.

In the absence of standardization in nomenclature, same items are also known by different names. The following are details of some items available, either as standard or optional, on blasthole drills.The order in which they appear is random. 7.5.4.1

Language name plate

Many instructions/warnings are printed or stuck through printed stickers at various places on a blasthole drill to fulfill safety regulations. These instructions are normally in the English language and in Roman Script. Blasthole drills are used in many countries where the English language is not well understood. If this language name plate option is chosen the instructions and warnings are printed in the language and script chosen by the purchaser. 7.5.4.2

Lighting

Almost all the blasthole drills are operated round the clock. On certain locations the light available during daytime hours is also insufficient for properly viewing drilling

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Figure 7.50 Centralizer supporting long drill pipe.

operations and various portions of a blasthole drill. Lighting arrangements are very essential in such circumstances. Most manufacturers provide a lighting system as a standard feature of their blasthole drill. Depending upon the size of the drill, mercury vapor, quartz-iodine or other types of lamps are provided at different parts of the blasthole drill. The number of lamps ranges from 7 to 10 and wattage of each lamp is about 75 to 100 W. Figure 7.51 shows night operation of a rotary blasthole drill from which one can get an idea about the placement of lamps. 7.5.4.3

Fast fuel fill system

This system is often known by other proprietary names. On a large blasthole drill diesel is consumed very rapidly. Even a large capacity fuel tank has to be refilled frequently without stopping the engine and drilling operation. If the tanks is filled by the usual method of rapidly pumping the fluid from a pump on

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Figure 7.51 Night operation of a blasthole drill.

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209

the tanker through a hose pipe, there is a danger of bubble formation in the tank. It can cause severe troubles in engine running. One component of a fast fuel fill system is mounted on the fuel tank itself. The other part is the fuel filling bayonet. The fuel starts flowing to the fuel tank automatically only after these two components are coupled together properly. The component on the fuel tank has an automatic level sensing device. When the tank is filled to the appropriate level the fuel flow to the tank automatically stops and the indicator light on the fuel filling bayonet gets illuminated. When the bayonet is disconnected from the fuel tank inlet there is no drop spillage of the fuel. Fast fuel fill systems are designed to ensure that fuel will be filled in the tank at rapid flow rate but without bubble formation and to a predetermined level. A schematic of a typical fast fuel fill system is shown in Figure 7.52. 7.5.4.4

Depth indicator

Many of the small and medium blasthole drills are not equipped with a computerized drilling system. In such event this item, which numerically shows the blasthole depth, becomes essential. These numerical readings are far more accurate than those judged by a driller through his visual observations or perhaps by some marking made on the mast. Instead of the old mechanical depth indicators that worked through a cable and took measurements from the position of a drill head within the mast, many drills are now provided with electronically operated depth indicators that indicate blasthole depth inside cab after sensing the drill pipe addition. More details about such depth indicators have been given in chapter 6. 7.5.4.5

Fast retract system

In rotary blasthole drilling very high feed force is required to be exerted on the drill string by the drill head. While doing so, the speed with which the drill head can travel downwards is often very low because progress of blasthole drilling itself is very slow.

Fuel Tank

Fuel Vacuum

Fuel Level Sensor Fuel Inlet

Fuel Gun or Bayonet

Fuel

Figure 7.52 Fast fuel fill system.

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For ease of controlling the feed rate the maximum feed rate at which the drill head can travel has to be kept limited. Since maximum feed rate and maximum retract rate are related to each other, due to the characteristics of actuators like electric or hydraulic motors or hydraulic cylinders, the retract speeds are also limited. It is desirable to withdraw the drill string rapidly so as to increase overall drilling speed. As blasthole depth is very shallow, the force required to withdraw a drill string from a blasthole is also very low and it is possible to retract the drill string at fast speed by using some specialized mechanism or devices. Many rotary blasthole drills provide such a device or mechanism as an optional item that can be chosen by the customer. It is called a fast retract system. A fast retract system supplies a higher volume of hydraulic oil to the hydraulic cylinder or hydraulic motor at lower pressure, so the speed of retraction is increased considerably. 7.5.4.6

Tow hook

Blasthole drills are self-propelled and have sufficient traction power to move in rough terrain conditions. However, occasionally a blasthole drill has to be pulled or towed by another vehicle like a dozer or a tractor. To enable such pulling or towing operation one or more tow hooks must be provided on the blasthole drill at a specific position so the operators do not use incorrect positions for slinging of the wire rope. Most rotary blasthole drills have this item as an optional extra. Apart from the pull load sustaining ability, a tow hook must have self locking arrangements for safety. 7.5.4.7

Fire extinguishers

Fire gives rise to smoke, flames and heat that cause damage to equipment. Fire in a blasthole drill can be disastrous not only because a blasthole drill is an expensive piece of equipment but more so because its repair can extend over a long period. Replacement of a rotary blasthole drill by a rented piece is very difficult. In either case it results in very high production losses. An automatic fire detection and extinguishing system described in the next section is very effective, but many small blasthole drills are equipped with lightweight fire extinguishers placed at different locations in a blasthole drill that can be operated manually like the one shown in Figure 7.53. In case of fire, the nozzle of the fire extinguisher is directed towards the root of the flame and a switch on the handle is pressed. The orifice in the nozzle opens and the gas stored in the fire extinguisher under high pressure is very rapidly blown at the fire to create an inert gas surrounding the fire. Fire usually gets extinguished when oxygen concentration around the fire drops below 15%. Gases used in fire extinguishers are required to be the inert type, not only so that they should not burn themselves at high temperature or aid burning of other gases, but they must not decompose to form any component having such undesirable properties. Further, they must be safe for human beings.

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211

Figure 7.53 Lightweight fire extinguisher.

Inert gases are best suited for fire extinguishing purposes. Commonly used gases in a fire extinguishers are nitrogen and argon. Carbon dioxide also does not decompose very easily and hence finds use in fire extinguishers. To reduce the cost of refilling, some fire extinguishers have a mixture of the above gases. In such a mixture, a high argon proportion is very essential because it is heavier and settles down on the floor where the root of fire normally exists. Some fire extinguishers have gases in liquid form, but in most handheld fire extinguishers, chemicals that generate a high volume of specific gases are used in place of gases in liquid form. To increase their effectiveness in extinguishing Fire, some fire extinguisher also have arrangements of spraying non inflammable foaming agents. These foaming agents form a film around the gas bubbles and do not allow the oxygen in the atmosphere to mix and continue the combustion.

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7.5.4.8

Automatic fire suppression system

Sensing of fire is possible through different types of electronic sensors some of which are schematically shown in Figure 7.54. Photoelectric or ionization type smoke detectors detect fire by sensing smoke generated by fire. In photoelectric sensors a ray of light from a light source is normally not sensed by the light sensors but when smoke particles enter the chamber they scatter the light from the light source and it is sensed by the light sensor. In ionization-type sensors the air in the chamber is continuously ionized by a radioactive source. So long as the air does not contain smoke a small current flows continuously from one plate to the other due to electric conductivity of the gases in the air. When smoke enters the space between two plates the conductivity of gases reduces and electric current discontinues. Smoke sensors can give a false alarm if the dust concentration in the air increases above certain level. Heat sensors are better suited in dirty or dust-polluted environments that often surround blasthole drills. The most common heat detectors either react to a broad temperature change or a predetermined fixed temperature. Heat detectors use a set of temperature-sensitive resistors called thermistors, the resistance of which decreases as the temperature rises. One thermistor is sealed and protected from the surrounding temperature while the other is exposed. A sharp increase in temperature reduces the resistance in the exposed thermistor, which allows a large current to activate the detector’s alarm. Flame detectors are devices that look for specific types of light (infrared, visible, ultraviolet) emitted by flames during combustion. When the detector recognizes this light from a fire, it sends a signal to activate an alarm. In automatic fire suppuration system one or more large size fire extinguishers are kept together, as shown in Figure 7.55, and connected to a network of tubes. These

Ionization Type Sensor

Photoelectric Detector

Radioactive Source

Ionized Gas

Unobstructed Current

Light Source Smoke Particles

Light Sensor

Heat Detector

Sealed Thermistor

Exposed Thermistor

Figure 7.54 Fire sensors.

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Sensor Electric Line

Nozzle

213

Fire Sensor

Alarm Gas Line

Control Cylinders Filled with Fire Extinguishing Gas

Figure 7.55 Fire extinguishing set up.

tubes have outlets at many places on the blasthole drill where the possibility of fire exists. One or more types of sensors are also provided at such places. When fire is sensed by sensors at a particular Location, the electronic device gives an audible alarm and automatically operates the fire extinguisher nozzle opening mechanism, and the gas contained in the extinguisher flows through the particular tubing to one or more outlets. When the gas comes out, the oxygen concentration at that point reduces below 15% and the fire is put out. An automatic override to the system is also provided when the heaters start heating the air or the atmosphere becomes very hot. 7.5.4.9

Welding outlet

In blasthole drilling practice, sometimes it becomes necessary to resort to welding for temporarily overcoming difficulties. Welding requires direct current of high amperage. Some drills, particularly the extra large all-electric type, have DC electric power for driving the DC electric motors. An outlet made available for such current is the welding outlet. 7.5.4.10

Automatic leveling system

Nowadays automatic leveling systems are being made available, as an option, on more and more rotary blasthole drills. Precision and speed in leveling a blasthole drill as well as relief to the operator are the main reasons for which automatic leveling systems are preferred by more and more users. Automatic leveling systems have different ways of working. They have sensors to sense hydraulic oil pressure or the actual inclination of the blasthole drill frame and its direction as the leveling operation begins. The logic circuit then determines the quantity

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of hydraulic oil to be pumped into or out of each of the four leveling jacks. The flow to the jacks is controlled by electrohydraulic control valves. This operation is carried out in a cyclic iterative manner so the level of the blasthole drill is corrected in stages. In every stage the blasthole drill attains a more and more precisely leveled position. The operation is stopped when the main frame of the drill is within about 0.1° from a perfectly level position. The operator is given an audio alarm after the leveling operation by the automatic leveling system is completed. He then can lock the hydraulic cylinders and continue with the next step in the operational sequence. Automatic leveling systems have an override so an operator can take charge of the drill leveling operation instead of the leveling system. 7.5.4.11

Remote propel control

Placing a blasthole drill on the exact location of a blasthole is very important from the viewpoint of effectiveness of subsequent blasting. Attaining exact placement of the drill is relatively easy in the case of vertical blastholes, but when it comes to inclined blastholes the task is difficult and a good bit of judgment is required in the operation. When a blasthole drill propels to the location of a new blasthole, the blasthole position mark gets masked by the dusthood and becomes invisible to the operator. Guidance of an assistant standing on the ground is inevitable. Small size blasthole drills are usually propelled in this manner i.e. under the guidance of an assistant. Some more sophisticated large blasthole drills have a viewing window in the dust hood and lighting arrangements so the site of the blasthole marking is visible to the operator sitting in the cab. Thus, he can position the drill quite easily. In any case it is much safer to be away from the drill while it is being propelled to the new location. To enable propelling a blasthole drill by standing on the ground, some blasthole drills have an option of remote propel. Remote controls are either cable-connected type or wireless communication type . A remote control includes a small handheld box, as shown in Figure 7.56. The box contains a device for emitting different types of digital signals based on which lever

Figure 7.56 Box for remote propel control.

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has been operated and how much it has been pressed by the driller. These signals are transmitted to a receiver on the drill through a cable or wireless transmission, depending upon the type of remote control. When the receiver on the blasthole drill receives these signals the hydraulic pumps in the drill supply hydraulic oil to the propel motors in appropriate directions. If both the hydraulic motors rotate in the same direction at same speed the drill moves in a forward (or reverse) direction. If the motors rotate in the same direction but at different speeds the drill moves in a curved path. It is also possible to rotate the drill around a central axis without moving by rotating the motors at the same speed but in reverse directions. With the help of a remote control box, a driller can stand on the ground and move the drill precisely onto the location of the blasthole collaring point. Many remote propel control boxes have automatic leveling control. The signals sent from the box are received by the automatic leveling system and the drill is automatically leveled at the collaring point. The operator can then get into the cab and start drilling operations 7.5.4.12

Air conditioners

The operator’s cab and machinery house are often equipped with air conditioners. The air conditioners are of the same type that are used for air conditioning passenger buses. Electric power to the air conditioners is supplied from a small generator in the case of diesel engine-operated drills or from the transformer in the case of electric drills. 7.5.4.13

Heaters

Heaters become essential when the drill has to operate in cold weather. Apart from the operator’s cab and machinery house, heaters may have to be fitted on the rotary head gear case and the cases of gears that reduce the speed of the propel motors. The heaters can be infrared type or heating coil type. In most cases a fan is provided near the heater so it spreads the hot air within all the space of the enclosures like cab, machinery house etc., or the enclosures provided around the gear cases. 7.5.4.14

Hydraulic test station

A rotary blasthole drill may contain only 2 to 5 hydraulic pumps but it has many hydraulic circuits meant for different operations. The pressure in each of the circuits may have to be known under certain circumstances. For this purpose an hydraulic test station, as shown in Figure 7.57, is provided on some blasthole drills. The hydraulic test station has several end points of hydraulic tubes, each emanating from a certain hydraulic circuit. A special pressure gauge can be inserted in the tube at the point and the pressure in the circuit can be easily read. 7.5.4.15

Video camera system

A driller or his assistant sitting in the operator’s cab of a blasthole drill have very very limited view of the surroundings. This is particularly true in the case of large rotary blasthole drills.

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Figure 7.57 Hydraulic test station.

From the viewpoint of safety of the blasthole drill as well as other equipment and the respective operators or uninvited intruders, it is necessary that the driller’s assistant keeps an eye on the surroundings of the blasthole drill. To enable the driller’s assistant to view the surroundings a video camera system is provided on some blasthole drills. The system comprises video cameras fixed in the blasthole drill at appropriate points. The display emanating from these cameras is shown on a screen provided in the operator’s cab as illustrated in Figure 7.58. The number of cameras is usually up to four but depending upon the necessity it can be reduced or increased.

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Figure 7.58 Display of video camera system.

7.6

EXTREME COLD OPERATION DEVICES

For operations in extremely cold weather conditions and high altitudes many devices, some of which are described here, can be specially provided on a rotary blasthole drill.

7.6.1

Engine starting aid

Diesel engines do not have any sparking mechanism. The heat generated by the compression of the air and diesel mist in the engine cylinders takes the temperature to flash point and ignition is caused to generate mechanical energy in the form of shaft rotation of the engine. When a diesel engine has to operate in extremely cold weather the temperature of the mixture itself is so low that the heat generated by compression is insufficient to take the mixture temperature to flash point. Thus, the engine does not start.

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To overcome such situations, normally an engine is provided with glow plugs within its cylinders or an alcohol vaporizer at the air intake of the engine. 7.6.1.1

Glow plugs

Glow plugs are located in the pre-combustion chamber of each of the engine cylinders. They generate heat from the electric supply and heat the air within the chamber. Thus, the temperature of the air and diesel sprayed into the chamber becomes sufficiently high for its ignition after compression. It is important to use glow plugs of specific rating only, otherwise the results can be disastrous. The glow plug electric supply is started with a switch so the glow plugs can be kept on only at the time of engine starting. 7.6.1.2

Manifold flame heater

On some engines a manifold flame heater is used for starting the engine in extremely cold weather. It consists of a housing, spark plug, flow control nozzle and two solenoid control valves. The flame heater ignition unit energizes the spark plug. The nozzle sprays fuel under pressure into the intake manifold assembly. The fuel vapor is ignited by the spark plug and burns in the intake manifold. The heat from this fire warms the manifold and the combustion chamber. Thus, the temperature of the air entering in the combustion chamber rises and the oil mist in the combustion chamber can ignite easily.

7.6.2

Batteries

In cold weather batteries lose the charge very rapidly. For extreme cold weather operation batteries of high AH rating and specialized construction may have to be used to minimize the troubles in starting a diesel engine.

7.6.3

Double wall machinery house

For protection of the electrical equipment from snow and rainwater, a machinery house is provided on most of the electric drills and all-electric drills.

Figure 7.59 Double wall of machinery house.

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Diesel hydraulic drills are rarely equipped with a machinery house as diesel engines are not so vulnerable to rainwater. Further, a machinery house on such drills can prove to be a hindrance in frequent maintenance needed by the engine and other components. When the drills have to be used in extremely cold weather even diesel hydraulic drills also have to be equipped with a machinery house by suitably directing the air intake to and air exhaust from the diesel engine. Usually a machinery house is constructed from steel plates of adequate thickness. For operations in extremely cold weather the machinery house can be made from panels that have two plates welded on square tube steel sections as shown in Figure 7.59. Air entrapped within the steel plates acts as an insulator. To reduce the heat conduction further, the plates and tubing are coated with special paints.

7.6.4

Machinery house heater

Whether single wall or double wall, a machinery house is required to be heated so the temperature of the air inside the enclosure is comfortably high for machines to work and service personnel to carry out maintenance procedures. Heaters can be provided in machinery house for this propose. These heaters are heating coil or induction type. They have fans for pushing the hot air in different directions. Depending upon the range of low temperatures and the size of the machinery house, heaters can have ratings even up to 50 kW. The heater has a control to increase or reduce the quantity of heat generated.

7.6.5

Operator cab heater

For use in cold weather, an operator’s cab is often equipped with a cab heater. For extremely cold weather this heater proves insufficient, so another heater is provided. The rating of this additional heater can be up to 10 kW. This additional heater can be switched off during the daytime or in summer when the temperatures are not at extremes. Cab heaters can also be in the form of a combination of an air conditioner and heater. As they are expensive compared to the normal heaters, they are chosen on very few occasions.

7.6.6

Other enclosures and heaters

Apart from the aforesaid specific cold weather protection components, many other assemblies in a blasthole drill store or circulate water or oil. They are not located within the machinery house. If the temperatures in cold weather are not at extremes, insulation of these assemblies can prove sufficient, but for operations in extreme cold weather these assemblies are required to have special devices in the form of enclosures and heaters. Heaters are heated coil or infrared type. Hot air generated by the heaters is circulated by means of fans in the enclosures so the temperature around the assemblies is kept at acceptable level. The following are some of the assemblies that are provided with such enclosures and heaters: 1 2

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Water Tank Hydraulic Oil Tank

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3 4


Propel Gear Case Rotary Head

7.6.7

Protective coating on steel components

Heat is always generated in many components of a blasthole drill when it is working. This heat keeps most of the drill components and structural members at temperatures at which the mechanical properties of the materials remain at an acceptable level. To reduce the flow of heat from the drill components and the structural members to the extreme cold weather around them a coating of special chemicals and paints are can be used. Manufacturers coat the components and structural members of their drill with such insulating chemicals/paints when a user orders such coating at the time of its purchase.

7.7

COMPARISON OF TYPES OF BLASTHOLE DRILLS

As seen earlier, blasthole drills work either on internally generated diesel engine power or externally available electric power. The power is then distributed to various actuators in the drill through a hydraulic or electric medium. The difference in the mode of power distribution makes it necessary to have different devices on respective drills as given in Table 7.8. The following are some comparative comments in respect of the power source: 1 2

Drills with diesel engines have unrestricted mobility, whereas the mobility of the electric drills is restricted by the length of power supply cable in the cable reel. Additional items required with all-electric drills are costly so all-electric drills cost more than equivalent diesel drills.

Table 7.8 Need for some devices on the drill.

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Aspect

Diesel engine power source

External electric power supply

Fuel Tank Radiator Air Cleaner Battery Starting Transformer Converter Cable Reel Fuel Oil Engine Oil Transformer Oil

Needed Needed Needed Needed Not needed Not needed Not needed Needed Needed Not needed

Water Coolant Enclosure

Needed Not needed

Not needed Not needed Not needed Not needed Needed Needed Needed Not needed Not needed Not needed for Dry Transformers Usually not needed Needed

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3 4

5

6 7 8

221

The cost of power generated by diesel engines is very high as compared to that of an external electric supply. Consumable requirements of diesel powered drills is very high compared to allelectric drills. Transformers used on modern drills are of dry type and do not require transformer oil. Reliability of power supply from an engine is excellent provided the regular maintenance procedures are followed. External electric supply is also equally reliable or perhaps better, but without much maintenance. Exhaust of a diesel engine causes air pollution to some extent. External power supply does not cause any local air pollution. Maintenance requirements of various components in the case of diesel engine power are far greater as compared to external electric power. Starting a diesel engine in cold weather is difficult. In extremely cold weather a diesel engine may have to be run all the time irrespective of the use of power. In case of electric motors and other devices that operate on external electric supply the difficulties in starting and running are much less.

A comparison between hydraulic power distribution systems and electric power distribution on a rotary blasthole drill is given in Table 7.9. Specific elaborations in this regard are as follows: 1

2

3

4

Hydraulic oil used in hydraulic power distribution systems is a consumable item. It is required to be changed periodically and incurs expenditure. No such consumable exists in electric power distribution. Hydraulic power distribution systems require many components to carry and control the hydraulic oil flow as compared to the electric power distribution system. Power cables used for electric power transmission from source to the actuator are virtually jointless but the tubes and hoses used for carrying hydraulic oil contain bents, tees and other similar components. Once the electric cables are laid properly they or components within them may not have to be changed for as much as two decades. This is because the electric flow does not exert any physical pressure on the component hence no question of contraction, expansion and resulting fatigue or rupture.

Table 7.9 Comparison between hydraulic power distribution and electric power distribution.

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Aspect

Hydraulic power distribution

Electric power distribution

Hydraulic Oil No. of Components Type of Control Distribution Losses Component Weight Control or Motor Speed Maintenance Component Life

Needed Many Electrohydraulic Very High Motors are Lightweight Excellent High Maintenance Relatively Short

Not needed Fewer Electric Low Motors are Very Heavy Good Very Low Maintenance Very Long

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5

6


In hydraulic systems physical pressure is exerted on tubes, hoses, valves etc. Apart from their susceptibility to rupture or leakage the fatigue resulting from the pressure changes does reduce the life of the component. Further physical movement of parts and hydraulic oil does result in a small wear of the component. As the hydraulic motors have very low weight and low inertia they stop rotation almost instantaneously once the power is switched off. Electric motors keep on running due to inertia for a very short duration after the power is switched off. Thus, hydraulic motors can be controlled precisely. Linear movement with heavy force can be easily generated by hydraulic means but not easily with electric means.

Above point nos. 5 and 6 make it desirable to have hydraulic power distribution systems even on all electric blasthole drills for leveling jacks and propel motors. In summary it can be said that all-electric drills are very economical to use as compared to equivalent diesel hydraulic drills. However, circumstances as under dictate the use of diesel drills: 1

2

Several small and medium size surface mines use small and medium size equipment. They choose to operate the equipment on diesel power as the expense of providing a high voltage electric grid are very high. The economy achieved by use of electric machines like draglines, shovels, drills etc. is not enough to offset the cost of laying electric grid. Electric supply is usually not available at new mine sites in the beginning of mining operations so there is no alternative to using diesel hydraulic blasthole drills.

If satisfactory electric supply is expected at a later Date, it is wise to choose diesel drills in the initial phase, which at a later date can be converted to electric drills by replacing the diesel engine with an electric motor. Once the electric supply is fully functional all-electric or electric blasthole drills can be procured.

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Chapter 8

Compressed air and air compressors

8.1

INTRODUCTION

Compressed air is the invariable medium for flushing the blasthole. It plays a very important role in the blasthole drilling process because in many cases the power consumed by the air compressor on a rotary blasthole drill is more than the power consumed by any other of its devices like rotary head, feed mechanism, propel mechanism etc. This chapter deals with many aspects related to the generation of compressed air in an air compressor. Blasthole flushing is covered in a separate chapter due to its specific importance.

8.2

COMPRESSED AIR

Compressed air is nothing but atmospheric air that is pressurized in a device called a compressor. A compressor sucks the atmospheric air and applies pressure to it. In this process the air becomes thicker i.e. its density increases. To keep the density at the desired level the air absorbs and stores energy supplied by the compressor. If the valve of a container containing compressed air at high pressure is opened and compressed air is allowed to flow to atmospheric air, the compressed air flows till such time as the pressure of air in the container reduces to the atmospheric pressure. Energy stored in the compressed air in the container is used for such flow. If compressed air is made to flow through pipes, the energy stored by the compressed air also flows with it. Thus compressed air can be used as a medium of energy transmission. Every atom/molecule of the atmospheric air is being pulled towards the center of gravity of the earth as per the laws of gravitation. The weight of the atom/molecule is being passed on to the atom/molecule just below i.e. nearer to the center of the earth. Thus atmospheric air also exists at a certain pressure. The atmospheric pressure is highest at the points on the surface of the earth that lie nearest to its center and keeps on reducing as we move away from the center. Thus atmospheric pressure at an altitude is always lower than the atmospheric pressure at sea level. If we imagine a container taken far away – say 5000 km from the surface of earth – there will be no air outside the container. Naturally, the level of pressure experienced at such height will be the minimum possible. Such pressure is called absolute zero pressure.

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A barometer is used to measure the pressure of atmospheric air. Pressure gages are used for measuring the pressure of compressed air above the barometric pressure. Therefore, Absolute Pressure = Barometric Pressure + Gage Pressure. In the International Standard (SI) System of units of Measurements, pressure is measured in terms of pascals, abbreviated as Pa. One pascal equals to a force of one newton applied on one square meter area. One kilogram of mass on the earth exerts 9.80665 newton force on the earth because the gravitational constant of the earth is equal to 9.80665 m2/s. Every system in the world contains energy. A fundamental law in this regard states that energy can neither be created nor destroyed but can only be transferred and converted. Energy has to be spent in doing every form of work. As a corollary it can be said that the amount of work done on or by a system is equal to the amount of energy transferred to or from the system. Like every other physical system that exists in the world, atmospheric air also has a certain temperature. Temperature is an indicator of the intensity of heat energy contained in the system. In the International Standard (SI) System of units of Measurement, temperature is measured in Celsius and is abbreviated as °C. Originally the same unit of measure was called centigrade. When the purest form of water is about to start freezing its temperature level is considered to be 0 and when it was about to start boiling it is considered to be 100. One hundredth part of the difference between these two levels is considered equal to 1°C. If, in some way, energy from a system is removed completely, the system will attain the minimum possible temperature called absolute zero temperature. The level of absolute temperature is 273.16°C below the freezing temperature of water. When a temperature of any system measured in °Celsius is TC, we have the absolute temperature of the substance as TC + 273.16. Kelvin is a unit of temperature measurement on the absolute scale basis. Magnitude of one degree Kelvin is the same as one degree of Celsius. Therefore, TK = TC + 273.16 Figure 8.1 clarifies the concept of absolute and gage pressure as well as absolute temperature measured in °K and thermometer temperature measured in °C.

Pressure Measured by Gauge

Temperature Measured by Thermometer

Gage Pressure

Thermometer Temperature in °C

Absolute Pressure

Absolute Temperature Atmospheric Pressure Measured by Barometer

Lowest Possible Pressure

273.16°C

Lowest Possible Temperature

Figure 8.1 Concept of absolute pressure or temperature.

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Pressure, volume and temperature are the fundamental properties of compressed air. The subject that deals with these is called Thermodynamics. When air expands or gets compressed the expansion or compression can take place in two manners viz. isothermal or isoentropic. In isothermal expansion or compression, the change in volume takes place without any change in the temperature of the system. In isoentropic expansion or compression the change in volume takes place without gaining or losing the heat in the system.

8.2.1

Gas laws

Behavior of gases is in accordance with the following laws: 1 2 3 4 5 6 7

Boyle’s Law Charles’ Law Guy Lussac’s Law Joule’s Law Poisson’s Law Amagat’s Law Avogadro’s Law The principles elaborated by these laws are as given below.

8.2.1.1

Boyle’s law

Boyle’s law states: “The absolute pressure of a given mass of perfect gas varies inversely as its volume when the temperature remains constant.” i.e. p ∝ 1/v. In other words, if the absolute pressure and volume of a gas system before the process were p1 and v1, and became p2 and v2 after the process, then p1 * v1 = p2 * v2 = Constant As the temperature of the gas remains constant the process is called an isothermal process. 8.2.1.2

Charles’ law

Charles’ law states: “The volume of a given mass of perfect gas varies directly as its absolute temperature when the pressure remains constant.” i.e. v ∝ T. In other words, if the absolute temperature and volume of an air system before the process were T1 and v1 and became T2 and v2 after the process, then v1/T1 = v2/T2 = Constant As the absolute pressure of the gas remains constant the process is called an isobaric process.

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8.2.1.3

Guy lussac’s law

Guy Lussac’s law states: “The absolute pressure of a given mass of perfect gas varies directly as its absolute temperature, when the volume remains constant.” i.e. p ∝ T. In other words, if the absolute pressure and absolute temperature of an air system before the process were p1 and T1 and became p2 and T2 after the process, then, p1/T1 = p2/T2 = Constant As the volume of the gas remains constant the process is called an isochoric process. 8.2.1.4

Joule’s law

Joule’s law states: “The change in internal energy of a perfect gas is directly proportional to the change of temperature.” i.e. ΔE ∝ ΔT. In other words, if the temperature of a mass of a gas before the process was T1 and it became T2 after the process, then ΔE = m * C * ΔT = m * C * (T2 − T1) Where m is the mass of the gas and C is a constant of proportionality called specific heat. 8.2.1.5

Poisson’s law

Poisson’s law states that in the case of a gas, for a process without any heat exchange with the surroundings, the relationship between pressure and volume is as per the following equation p1 * v1(Cp/Cv) = p2 * v2(Cp/Cv) In the equation p1 and v1 represent pressure and volume before the process and p2 and v2 represent pressure and volume after the process. Since no heat is exchanged with surroundings, the process is isoentropic compression or expansion. The quantities represented by Cp and Cv are the specific heat capacities for the gas at constant pressure and constant volume. Often the ratio Cp/Cv is represented by k and has values as follows: k = 1.66 for monoatomic gases k = 1.40 for diatomic gases k = 1.30 for triatomic gases Since more than 98% of the gases contained in atmospheric air are diatomic i.e. those gases where two atoms of the gas form a stable molecule (e.g. nitrogen, oxygen), for atmospheric air the ratio k is taken as 1.40. The ratio Cp/Cv is fairly constant at low pressures.

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8.2.1.6

227

Amagat’s law

Amagat’s law states that the volume of a mixture of gases is equal to the sum of the partial volumes that the constituent gases of the mixture would occupy if each of them existed alone at the total pressure of the mixture. 8.2.1.7

Avogadro’s law

Avogadro’s law states that under the same conditions of pressure and temperature, equal volumes of all gases contain an equal number of molecules. In other words 1 m3 of hydrogen, in its molecular form H2, will contain the same number of molecules as contained in 1 m3 of oxygen in its molecular form O2, when the temperature and pressure of both systems are the same. In such cases, since the molecular weights of hydrogen and oxygen are 2 and 32 respectively, the densities of hydrogen and oxygen will be in proportion of 2:32 i.e. 1:16. A mole is a scientific term given to a numerical quantity of molecules. The number of molecules contained in a mole is 6.02257 * 1027 and is often termed as Avogadro’s Number. It can be shown that at the NTP conditions i.e. temperature of 0°C and pressure of 101.325 kPa, the volume of one mole of any gas equals to 22.3933 m3. 8.2.1.8

General gas law

By combining Boyle’s law and Charles’ law we can write an equation p * v/T = Constant. This formula is followed in all the relevant gas calculations. Following example illustrates its use. EXAMPLE

Initial conditions of a gas before compression are volume = 1 m3, gage pressure = 300 kPa, temperature = 27°C. It is compressed to attain a volume = 0.2 m3, temperature = 90°C. What will be the final gage pressure? SOLUTION

The initial conditions of volume, absolute pressure and absolute temperatures are, v1 = 1 m3,

p1 = 300 + 101.325 = 401.325 kPa,

T1 = 27 + 273.16 = 300.16°K Post compression conditions are, v2 = 0.2 m3,

T2 = 90 + 273.15 = 363.16°K

Applying the general gas law we get, (401.325) *(1)/(300.16) = p2 * (0.2)/(363.16)

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This gives p2 = 2427.79 kPa. Since p2 is the final absolute pressure the final gage pressure will be 2427.79 − 101.325 = 2326.465 kPa.

8.2.2

Power required for air compression

An air compressor is a device that takes power from an external power source like a diesel engine or electric motor and apples it mechanically to the air so as to reduce its volume and increase its pressure. In such compression processes heat is neither gained nor lost by the system. The compression process is therefore adiabatic. A schematic illustration of a typical piston compressor is shown in the upper drawing in Figure 8.2. The lower drawing shows an indicator diagram of the compressor work in a complete cycle. The piston moves to left and right in the cylinder to compress the air. When the piston is at the beginning of the compression stroke (point C, pressure p1 and volume v1)

Outflow Pipe

Outflow Valve

Inflow Inflow Valve S

E

Piston at the end Piston at the of Compression beginning of Stroke Compression Stroke D pvn = Constant Δp

V P2 V2

T

B

A

C

P1 V1

Figure 8.2 Cycle for isothermal compression of air.

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the intake valve is open but as soon as it moves a little to left the air in the cylinder gets compressed and the inflow valve closes. From here the compression starts as per Poisson’s law along the curve CD. When the piston is at a certain intermediate position (point D) the pressure inside the cylinder becomes equal to the pressure in the outflow pipe and the outflow valve opens. From this point onwards as the air flows into the outflow pipe the pressure of the air does not increase. This continues till the piston reaches the end of the compression stroke (point E, pressure p2 and volume v2) As the piston starts moving to the right the outflow valve closes and the pressure of the air contained in the cylinder starts decreasing. At a certain intermediate position (point B) of the piston the pressure inside the cylinder becomes lower than the pressure in the inflow pipe and the inflow valve opens. The piston continues its movement towards the right till the end of the stroke but the pressure in the cylinder remains the same. The volume of air contained in the cylinder when the piston is at the end of compression stroke is called the clearance volume. For air compression or expansion process a general equation, p * vk = K, can be applied. For isothermal compression or expansion k = 1 as per Boyle’s Law. For isoentropic compression or expansion k = 1.4 as per Poisson’s Law. In actual air compression by an air compressor the compression process is associated with a change in temperature as well as a change in energy. Thus the process is neither isothermal nor isoentropic but called polytropic. For positive displacement compressors the value of k ≈ 1.3. The net power required by the compressor to compress air is equal to the work done in compressing the air. Most of the compressors used on rotary blasthole drills are single stage rotary vane or rotary screw compressors i.e. positive displacement compressors. For such compressors the net power requirement can be found from the following equation for work done in one cycle of compression. W = [k/(k − 1)] * p1 * v1 * [(p2/p1)[(k−1)/k] − 1] If the compressor is rotating at a speed of N rpm i.e. N/60 revolutions per second then the power requirement will be P = {k/[(k − 1) * 60]} * p1 * v1 * N * [(p2/p1)[(k−1)/k] − 1] In the above equation, W = Work done per cycle in Nm k = Exponent for Polytropic Process p1 = Initial absolute pressure in kPa p2 = Final absolute pressure in kPa N = Compressor rotary speed in rpm v1 = Initial volume sucked in compressor m3 P = Net power required by compressor kW The volume v1 is taken as the discharge capacity of the compressor.

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In some compressors a small volume of the compressed air does not go into the high pressure section but returns to the atmosphere. If this returned volume is taken as vs then the volumetric efficiency of the compressor is taken as ηv = (v1 – vs)/v1 If the compressor on a rotary blasthole drill is a two-stage compressor, it is normally equipped with an intercooler which cools the air compressed in the first stage and then sends it to the second stage for further compression. The cooling of the air in the intercooler can either be complete i.e. the temperature of the air compressed in the first stage is lowered to the atmospheric temperature or incomplete i.e. the temperature of the air compressed in the first stage is not lowered to atmospheric temperature. For compressors with complete intercooling the equation for the power requirement is P = {k/[(k − 1) * 60]} * p1 * v1 * N * {(p2/p1)[(k−1)/k] + (p3/p2)[(k−1)/k] − 1} and for compressors with incomplete intercooling the equation for power requirement is P = {k/[(k − 1) * 60]} * N * {p1 * v1 * [(p2/p1)[(k−1)/k] − 1] + p2 * v2 * [(p3/p2)[(k−1)/k] − 1]} In both the above equations, p1 = Initial absolute pressure in kPa p2 = Interstage absolute pressure in kPa p3 = Final absolute pressure in kPa all the other terms have the same meanings as in the earlier case of single stage rotary vane or rotary screw compressor. Usually the compressor characteristics state the compressed air volume delivered along with final pressure of compressed air. In such cases the initial pressure is treated to be the one defined by the rating conditions. For such situations the whole volume is considered to have occurred in one cycle and, therefore, the value of N is taken as 1 while evaluating the power requirement by the above equation. The following is a sample example that calculates the power required by a single stage rotary vane compressor. In many cases the conditions of temperature and pressure are expressed as NTP or STP rather than expressing specific values of temperature and pressure. In such instances NTP means temperature of 273°K and pressure of 101.325 kPa and STP means the standard atmospheric condition defined by the particular standard (SAE, BS, DIN etc.). These are specified for some standards in Table 11.3 of chapter 11. Two other terms that are often heard in connection with compressed air are SCFM and ACFM. Here CFM means the discharge from compressor in imperial unit cubic feet per minute. S stands for standard conditions.

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Due to the presence of moisture in the air the actual discharge is a little less than that given by the SCFM rating. This actual discharge is termed ACFM. Pressure is often given in terms of MPa, kPa, bar or psi. In such case 1 MPa = 10 bar = 1000 kPa = 145.038 psi. Similarly, for flow 1 m3/min = 35.31466 ft3/min. = 0.01666 m3/s = 16.666 L/s. EXAMPLE

A vane compressor on a blasthole drill delivers 30 m3 compressed air at a gage pressure of 6.89 bars at NTP condition. What is the power required by the compressor to compress the air? SOLUTION

In this case we presume that the compression is polytropic and the value of k = 1.3. For NTP conditions we have p1 = 101.325 kPa. Final gage pressure is 6.89 bars i.e. 689 kPa. Thus, absolute pressure p2 is 689 + 101.325 = 790.325 kPa. Volume delivered in one minute is equal to 30 m3. Evaluating the above equation with these values gives the net power requirement as P = 133.138 kW.

8.3

FLOW OF COMPRESSED AIR

The compressed air generated by a compressor on a rotary blasthole drill has to flow through a specific flow path. Conduits of circular cross section are far better in withstanding the high pressure of compressed air without undergoing distortion than conduits with other shapes of cross section. Therefore, almost invariably a circular cross section is chosen for the flow path. As far as rotary blasthole drilling is concerned, compressed air is used for flushing. For this purpose it is made to flow through circular pipes and pipe fittings as shown in Figure 8.3. If the flow path schematic in Figure 8.3 is minutely observed one can easily comprehend the following. After a little drop in elevation the compressed air flow path horizontally goes up to the mast. Steel tubes containing valves, bends etc. are used as a flow path. The length of this flow path can be from 4 to 17 m depending upon the size of the blasthole drill. A further part of the flow path, that takes compressed air up to to the steel tubes on the mast, has to be curved and must have flexibility because the mast may have to be occasionally lowered and raised. Therefore, this part of the flow path is made up of hose pipe. The length of this part is between 1 to 3 m and a small part of the hose is vertical. Further flow path is formed of straight steel tubes. It remains vertical during the drilling operation. The length of this part can be between 6 m to even 20 m depending upon the drill size.

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Figure 8.3 Flow path in a typical blasthole drilling process.

From the vertical tubing the flow path to the moving rotary head of the blasthole drill is in the form of a hose. The length of the hose is between 8 to 20 m. This hose is always bent through a 180° curve. Naturally the air flows through it vertically, in the up as well as down direction. In the drill head again some steel tubes and other components take the compressed air to the drill pipes. The length of this part is between 1 to 3 m. Some of the portion is horizontal and some is vertical. The drill pipes take the compressed air to the tricone bit. This flow path can have a length from 6 to even 60 m. In this section air flows in a vertically downward direction. In order to clean the bottom of the hole very effectively, the flow of air is passed through the three nozzles of the tricone bit. In this process the pressure energy stored

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in the compressed air is converted into kinetic energy. Thus the speed of air increases considerably and the jet is able to clean the blasthole bottom in an effective manner. Finally the compressed air flows upward in the annulus along with cuttings of rock and escapes to the blasthole top. The large cuttings fall on the ground around the blasthole, whereas the air mixed with small sand and dust size cuttings is made to flow through a dust collector. After the filtration dustless air escapes to the atmosphere. The flow of compressed air is studied through the subject called ‘Fluid Mechanics’. From the above description of flow path in a blasthole drill, it is obvious that following topics of fluid mechanics must be considered. Flow through steel pipes Flow through fittings Flow through hose pipe Flow through nozzles Flow through annulus. Of the above five topics the last i.e. flow through the annulus is considered in chapter 11. When the pressure of flowing compressed air remains almost constant, the flow is considered as incompressible flow. In the compressed air flow path described above the pressure drop between the compressor to the top of the drill bit is usually very small, of the order of 35 to 50 kPa. Thus, the flow of compressed air through pipes comes under the realm of incompressible flow. When the pressure at a point on the flow path differs significantly from that at other point on the flow path, the flow is treated as per the technical formulation of compressible flow. Thus, the flow of compressed air through nozzles or orifices is compressible.

8.3.1

Compressed air flow in steel pipes

When compressed air flows through a flow path, energy stored in the compressed air is lost due to the friction of air molecules with the walls of the flow conduits. As a result, the pressure of compressed air reduces. It is important to know the quantum of pressure lost in the passage of compressed air up to the drill bit, or ideally up to the point where the compressed air emerges out of the blasthole i.e. at the mouth of the blasthole. The properties of air, the velocity of air flow, and the roughness of wall that comes in contact with the atoms of flowing air, make the state of the flow either smooth or disturbed or transient (when it is changing from smooth to disturbed). These states are referred to as laminar or turbulent or transitional, respectively. In laminar flow the molecules of air move in a straight line parallel to the axis of the pipe, but in turbulent flow the molecules move in a very haphazard manner, and in addition to the axial velocity also have velocity components in a plane perpendicular to the axis of the pipe. As far as the roughness of pipe is concerned, on a microscopic dimensional basis even the best polished surface of any pipe has ups and downs in the surface as shown

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in Figure 8.4. The average distance between the extent of these ups and downs is called absolute roughness of the pipe. It is commonly expressed by symbol ε. Further, a pipe has some inner diameter D. Relative roughness is the ratio ε/D. It can be easily calculated for any pipe from the values of absolute roughness of different pipe materials given in Table 8.1. The most important indicator of the type of flow is called Reynold’s Number. Reynold’s number is actually a dimensionless number, usually denoted by Re. It is related to some variables of the pipe, flowing fluid and the flow as follows: Re = (D * W * ρ)/ν where, Re = Reynold's number D = Inner diameter of pipe in m W = Flow velocity in m/s ρ = Density of fluid in kg/m3 ν = Dynamic viscosity of fluid in Pa · s Centreline of the pipe Relative roughness works out to ε/D Radius of the pipe = D/2 D is the diameter of the pipe Absolute roughness ε

Figure 8.4 Absolute and relative pipe roughness. Table 8.1 Absolute roughness values (ε) of some pipes.

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Material and type of pipe

Absolute roughness in mm

Copper, Lead, Brass, Aluminum (new) PVC and Plastic Pipes Stainless steel Steel commercial pipe Stretched steel Weld steel Galvanized steel Rusted Steel New Cast Iron Worn Cast Iron Rusty Cast Iron Sheet or Asphalted Cast Iron Smoothed Cement Ordinary Concrete Coarse Concrete

0.001–0.002 0.0015–0.007 0.015 0.045–0.09 0.015 0.045 0.15 0.15–4 0.25–0.8 0.8–1.5 1.5–2.5 0.01–0.015 0.3 0.3–1.0 0.3–5

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Dynamic viscosity of compressed air at various temperatures and pressures is given in Table 8.2. When Reynold’s number of flow is less than 2300 the flow is laminar. For Reynold’s number of 4000 or more, the flow is turbulent and in between 2300 and 4000, the flow is transitional. The most widely used equation for determining pressure loss in pipes is the DarcyWeisbach equation, viz. Δp = f * (L/D) * 10−3 * {(ρ * W2)/2} where, Δp = Pressure loss in kPa f = friction factor L = Length of pipe in m D = Hydraulic diameter of pipe in m ρ = Density of flowing air in kg/m3 W = Average flow velocity in m/s The assumptions made while deriving the above equation from fundamentals are 1 2 3

The pipe is without any curvature. Pipe roughness is uniform throughout its length. The pipe is horizontal. Thus the gravitational forces acting on the mass of flowing air do not change and can, therefore be neglected.

In the above equation all the other factors except D and f are either directly or indirectly known. The factor D in the Darcy-Weisbach equation, i.e. hydraulic diameter of the pipe, can be found by an equation viz. D = 4 * A/P where A = Cross sectional area of flow path. P = Wetted perimeter of the flow path.

Table 8.2 Dynamic viscosity of dry air.

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Absolute pressure in kPa

Value of dynamic viscosities of dry air in Pa · s ( i.e. kg/(m · s)) at different temperatures in °C −25

−20

−10

0

10

20

25

30

40

50

100 300 500 700

15.9000 15.9339 15.9700 16.0083

16.1504 16.1824 16.2166 16.2530

16.6348 16.6648 16.6965 16.7299

17.1000 17.1296 17.1600 17.1913

17.5492 17.5792 17.6091 17.6390

17.9856 18.0161 18.0458 18.0748

18.2000 18.2304 18.2600 18.2887

18.4124 18.4426 18.4718 18.5001

18.8328 18.8612 18.8893 18.9170

19.2500 19.2743 19.3000 19.3270

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For a circular tube it is 4πr2/πr = 2r i.e. diameter of the pipe. For the annular flow path between a blasthole of radius r and a drill pipe of outer radius r1 it is 4π(r2 − r12)/2π(r + r1) = 2(r − r1). For other shapes it can be easily calculated. When the flow conduit is in the form of a pipe with circular cross section the Darcy-Weisbach equation can be written as, Δp = f * (L/D) * 10−3 * {(8 * ρ * Q2)/(π2 * D4)} This form is particularly suitable for pneumatic calculations because here Q is in terms of m3/s and the compressor discharge, which is usually given as m3/min can easily converted into m3/s by dividing the discharge value by 60. If the flow rate of air is given in terms of mass of compressed air then the above equation gets converted into Δp = f * (L/D) * 10−3 * {(8 * m2)/(ρ * π2 * D4)} Where m is the mass flow in kg/s. In aforesaid Darcy-Weisbach equation the factor {(ρ * W2)/2} is often called dynamic pressure. Obviously it is equal to {(8 * ρ * Q2)/(π2 * D4)} or {(8 * m2)/ (ρ * π2 * D4)} as the case may be. The value of the friction factor f has a complex relationship with Reynold’s number of the flow and relative roughness. The approximation that is considered to be the most accurate is given by the Colebrook Equation, which is (1/f0.5) = −2.0 * log[{(ε/D)/3.7} + {2.51/(Re * f0.5)}] This equation is implicit. Therefore, the evaluation of f from this equation can be done only through an iterative process. In the days when electronic calculators or computers did not exist, a chart called Moody’s chart, shown in Figure 8.5, was used for evaluating f. It is almost a graphical representation of the Colebrook equation. In the era of calculators i.e. during 1970 to 2000 some simplified and explicit formulae were suggested for determining the friction factor f. Now, in the present era of very fast personal computers, generally f is calculated by using the Colebrook equation. Computer software, some of which is totally free, is available on the internet and can be downloaded for the purpose of calculation of pressure loss. It must be remembered that the friction factor f is sometimes referred to as Moody’s Friction Factor. The pressure loss measurements taken in the field and those calculated by the above equations do tally acceptably well in the case of liquids, but for air and gases the differences are quite high. In almost all blasthole drills, the capacity rating of the compressor and the diameters of the pipes used for carrying the compressed air to the drill head, the drill pipe, and the annulus formed by the walls of the hole and the drill pipes, are such that the flow through them is turbulent. The following example will clarify the method of calculating pressure loss in a pipe.

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Moody Friction Factor

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Reynold's Number Figure 8.5 Moody’s chart showing relation between friction factor, Reynold’s number and absolute roughness. 11/22/2010 2:38:16 PM

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EXAMPLE

A straight galvanized pipe of circular cross section with a diameter 140 mm and length 75 m is used for carrying discharge from a compressor having a capacity of 25.49 m3/min at a pressure of 1034 kPa. The temperature of the compressed air is 15°C and atmospheric pressure is 101.325 kPa. Find out the pressure loss. SOLUTION

Since the atmospheric pressure is 101.325 kPa, we get the actual air flow density ρ = (1034 + 101.325)/ (287.1 * (273.16 + 15)) = 13.723134. As the compressor discharge is 25.49 m3/min, the actual volume flowing through pipe will be = (25.49/(13.723 * 60)) * = 0.0309575 m3/s. A = 3.14159 * 0.142/4 = 0.0153937 m2 Therefore, W = 0.0309577/0.0153937 = 2.011033 m/s. With the above, the dynamic pressure ρ * W2/2 = 13.723134 * 2.0110662/2 = 27.749934. Since the pipe is galvanized, ε = 0.15 mm and relative roughness ε/D = 0.15/140 = 0.0010714. Dynamic viscosity at STP conditions μ will be 1.802 * 10−5. Reynolds Number Re will be ρ * D * W/μ 13.723 * 0.14 * 2.011033/1.802 * 10−5 = 214410. From the Moody chart the friction factor f will be 0.0195 and the pressure loss will work out to Δp = f * (L/D) * [(ρ * W2)/2] Δp = 0.022 * (75/0.14) * 27.749934 Δp = 313.96 Pa. EXAMPLE

Find out the air pressure loss experienced in a drawn circular galvanized pipe of diameter 152 mm and length 50 m when compressed air at the mass flow rate 3.7 kg/s, gauge pressure of 448 kPa and temperature of 20°C is flowing through it. The atmospheric conditions are STP. SOLUTION

From thermodynamics fundamentals we have p*v=m*R*T In this case p = 448 + 101.325 = 549.325. m = 3.7 and T = 20 + 273.16 = 293.16. Therefore, v = 3.7 * 287.1 * 293.16/(549.325 * 103) = 0.556905 m3/s.

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Since the diameter of the pipe is 152 mm, the cross sectional area of the pipe A will be A = π * 1522/(4 * 106) = 18145.8 * 10−6 m2 and the velocity of the flow W will be W = v/A = 0.55690/(18145.8 * 10–6) = 31.24195 m/s As p * v = m * R * T, p = ρ * R * T or ρ = p/(R * T). In this case, ρ = 549.325 * 103/(287.1 * 293.16) = 6.52667 kg/m3 The dynamic viscosity of air for 20°C temperature and 448 kPa pressure, is 1.825 * 10−5. Reynold’s number, therefore, will be Re = 0.152 * 31.24 * 6.52667/(1.825 * 10−6) = 1698264. As the pipe is a galvanized pipe the absolute roughness will be 0.00015 m and relative roughness will be 0.00015/0.152 = 0.00098684. From Moody’s chart shown in Figure 8.5 we get the friction factor as 0.0195. With all the data evaluated above the pressure loss works out to Δp = 0.0195 * (50/0.152) * (6.52667 * 31.242/2) i.e. Δp = 8172 Pa. 8.3.1.1

Simplified formulae

As stated earlier, many explicit equations have been set forth for determining the pressure loss in the flow of compressed air in circular pipes. Popular amongst them are: 1 2

Harris Formula Engineering Toolbox Formula

These formulae are semi-empirical in nature. They are not derived on a purely theoretical basis but contain some part that is based on experimental findings. None of these contain any reference to the absolute or relative roughness of the pipe. Both these formulae involve simple calculations of some readily measurable parameters. 8.3.1.1.1

Harris formula

The original Harris formula is in imperial units as given below. Δp = 0.1025 * Q2 * L/(d5.31 * p)

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where Δp = Pressure loss in kg/cm2 Q = Flow volume in ft3/s L = Length of pipe in ft d = Inner diameter of pipe in inches p = Compression ratio The following is an SI conversion of the formula. Δp = 8.333307 * 1013 * Q2 * L/(d5.31 * p) where Δp = Pressure loss in Pa Q = Flow volume in m3/min L = Length of pipe in m d = Inner diameter of pipe in mm p = Compression ratio 8.3.1.1.2

Engineering toolbox formula

Δp = 728 * Q1.85 * L * 1012/(d5 * p) where Δp = Pressure loss in Pa Q = Flow volume in m3/min L = Length of pipe in m d = Inner diameter of pipe in mm p = Initial gauge pressure of air in Pa It can be easily seen that the above formula does not take into consideration the temperature of the flowing air. The above equation is based on the absolute roughness for standard seamless carbon steel pipes of schedule 40 or schedule 80. More information and the air pressure loss values for steel pipes are given in Appendix 4 at the end of this book.

8.3.2

Compressed air flow in hose pipes

A pipe is meant to carry fluid flow from one point to the other. If the distance between these two points and their position remains unchanged it is highly beneficial to use steel pipe for reasons of cost, safety, long life etc. However, if the distance between the points and/or their position with respect to each other changes, the only alternative is to use a flexible hose pipe. In past two decades several new materials have been used in the manufacture of hoses. Due to their superior characteristics such as durability, flexibility etc., they are often preferred to the rubber hose pipes used earlier.

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Most of the modern generation of hose pipes that are used for carrying fluids at high pressure are braided, either with nylon strings or steel wires. This braiding prohibits their enlargement in diameter and allows them to carry fluids at high pressures without losing their flexibility. On a rotary blasthole drill it becomes essential to use hose pipes at two distinct places as follows: 1

2

Where the flow conduits have to cross from the main frame to the mast, which has to change its position with respect to the frame. This section is not very critical because the length of hose required for this crossover is usually rather short. Besides this, the mast is not raised or lowered frequently hence hose bending takes place only rarely. Where the flow conduit on the mast has to go to, and remain attached to, the moving drill head. This section is very critical because the distance between the points changes by a very great length. Therefore hose pipes have to be very long. Further, the relative position of the points with respect to each other changes every fraction of a second during most of the time when the drilling is carried out.

On some modern blasthole drills a hose tray as shown in Figure 8.6 is provided to carry hydraulic hoses, electric cables and air hoses from the mast to the drill head. This device greatly reduces the maintenance and ensures longer life to all the air hoses, hydraulic hoses and electric cables carried through it. Every hose has some straight portion and some curved portion. No separate fitting is involved in a hose except the end connectors. For a straight portion of the hose, pressure loss can be calculated with the same equations given in the last section for steel pipes but by treating the pipe as having very low roughness. Appropriate values of absolute roughness to be used are to be obtained from the manufacturers of the hose pipe. Since hose pipes are made from plastic or rubber with a special smooth inner lining, absolute roughness of 0.0015 mm can also be used. Appendix 5 at the end of this book gives pressure loss experienced when compressed air flows through hose pipes.

8.3.3

Compressed air flow in pipe fittings

Whenever a pipe fitting is built into a pipe line it creates an obstruction in the flow path. This results in additional pressure loss in the pipe lines. Two identical pipe lines are shown in Figure 8.7. The upper pipeline has a valve, whereas, the lower pipeline is without a valve. Other parameters of the two pipelines are same. Since the part of the pipelines on the left hand side of the plane AA is identical, the average pressure P1 across the pipe cross section in the plane AA is the same in either case. However, the pressure at the pipe cross sections in the plane BB differs in the two cases. At this plane the pressure P2 in the pipeline without a valve is higher than the pressure P3 in the pipeline with a valve.

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Figure 8.6 Hose tray to carry hose pipes from mast to rotary head.

A

B

Pressure P3 Pressure P1 Pressure P2

A

B

Figure 8.7 Pipe line with and without pipe fitting.

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This additional pressure loss (P2 – P3) is solely due to the presence of the valve. Let this difference be BF. It can be shown that BF = KL * (ρ * W2/2) where, BF = Additional pressure loss due to valve KL = Resistance coefficient ρ = Density of flowing fluid W = Average flow velocity Fittings in pipeline are generally divided into four main categories on the basis of their purpose as follows: 1 2 3 4

To change the direction of flow To control the flowing volume To connect pipelines of different diameters To divide the flow

Due to the complexity of flow, theoretical calculations of the resistance coefficient are virtually impossible. Hence, the values of resistance coefficient for different types of pipe fittings are expressed numerically. If a pipeline has length L and diameter D, and contains three fittings having resistance coefficients of KL1, KL2, KL3 then the total pressure loss in the pipeline can be equated as, ΔP = (f * (L/D) + KL1 + KL2 + KL3) * ρ * W2/2 In many cases, particularly in the case of valves, the resistance coefficient is expressed as the equivalent length Le. Equivalent length means the additional length of the pipe that would have given the same additional pressure loss occurring due to the presence of the fitting. In other words, KL1 = f * (Le1/D) In such cases the pressure loss equation will become ΔP = (f * (L/D) + f * (Le1/D) + f * (Le2/D) + f * (Le3/D)) * ρ * W2/2 Appendix 6 also gives the values of equivalent lengths for different types of valves.

8.3.4

Effect of ups and downs in the flow path

In addition to the horizontal part, a flow path often contains sections that are not horizontal. They may be inclined downward or upward or can even be vertical in upward or downward directions. This is very much so in a blasthole drill. An inclined pipe is shown in Figure 8.8. The fluid flow in this pipe is in an upward direction as shown by the arrow. It can be proved from fundamentals that if ΔPo is the pressure loss in the fluid over the length L then because of the rise in the pipeline the pressure loss is equated as

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ΔPo = ΔP + ρ * μ * L * sinθ For a flow path inclined downwards, the angle θ will be negative, therefore the equation works out to ΔPo = ΔP − ρ * μ * L * sinθ If the pipes are in a vertical direction and the flow is in an upward or downward direction, then the pressure losses will be ΔP + ρ * μ * L and ΔP – ρ * μ * L respectively. The above considerations of ups and downs in the flow path are very important in the case of flow of liquids but are negligible in the case of air flow because the density and viscosity of air are of very small values. In blasthole drills the pressure loss due to ups and downs in the flow path can be safely neglected because the length is only of the order of 30 m or so.

8.3.5

Effect of curvature in the flow path

Pipe fittings, like elbows, bends etc., used for changing the direction of the air flow, have fixed dimensions. The dimensions of their curvature are invariably dependent upon their diameter. In such case the values for additional pressure loss due to their presence is calculated by use of a resistance coefficient as stated earlier. However, when a hose pipe is used as a conduit for air flow, the dimensions of the curvature are not fixed and bear no relation to the hose diameter. Therefore, the pressure loss has to be evaluated on an individual basis. The resistance coefficient of the curved hose, denoted by ζ is very similar to the resistance coefficient of pipe fittings. Additional pressure loss incurred by the presence of curvature in the hose is given by BH = ζ * (ρ * W2/2) where, BH = Additional pressure loss due to curvature of the hose ζ = Resistance coefficient ρ = Density of flowing fluid W = Average flow velocity The resistance coefficient is dependent upon the hose diameter and radius of hose curvature.

L H Point B

θ

Figure 8.8 Forces acting on fluid in inclined pipe.

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Referring to the dimensions of the hose in a curved state as shown in Figure 8.9, the values of the resistance coefficient ζ are dependent upon the diameter of hose, the radius of curvature of the hose path and the angle through which the curved path passes. Table 8.3 gives values of the resistance coefficient for different values of the ratio R/d and different angle of curvature. It is to be clearly understood that the resistance coefficients are for a circular curved path of the pipe in a horizontally-curved alignment. Effect of height is to be treated separately.

8.3.6

Compressed air flow through nozzles

Compressed air always flows through an opening of any shape from a high air pressure container to a low pressure container. An orifice is simply a hole, made in the wall of a high pressure container that allows fluid to flow. It is shown in the right hand side drawing in Figure 8.10. A nozzle, as shown in the left hand drawing in Figure 8.10, is also an opening and allows fluid to flow, but as can be seen in the drawing, it is specially designed to streamline the flow. Therefore, it allows the fluid to flow at high efficiency. Considering the open atmosphere as the low pressure container, a schematic of the conditions is shown in Figure 8.10. An equation that correlates various influencing factors is as follows: M = α * ψ * p1 * A * [2/(R * T1)]0.5

L

d

θ R

Center of Hose Curvature

Figure 8.9 Forces acting on fluid in curved pipe.

Table 8.3 Values of resistance coefficient ζ applicable for curved hose pipes.

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Different values of the angle of curvature

For different values of the ratio R/d 1

2

4

6

10

30° 60° 90° 120° 150° 180°

0.1440 0.2520 0.3600 0.4500 0.5400 0.6120

0.0880 0.1540 0.2200 0.2750 0.3300 0.3740

0.0680 0.1190 0.1700 0.2125 0.2550 0.2890

0.0600 0.1050 0.1500 0.1875 0.2250 0.2550

0.0520 0.0910 0.1300 0.1625 0.1950 0.2210

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p1T1

pnTn

p1T1

pnTn p2T2

p2T2 D should D be more D should

than 2.5 dn dn

be more

dn

than 2.5 dn

Figure 8.10 Nozzle and orifice.

where M = Mass flow rate in kg/s α = Nozzle efficiency ψ = Flow coefficient p1 = Upstream absolute pressure in Pa T1 = Upstream absolute temperature in K A = Area of nozzle opening in m2 R = Gas constant in J/(kg · K) Some basic points to be understood clearly while using this equation are given below. The shape of the opening influences the flow rate to a very great extent. In practice, almost all man-made openings are of circular cross section. Whatever the type of opening it may be, there is always friction between the molecules of air and the walls or edges of the flow conduit. Due to this, resistance is caused to the flow and the actual flow decreases. Very well-shaped and highly polished circular nozzles can have an opening efficiency α as high as 0.97, whereas well-shaped circular orifices can have opening efficiency α of 0.61 at best. Flow volume also depends upon the ratio of downstream absolute pressure p2 to upstream absolute pressure p1. If the downstream pressure p2 is reduced, the pressure ratio p2/p1 also reduces and the flow through the nozzle keeps on increasing till the ratio reduces to a certain value. At this value of the pressure ratio, the flow is at maximum because for any lower value of the ratio the flow remains the same as that of the maximum. This pressure ratio p2/p1 is called the critical pressure ratio. Thermodynamically it can be shown that in the case of compressed air this critical pressure ratio is 0.5282, and for this value the flow coefficient ψ is equal to 0.4841. For air, the gas constant R is equal to 287.1 J/kg · K. In this equation the flow rate is given in terms of the mass and not volume of the compressed air, because the mass of flowing compressed air remains constant while the volume of compressed air changes at every point in the flow path. Here the mass of the flow is with respect to a pressure of p1. As the mass flow is in terms of kg/s, to convert it into conventional volume flow in standard conditions used for air compressors i.e. m3/min, we must multiply by 60 and divide by 1.225. This multiplier works out to 48.97959.

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EXAMPLE

From a high pressure vessel containing compressed air at 3 MPa gauge the air flows through a well made circular nozzle of diameter 5 mm to a low pressure vessel having air pressure of 600 kPa gauge. The temperature of the air in the high pressure vessel is 35°C. What will be the mass flow rate? SOLUTION

In this case the pressure ratio is (600 + 101.325)/(3000 + 101.325) = 0.2336. Therefore, the flow will remain maximum and is to be calculated by using the flow coefficient ψ = 0.4841 Since the nozzle is well-made we can assume the nozzle efficiency to be 0.95. We also have R = 287.1 J/kg · K. p1 = 3000000 + 101325 = 3101325 Pa T1 = 35 + 273.16 = 308.16°K A = π * (5/1000)2/4 = 0.00001963 m2 Hence M = 0.95 * 0.00001963 * 0.4841 * 3101325 * [2/(287.1 * 308.16)]0.5 or M = 0.133151 kg/s or Q = 6.5217 m3/min.

8.3.7

Leakage of compressed air

Leakage of compressed air from the system is to be Avoided, as leakage of compressed air means waste of energy that has been used in its compression. Pipes used for the flow of compressed air are coupled either by welded or by threaded connections. A carefully welded connection does not have any leakage. Good threaded connections can have some leakage but being small it is treated as no leakage. How much compressed air is lost in leakage through leak paths depends upon many factors such as size and shape of the path, atmospheric pressure, compressor pressure etc. Table 8.4 gives the estimated volume of compressed air leaked in NTP atmospheric conditions through apertures of various sizes. The table also gives energy wasted in the leakage. Calculations are based on compressed air pressure of 700 kPa and yearly operation of the compressor for 2000 hrs.

8.4

COMPRESSORS USED ON BLASTHOLE DRILLS

To a rotary blasthole drill, a compressor is as good as a heart to a human being. Rotary blasthole drills use compressors with discharge capacities ranging between 12.5 to 107.6 m3/min and pressures ranging between 345 to 862 kPa. In some blasthole drills, compressors are used capable of delivering air at high

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Table 8.4 Air leakage and energy wastage in air system. Aperture diameter in mm

Quantity of air lost in leaks in liters/s

Annual energy wasted in kWh

0.4 0.8 1.6 3.2 6.4 12.7

0.2 0.8 3.2 12.8 51.2 204.8

133 532 2128 8512 34040 136192

pressures – even to 3500 kPa – so that the drills are also capable of DTH hammer drilling. In the early days, rotary blasthole drills used to have piston type reciprocating compressors. By the 1960s many of the drills changed over to sliding vane type rotary compressors. From 1980 onwards preference has shifted to rotary screw compressors for the many advantages that they offer.

8.4.1

Sliding vane rotary compressors

The working of sliding vane rotary compressors is quite easy to understand. A schematic illustration of a sliding vane rotary compressor is in Figure 8.11. The working of a vane compressor can be explained from Figure 8.12. The vanes of the compressor are made from a tough polymer and are free to slide in and out from the slots in the rotor. The rotor rotates in the direction shown and the centrifugal force keeps the vanes out of the slots to the extent made possible by the steel housing in which the rotor and vanes are confined. The rotor is mounted in the housing eccentrically in such a manner that at the lower end the gap between the rotor and the housing is almost nil. Usually there are 8 vanes in 8 slots of the rotor. During rotation when the vane is at point A, it creates an enclosure of air between the rotor, housing and the previous vane that has moved to the point B. This enclosure opens in the suction port as shown and the air rushes into the enclosure. As the rotation continues both the vanes move outward and the volume of the enclosure increases. More air is sucked into it from the suction. When the vane moves from point A to point C the enclosure loses the contact with the suction port and the air in the enclosure is trapped. Further rotation forces the vane to slide inside and the volume of the enclosure keeps on reducing. The entrapped air in the enclosure is thus compressed. When the vane originally at point B reaches point D the enclosure gets opened to the discharge port and the compressed air moves into the discharge port. The volume of the enclosure keeps on reducing till it becomes virtually nil. Thus, compressed air is completely forced into the discharge port.

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Rotor

Vanes

249

Cooling Medium

Air Inlet

Air Outlet

Steel Body

Cooling Medium

Figure 8.11 Schematic of sliding vane rotary compressor.

C

Suction Port B A

Discharge Port

Figure 8.12 Working of sliding vane rotary compressor.

Some advantages of sliding vane compressor are: 1 High volumetric efficiency, up to 99.8% is retained through the life of the compressor. 2 Very compact construction, require less space.

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3 Can be driven by V belts. 4 The component alignment is only at one place and that too is not very critical. 5 Lubrication is required at three places viz. the contact of slots and vanes, vanes and housing, and the bearings of the rotor shaft. This can be accomplished by force feeding the lubricant at appropriate points. 6 Lubricant mixes with compressed air only to a negligible extent. 7 Very simple construction through usually 8 vanes and two roller bearings. Therefore fewer wearing parts. 8 Lower manufacturing cost and hence relatively low price. 9 No cold weather starting problems. 10 Low noise level. 11 Negligible pulsation of air in discharge side. 12 Low heat generation as the vanes are made from non metallic materials. The main drawbacks of the sliding vane rotary compressors are: 1

2 3

Due to the limitations of the strength of vanes made from polymer, the eccentricity of the rotor cannot be increased beyond a certain limit. Therefore, vane compressors cannot achieve a very high compression ratio, and hence it is difficult to achieve pressures beyond about 400 kPa in the first stage. It is usually not possible to change the pressure discharge characteristics. Compressed air becomes hot as it absorbs the heat generated by the friction between the vanes and the housing.

Due to these disadvantages, most of the blasthole drills that were equipped with vane compressors have now switched over to screw compressors.

8.4.2

Rotary screw compressors

The working of rotary screw compressors can be explained with a series of figures. A screw compressor consists of one male and one female rotor as shown in Figure 8.13. The male rotor has four lobes, whereas the female rotor has six grooves. The male rotor is driven in the direction as shown and the female rotor follows it in the opposite direction. When the groove of the female rotor is opened to the suction port, atmospheric air fills the groove enclosure formed by the casing of the rotors and the end flanges. As the male rotor rotates and the female rotor follows, the lobe of the male rotor intrudes into the female rotor groove at the front flange end. The male and female rotors are so machined that the gap between male and female rotors is very small. Further, it is filled with lubricant circulated through the compressor housing for the purpose. Thus, air cannot escape through this gap. As the male rotor rotates further and female rotor follows, the contact between suction port and the female rotor groove is lost and the air in the female rotor is entrapped. Further rotation causes the reduction of volume of the groove between female rotor, male rotor and the rear end of the flange as shown in Figure 8.14. Air is thus compressed further as shown by the dark gray color in the figure. Further rotation causes contact between the groove filled with compressed air and the discharge

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Atmospheric Air Female Rotor in the Female Groove

Gap Between Rotors

Front End Flange

251

Suction Port Rear End Flange

Male Rotor

Figure 8.13 Screw compressor working 1.

port at the lower side of the compressor casing as shown in Figure 8.15. The compressed air thus rushes completely into the discharge port. It is quite easy to visualize that the compression ratio depends upon the length of the male and female rotors. The length of these rotors cannot be increased beyond a certain limit that is imposed for many reasons such as difficulties in manufacture, alignment etc., but in actual practice it can be made sufficiently long to give pressures of the order of about 1500 kPa in one stage. The compressed air received in the discharge port is always mixed with oil. Rotor profiles can be machined very precisely so the need for a film of oil between the male and female rotors is eliminated. Most of the screw compressors used on rotary blasthole drills are large oil-flooded type. The proportion of oil in the air is quite excessive even for air flushing operations. For this reason and for the purpose of economy of operations the oil must be separated and retrieved. Oil separation of the compressed air is carried out in a subassembly called an oil separator and is shown in Figure 8.16. The oil-mixed air discharged by the compressor flows to the oil separator where the air is filtered to remove the oil. Usually 99.5 to 99.8% of the oil is removed from the compressed air and recirculated back to the compressor as shown in the figure. Oil that flows with compressed air to the drilling string and eventually the bit, serves the purpose of bit lubrication. When blasthole drills, earlier equipped with vane compressors, were modified with a screw compressor a significant increase in bit life was observed. Compressed air discharged from the compressor loses pressure while flowing through the piping up to the oil separator and subsequent oil filtering operation. Depending upon the design of the oil separator assembly, the pressure loss is about 60 to 100 kPa. Thus the pressure developed at the compressor is higher by this differential than the air discharged to the outlet for flow to the drill string.

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Compressed Air


Suction Port at Top

Compressed Air Discharge Port at Bottom


The main advantages of screw compressors are: 1 2 3

4 5 6

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The moving parts are only the two rotors, two roller bearings and two thrust bearings. Screw compressors can attain a very high compression ratio – even up to 20:1. Thus, it is possible to get high pressures up to even 1500 kPa in the first stage. Since oil is injected in the compressed air for the purpose of sealing and lubrication, it also reduces the temperature of compressed air sent into the discharge pipes. The size of the compressor is very compact. During operation the noise and vibration levels are very low. The pressure discharge characteristics can be easily varied by means of a sliding valve control placed at the bottom of the compressor.

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Radiator Oil Cooler

Minimum Pressure Orifice

Fan Motor

Scavenging Line Air Outlet

Inlet Regulating Valve

Air Inlet

Strainer

Oil Separator Strainer

Heavy Duty Inlet Air Filter Sight Glass Orifice Oil Filter

Thermostatic Flow Control Valve Screw Compressor

Oil Level Gage Air/Oil Receiver

AIR FLOW DIAGRAM

Oil Pump Main Oil Filter

Bearing Oil Filter

Air Air/Oil Oil

Figure 8.16 Oil separation in screw compressors.

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254

7

8 9


In screw compressors it is also possible to reduce capacity without making any change in the discharge pressure by means of a sliding valve which causes recirculation of the air between adjacent female grooves while the air is being compressed. This decreases power requirements significantly. Maintenance of the screw compressors is very easy. It is required only to a negligible extent. Pulsation levels of compressed air are very low. The main drawbacks of screw compressors are:

1 2 3

Since the compressed air contains a significant quantity of oil, an oil separator is essential. As the temperature of oil is increased during air compression, the screw compressors require an oil cooler. Precise profiling of the male and female rotors increases manufacturing costs and so the price of screw compressors is higher than equivalent sliding vane compressors.

In the early days, screw compressors were designed for operating speeds of for 3550 with two-pole electric motors, but later their design was changed for operation at about 1800 rpm. With this they could be used on blasthole drills. Today, most of the blasthole drills have screw compressors mainly for their power-saving features. Screw compressors on very large blasthole drills operate at about 450 kPa air pressure whereas those on smaller drills operate at higher pressures ranging from 689 to even 758 kPa. Some two-stage screw compressors can also operate at either low pressure high discharge mode for rotary drilling, or high pressure low discharge mode for DTH drilling.

8.4.3

Discharge and pressure control

From the viewpoint of rotary drilling the compressor must discharge an adequate volume of compressed air at sufficiently high pressure. How much discharge and pressure is sufficient depends upon several factors related to blasthole flushing. Due to its importance this topic has been elaborated in a separate chapter. It is essential that compressed air generated by the compressor must be controlled to give the desired discharge at the desired pressure. For this purpose one of the three types of control systems viz. load/no load, modulation, or combination of load/no load and modulation, are used. 8.4.3.1

Load/no load control

This system consists of a valve at the suction of the compressor (Valve1), a pressure sensor that senses the pressure at the discharge port of the compressor and another valve at the discharge port of the oil separator (Valve2).

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When the compressor drive motor or diesel engine is started initially, Valve1 is closed and Valve2 is open. As Valve1 is closed no air is sucked into the compressor and the compressor starts without any load. Thus, no overloading or overheating of the drive motor or engine occurs. As the rotors start rotating, a very small quantity of air accumulated in the housing of the compressor does get compressed and is discharged. When the pressure sensor senses this pressure, Valve1 is opened and Valve2 is closed. The compressor then starts fully loaded. Pressure in the piping and air separator sub-assembly starts increasing. When the pressure sensor senses that pressure has reached an adequately high level and the oil is being forced to the compressor, Valve2 is also opened and Valve1 is closed. Thus, the compressor works in unloaded condition. When the operator opens the valve in the discharge pipe located beyond the oil separator Valve1 is opened again and the compressor starts working in loaded condition. 8.4.3.2

Modulation control

This system consists of a control valve (CV1) at the suction of the compressor, a regulator with pressure sensor and a control valve (CV2) at the discharge port of the oil separator. The controller senses the pressure at the discharge port of the compressor and partially closes both the control valves CV1 and CV2 as needed depending upon the pressure sensed. When the motor or diesel engine driving the compressor is started initially, CV1 is fully closed and Valve2 is open. As CV1 is fully closed no air is sucked into the compressor and the compressor starts without any load. Thus, no overloading or overheating of the drive motor or engine occurs. As the rotors start rotating, a very small quantity of air accumulated in the housing of the compressor does get compressed and is discharged. When the pressure sensor senses this pressure, the CV1 is fully opened and Valve2 is closed. The compressor then starts fully loaded. Pressure in the piping and air separator sub-assembly starts increasing. When the pressure sensor senses that pressure has reached an adequately high level and the oil is being forced to the compressor, Valve2 is also opened and CV1 is closed. Thus, the compressor works in unloaded condition. When the operator opens the valve in the discharge pipe located beyond the oil separator the pressure in the system drops down. At this stage the regulator comes into action and partially opens CV1, depending upon the pressure differential. Modulation control is very effective in maintaining a stable discharge pressure. 8.4.3.3

Load/No load and modulation control

This is a combination of the two control systems described above. When the compressed air volume required to reach the pressure level is more than about 70%, the control is as per the Load/No Load System described above, but the volume requirement is less and the control is achieved by modulation. When the compressor is working in unloaded condition it consumes about 25% to 30% of the full load power. In all-electric drills the system may also consists of

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Percent of Full Load Power

120 B

100 80

E

D

A

C G

F H

60 I

40

J

20 K 0

0

20

40 60 80 100 Percent of Capacity Load Unload Control - ABCDEF Modulation Control - AGHIJK Combination Control - ABHIJK

Figure 8.17 Comparison of control system.

an adjustable timer. If this timer senses that the compressor is working in unloaded condition for a long time the electric motor driving the compressor is automatically switched off, after closing Valve1 or CV1 as the case may be. Figure 8.17 illustrates a comparison of the energy consumption with each control system.

8.5

MEASUREMENTS OF COMPRESSED AIR

As can be seen from the earlier elaboration, the calculations in respect of compressed air are rather complex and are not very accurate. For this reason many manufacturers rely more upon measurements of flow volumes, temperatures, pressures etc. than finding theoretically calculated values of such parameters. For this purpose a measuring kit, like the one shown in Figure 8.18, is used. It consists of a unit containing a pressure gage and a temperature gage, a short cylinder with a polished inner surface, precisely shaped sharp orifices of several opening diameters, and an adapter to connect the cylinder at the bottom of the drill head or drill pipe. If pressure loss is to be measured, one more pressure gage is required to be attached in the pipeline on the deck at a location where the pipeline is connected to the compressor discharge. By fitting the orifices of different sizes one after another in the kit assembled at the bottom of the drill pipe, and allowing the compressor discharge to flow through the pipes, the reading of pressure and temperature can be taken on the gages provided in the kit. The reading should be taken at such time when the readings have stabilized on the gages. This may take several minutes. At this time the reading on the additional pressure gage installed in the beginning of the flow passage is also taken.

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Figure 8.18 Compressed air measuring kit.

By using tables provided with the kit, the discharge of compressed air through the particular nozzle can be found from the measured readings on the gage. Thus, the pressure loss for a particular flow volume through the pipeline is known from the pressure difference in the two pressure gages.

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Chapter 9

Mechanics of rock fracture under a drill bit

9.1

INTRODUCTION

It is important to understand the process of rock fracture with good conceptual clarity because one can get to know the advantages, disadvantages and limitations of the equipment, and procedures used for blasthole drilling. Rock breakage takes place when energy is applied to a rock mass. The intensity and rate at which the energy is applied to the rock mass plays an important role in the rock fracture process. In drilling as well as subsequent blasting, rock is broken and moved. The mechanics of rock fracture, however, is radically different in drilling than in blasting. In this chapter both these types of rock fracture mechanisms are elaborated to the necessary extent.

9.2

ROCK FRACTURE IN DRILLING

Energy in many different forms can be applied to a rock mass for creating a blasthole. To ensure that subsequent blasting is very effective, a blasthole must have some particular dimensional properties. As a blasthole is required to have set dimensions, energy has to be applied in a very controlled form. Three forms of energy can be applied to a rock mass in controlled manner. They are 1 2 3

Mechanical Energy Heat Energy Chemical Energy

Based on these three forms of energy application, many drilling methods as shown in Figure 9.1, are recognized. Of these three forms of energy, mechanical energy is almost exclusively used in actual drilling practice. The other two forms viz. heat and chemical energy are yet to prove their potential from an economic angle. Therefore, the methods that use mechanical energy for formation disintegration are called conventional drilling methods. Methods that use other two forms are non-conventional drilling methods.

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Drilling Methods Conventional Drilling Methods Rotary Drilling

Non Conventional Drilling Methods

Cable Tool Drilling

Rotary Percussion Drilling

Top Hammer Drilling

Down the Hole Hammer Drilling

Uncased Drilling Drill Rod Drilling

Full Face Tunneling

Water Jetting

Cased Drilling

COPROD Drilling

OD Drilling

Rotary Drilling

Large Diameter Rotary Drilling

ODEX Drilling

Augur Drilling

Calyx Drilling

Diamond Drilling

Jet Piercing

Induction Drilling

Figure 9.1 Drilling methods.

Microwave Drilling

Laser Drilling

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261

About five decades ago, when the metallurgy of drilling tools had not sufficiently developed, a drilling method called jet piercing, where heat energy was used for the disintegration of the formation, was adapted for drilling blastholes in extremely hard formations such as taconites. Soon thereafter great advances were made in the concepts, metallurgy and machining of drilling tools. This resulted in a huge reduction in the cost of drilling with these tools, and now the jet piercing method is totally discarded from blasthole drilling practice. Use of chemical energy for formation disintegration is still in the experimental stage. It is very unlikely that any drilling method based on heat or chemical energy application will become economically viable in next two decades or so.

9.2.1

Modes of fragmentation

Whichever may be the type of energy used, in the drilling process rock is fragmented to create a hole. The three forms of energy that have potential in being used for drilling are described here for academic purposes. 9.2.2.1

Mechanical energy

When mechanical energy is applied to a rock mass it causes various types of stresses. These stresses cause strains in the rock mass and subsequently it breaks into small pieces as the cracks propagate. Different modes of passing on the mechanical energy to the formation are practiced. These are described below. 9.2.2.1.1

Crushing

Crushing takes place when heavy and somewhat steady force is exerted on the formation through a hard drilling bit. The strain required to cause crack propagation can be adequately high only when the stress generated by the drilling tool is high. This can be achieved by having a very low contact area between the rock and the drilling tool or by exerting a very high force on the drilling tool. Figure 9.2 shows the disintegration pattern of the rock mass caused by the exertion of steady force on it. It must be noted that just along the contact between bit and

Increased Feed Force

Feed Force

Indenter

Zone of Very Fine Cuttings

Zone of Fine Cuttings Zone of Coarse Cuttings

Zone of Coarse Cuttings

Figure 9.2 Crushing of rock under an indenter.

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the formation and within some portion beneath it, the stresses developed are compressive, but along the fracture lines the stresses are predominantly shear stresses and the fracture takes place due to shear stresses. Crushing is the basis of material disintegration in rotary drilling using tricone bits. 9.2.2.1.2

Impact crushing

When a hard material piece impinges on the rock mass, the impact force is increasing for a very short duration and then it reduces suddenly. Thereafter the contact between the tool and rock mass is lost. As the impact force is increasing, the particles of the rock mass are compressed and a compression wave is generated when the particles near the drilling tool compress the particles a little farther from the drilling tool. As the force exerted by the drilling tool suddenly reduces, the particles in contact with the drilling tool try to regain their original position by moving back. Thus, a tension wave is also generated. Rocks by nature are made up of such materials that their tensile strength is very low. Since the tensile forces in the vicinity of the drilling tool are high, cracks are generated in the vicinity of the drilling tool. Further impacts of the drilling tool propagate the cracks and eventually a piece of rock is separated from the rock mass. Impact crushing is the basis of material disintegration in percussion drilling. 9.2.2.1.3

Scratching

Scratching of rock mass by a drilling tool is schematically illustrated in Figure 9.3. For scratching to take place, a vertical force that tries to penetrate the drilling tool into the rock mass and a horizontal force that tries to move the drilling tool along the surface of the rock mass, are necessary. Cracks are generated and propagated by the horizontal force along a shear surface as shown in the figure. The vertical force is necessary to initially cause penetration of the drilling tool into the rock mass and then keep the drilling tool in the penetrated position. If the vertical force is removed the drilling tool slides up along the fractured surface and continues its movement on the top of the rock surface without generating cracks and scratches.

Vertical Force

Separated Material Horizontal Force

Surface Where Potential Cracks Exist

Figure 9.3 Scratching of rock by a plow tool.

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Scratching action for rock disintegration takes place in rotary drag bit drilling or diamond bit drilling. More details of rock fragmentation by application of mechanical energy are given later in this chapter. 9.2.2.2

Heat energy

When heat energy is transferred to the rock mass, disintegration takes place either by spalling or fusion and vaporization. 9.2.2.2.1

Spalling

A rock mass consists of varied constituent minerals. These minerals have different coefficients of linear expansion. When formations having such minerals are rapidly heated, due to different levels of thermal expansion of the mineral particles, high stresses develop in the rock mass and result in disintegration. Most of the rock-forming minerals are poor conductors of heat. If a heat source with high potential such as a high temperature flame is brought very near the rock surface, the rock particles in a layer near to the flame get heated very quickly to high temperature. Since the heat is not effectively transmitted by these hot particles to the particles situated a little deeper in the rock mass, there is a great disparity between the thermal expansion of the particles in these two (or more) layers. This gives rise to flakes. The phenomenon of flaking caused by heat is called spalling. If a rock mass consists of minerals that are better conductors of heat and have uniform coefficient of thermal expansion, it does not spall easily. Limestone is one such rock that does not spall easily. Spalling is the basis of material disintegration in jet piercing. 9.2.2.2.2

Fusion and vaporization

If a rock mass is heated to very high temperature, it melts. Further heating causes vaporization of the rock mass. In vapor form the formation can be displaced easily by the gases used for the combustion. As the vaporized minerals move away from the flames with the exhaust gas flow, the cooler surroundings turn the mineral back to solid form and the cuttings formed in the solidification come out of the hole with the exhaust gas flow. This type of heat energy transfer to formation for the purpose of drilling is as yet under laboratory testing. 9.2.2.3

Chemical energy

Some minerals that form rock easily react with certain chemicals. A very common example of this is the reaction of marble with carbonic acid. As marble is mostly composed of calcium carbonate, i.e. CaCO3, it easily reacts with carbonic acid i.e. H2CO3. The chemical equation of the reaction is CaCO3 + H2CO3 = Ca(HCO3)2

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The minerals that result from such a reaction are in loosely bonded crystal form and can be easily removed from the place of formation. Many such reactions of different chemicals with a variety of minerals are known. It is, however, not very likely that chemical energy will be used for primary rock fracture in drilling for the following two main reasons. 1 2

The chemical reactions are invariably slow The composition of rock mass varies and even a small layer of mineral that does not react with the chemical can halt further progress.

The subject of this book being rotary blasthole drilling, all further discussion about rock fracture in this chapter is in respect of crushing and scratching as these mechanisms fracture rock in rotary drilling.

9.3

BASIC THEORY OF SOLID FRACTURE

When external or internal forces act on a solid body, stresses are generated at virtually every imaginable surface within the solid body. There are three basic types of stresses viz. compressive stress, tensile stress and shear stress. Figure 9.4 shows the three types of stresses. If a surface under stress is considered, compressive stress tries to push the atoms in the solid body towards each other in the direction perpendicular to the surface. Tensile stress tries to pull the atoms in the solid body away from each other in the direction perpendicular to the surface. Shear stress tries to move the atoms in the solid body away from each other along the surface. Whenever stresses are generated in a body it changes its shape. This is called strain. Solid bodies can be perfectly elastic, perfectly plastic or elasto-plastic i.e. combination of the two types. In a perfectly elastic body strain goes on increasing with increasing stress but when the external or internal forces that cause the stress are

External Force External Force

External Force

External Force Compressive Stress

Tensile Stress

Shear Stress

Figure 9.4 Three basic types of stresses.

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reduced to zero magnitude the strain also proportionally reduces to zero magnitude. In perfectly plastic bodies the strain goes on increasing so long as the body is under stress. Later when the external or internal forces that cause strain are reduced to zero the increase in strain stops but the magnitude of strain does not become zero as in elastic bodies. In other words the body remains in deformed shape. In reality, all materials are elasto-plastic. Atoms in a solid body are bound together by bonding forces. When the atoms of a material move away from each other and the bond between adjacent atoms breaks down, fracture is said to have taken place. Fracture generates cracks. When cracks propagate and reach the surface of the body, the body disintegrates into separate small pieces. Since the bonding forces between atoms must be broken for fracture to take place, in any fracture it is essential that the atoms move away from each other. This happens only in case of tensile strain or shear strain. Thus, a solid body can fail by tensile or shear strain but not by compressive strain where the atoms move near each other and bonds do not break. Experiments carried out by Bridgman and Crossland have proved that if a solid body is subjected to external compressive forces perpendicular to it in all directions, it does not disintegrate even when the forces are extremely high. Some properties of the material, however, change dramatically. When a sample of material subjected to compressive load fails, the failure occurs through the tensile and shear stresses generated in the sample as a consequence of the compressive load. Many theories about fracture of solids have been suggested so far. For brittle materials like rocks, the Mohr-Coulomb Theory for failure in shear and the Griffith Theory for failure in tension are very widely accepted. A complete description of both the theories is beyond the scope of this book. The following are brief summaries of the two failure theories.

9.3.1

Mohr-Coulomb theory of shear failure

While studying the magnitude of pressure exerted on a retaining wall by the soil behind the wall, Coulomb proposed in 1773 that a shear stress is always associated with a compressive stress. The equation for the shear stress proposed by him is S = C + σ tan φ where S = Shear stress C = Cohesion between the soil particles σ = Normal (compressive) stress φ = Angle of internal friction A graphical representation of the above equation is shown in Figure 9.5. As Coulomb was dealing with earth pressures he concluded that failure will occur when the shear stress exceeds shear strength. Thus the line AB in Figure 9.5 is the envelope that determines failure criteria.

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B S σ tan φ A C

φ σ

Figure 9.5 Mohr Coulomb failure criteria.

Compressive Stress mechanically applied on the two circular surface of a cylindrical rock sample.

σr

Compressive Pressure i.e. stress hydraulically applied from all the sides on the curved surface of a cylindrical rock sample

Figure 9.6 Triaxial compression test.

Mohr considered the concept further and in 1882 concluded that failure within a solid body was not caused by normal compressive stress alone reaching a certain maximum or yield point, or by shear stress alone reaching a maximum, but a combination of both shear and normal stress. Even though the failure is essentially caused by shear stress, the critical shear stress is governed by the normal stress acting on the potential surface of failure. The critical combination of shear and normal stress, when plotted on a graph will give a line known as Mohr’s envelope of failure. For most rocks the line is a curve. There no equation for plotting the line. The line is actually drawn on the basis of results on the rock samples obtained in a triaxial compression test. Forces applied on the cylindrical rock sample in a triaxial compression test are shown in Figure 9.6. During the test the hydraulically applied pressure is kept constant but the stress due to mechanical force applied on the end of the cylindrical rock specimen is slowly increased till the failure takes place. Typical results of three tests carried out on three cylindrical samples of the same rock with different hydraulic pressures are given in Table 9.1. These results can be plotted to give Mohr’s envelope as shown in Figure 9.7.

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Table 9.1 Results of triaxial compression test. Mechanically applied comp. stress σc in MPa

Hydraulic pressure σr i.e. comp. stress in MPa

Difference (σc − σr) in MPa

500 900 1400

100 300 600

400 600 800

Mohr’s Envelope

The diameters of the circles are equal to the difference between the stresses. The centers of the circles are at a distance equal to the sum of stresses divided by 2.

300 600 1000

Figure 9.7 Mohr’s envelope plotted from test results.

The Mohr Coulomb theory was originally developed for soils, which by nature have very weak bonding forces amongst the particles. It can, therefore, be applied with reasonably good accuracy for slope failures in soils and rocks. The theory, however, does not go to the desired depth in the actual mechanism of fracture of the rock.

9.3.2

Griffith theory of tensile failure

This theory was proposed by Griffith in 1920. It was developed for the actual mechanism of fracture in glass which is a highly brittle material. Griffith considered that brittle materials do have very fine randomly located and oriented cracks and discontinuities within them. For the material to break these cracks must propagate i.e. extend in their length till they join each other and create a failure surface that extends up the outer surface of the material body. The parts of the body on the two sides of failure surface must also separate. The assumption made by Griffith is very near to reality because if we consider a perfect material having a lattice of atoms in the atomic structure as shown in Figure 9.8, the formula for theoretical strength of the material is σt = (Eγ/b)0.5

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σ

σ b

Figure 9.8 Atoms and bonds in an atomic lattice of material.

where σt E γ b

= = = =

Theoretical Strength of the Material Modulus of Elasticity of the Material Actual Inter-atomic Separation Inter-atomic Distance at Stable State

For quite a few materials the inter-atomic distance γ required for breaking the bonds is equal to Eb/40, hence the theoretical strength of the material works out to E/6.3245. Since E is very high the theoretical strength also works out to an unrealistically high value. Thus, the actual low strength of the material can only be explained by presuming cracks in the material. Griffith’s theory has been extended and improved by many, but the basic assumption of the existence of cracks and imperfections in materials has been recognized in each of such extensions. Therefore before going ahead with Griffith’s theory one must understand the characteristics of a crack. If we presume a circular hole in a perfect material that is subjected to tensile stress σ as shown in Figure 9.9, it can be shown from the theory of elasticity that a stress −σ (negative sign shows that actually the stress is compressive) exists at the top and bottom of the hole, and a tensile stress of magnitude 3σ exists at the sides of the hole. The stress concentration arises from the lack of load-bearing capacity of the hole. The magnitude of the stress concentration is determined by the geometry of the hole but is irrespective of the size of the hole. If the hole is of elliptical shape with dimensions b and c as shown in Figure 9.10 then the stress concentration factor x is given by an equation, x = (1 + 2c/b) and the tensile stress at the narrow ends of the elliptical crack will be σe = σ(1 + 2c/b) Griffith considered the energy balance during propagation of cracks. The left side illustration in Figure 9.11 shows an elastic body that contains a crack of length 2c.

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σ

−σ 3σ

Figure 9.9 Concentration of stress around a circular hole.

σ

2b xσ 2c

Figure 9.10 Concentration of stress around an elliptical hole.

σ

Us = 2Cy

δc U 2C U

−Ue

Figure 9.11 Energy balance in crack propagation.

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The internal strain energy of the body is Ue. The body is loaded by a force that causes tensile stress. If the crack extends by a distance δc due to the action of external force, then work W will be done by the external force. A new surface will be generated. Energy spent in generating the new surface will be Us. The total energy of the system U can be equated as U = (Ue–W) + Us The term (Ue–W) in the above equation is often called mechanical energy. It can be shown that mechanical energy Um is related as Um = πc2σ2/E where c = Crack length σ = External force E = Young’s Modulus in the crack plane Similarly, as two surfaces will be formed, the surface energy will be Us = 4cγ If the cohesion between the incremental extension surfaces δc did not exist, the crack would grow outward without much work being done. Thus the mechanical energy would reduce. But at the same time the surface energy would increase. For a condition of equilibrium dU/dc must be 0. With this condition we get the stress σf at the time of failure, 0 = −πc2σf2/E + 4cγ σf = (2Eγ/πc)0.5

or

This is considered as the Griffith criteria of failure of a brittle body under tensile stress. The most important extensions to Griffith’s Theory are by Obrlemoff (1930), Wastergaard (1939), Irwin and Orowan (1948) and Rice and Cherepanov (1965).

9.4

FRACTURE OF ROCK IN DRILLING

Rotary blasthole drilling is carried out either by tricone bits or on fewer occasions by drag bits. In either cases the cutting edge of the teeth of the bit must penetrate the rock mass. A clear concept of penetration of teeth in rock is therefore essential.

9.4.1

Penetration of an indenter

When an indenter is forced into rock, stresses are generated in the rock. Even if the contact surface between the rock and the indenter is perfectly plane, the stress distribution in the rock is usually uneven. The nature of stress distribution depends upon the rigidity of the indenter. For a perfectly rigid indenter the stress distribution

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271

is shown in Figure 9.12. The isobar curves show the lines (or surface sections in case of three dimensions) where stresses generated are equal. As can be seen, the maximum stress is always generated near the edge of the indenter. The distribution of pressure, called contact pressure, under a perfectly rigid indenter is shown in Figure 9.13. If the rock mass has infinite strength the distribution of pressure will be as shown by the dotted line.

Vertical Force Indenter with Radius r 0

2r

r

r 1

0.7 0.55 0.5 0.4

r

2r

0.6

0.3 2r 0.2 3r 0.1 4r

Figure 9.12 Isobars created by an indenter.

Vertical Force Rigid Rock Mass

Pult

Practical Rock Mass

Ideally Rigid Formation

Figure 9.13 Contact pressure below an indenter.

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In actual practice the rock has definite ultimate strength and the maximum contact pressure is as per the continuous line shown. When an indenter is forced into rock, in the initial stage the point contacts between the indenter and the rock disintegrate the rock surrounding the points. The displaced rock particles act as a very thin distribution layer and force further disintegration of the rock. Then the particles below the indenter edge disintegrate. As the particles get loose they are unable to resist the force of penetration and the particles towards the centerline of the indenter take the load and disintegrate. Ultimately a small cone of fragmented material is formed below the indenter as shown in Figure 9.14. An indenter is not able to generate cracks in rock. It creates more fine particles of the rock mass. That is the reason why a wedge shape indenter is better from the viewpoint of formation fracture.

9.4.2

Penetration of a wedge

Considerations about penetration of a wedge in the rock are more important than that of an indenter because in actual drilling practice all the teeth of a tricone bit or even the cutting edge of a drag bit have a wedge shape. This is because a wedge has a very small contact area and more importantly an ability of exerting side pressures on the rock. With this, a wedge is capable of forming larger chips on the side and is far more effective in disintegrating the rock as compared to an indenter with a large contact area and inability of generating side pressure on the rock. When a wedge penetrates rock, in the very initial stage the contact area of the edge of a wedge disintegrates a small area underneath and a very small bulb of disintegrated rock particles is formed. This is shown in Figure 9.15 on the left side sketch of the wedge. As the wedge is forced further into the rock the sides of the wedge break the rock at their contacts. Owing to the friction between the wedge surface and the particles formed in the process, the particles have a natural tendency of moving towards the tip. The pressure inside the small bulb at the tip increases. This causes the particles in the bulb to break rock on the surroundings, and the size of the bulb increases. As the pressure in the surroundings is increased further cracks are formed at the discontinuities. This is shown by the middle sketch of the wedge in Figure 9.15. Further penetration of the wedge results in higher growth of the cracks near the surface because the rock mass on their upper side is unable to offer necessary resistance

Figure 9.14 Cone of fractured material below an indenter.

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Figure 9.15 Process of chip formation by a wedge.

to the crack growth and finally chips are formed, as shown on the right side sketch of the wedge in Figure 9.15. A wedge generates more coarse chips of the rock mass than an Indenter. When a wedge is forced to penetrate rock at slow constant speed, the resistance of the rock to penetration varies. The resistance keeps on increasing until fracture takes place; immediately after fracture, chips are formed and the resistance decreases. Upon separation of the chips the resistance starts increasing once again and keeps on increasing till the fracture takes place once again and another new set of chips is formed. This is clearly visible in the plot of resistance to wedge penetration shown in Figure 9.16. The curves presented in the figure are for wedges with different wedge angles. It can be observed from the figure that in most of the cases the depth of wedge penetration is inversely proportional to the wedge angle. How big is the volume of crushed zone and how much is the depth of cracks, depends upon the engineering properties of the rock being penetrated. In brittle rocks that have high hardness and low toughness, the volume of the crushed zone is less but the length of cracks is longer. For such rocks very high feed force is needed to initiate fracture because of their high compressive and shear strength. When a wedge penetrates rocks like basalt that have higher toughness but comparatively lower mineral hardness, the volume of the crushed zone is slightly higher but the length of cracks is somewhat shorter. Very high feed force is needed for chip formation in such rocks in view of their high toughness that gives limited extent of crack propagation. Many rocks e.g. limestone, sandstone etc., have a plastic nature. They have much lesser compressive strength. Such rocks form a large crushed zone and comparatively much shorter crack lengths. A careful observation of the experimental feed force vs. penetration curve for a wedge shown in Figure 9.16 gives clear indication that chips are formed in succession as the wedge penetrates more and more into the rock mass. The magnitude of the feed force required for formation of successive chips increases more and more. Several researchers have tried to tackle the problem analytically and have arrived at a relationship between different geometrical aspects of the wedge with depth of

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Feed Force in Thousand Pounds

274

14 13 12 11 10 9 8 7 6 5 4 3 2 1 0

Decrease to resistance shows that chips are being formed under the wedge 120° Wedge Angle Increase in resistance shows that chips are not being formed under the wedge

105°

90° 75° 60°

0

0.01 0.02 0.03 Wedge Penetration in Inches

0.04

Figure 9.16 Static feed force vs. penetration of a wedge.

wedge penetration based on some rock properties. Finite element analysis has also been introduced in establishing the trends of feed force vs penetration relationship. Peeping into such mathematical treatment of the wedge penetration has been avoided here because the solutions obtained have not yet reached a stage where the calculations give better results than the experimental procedures or empirical formulas used for predicting penetration rates in drilling.

9.4.3

Formation fracture below a drill bit

Whether a drag bit or a tricone bit is used, fracture of the formation takes place by a combination of the shear and tensile stresses. As shown in Figure 9.3, a tooth of a drag bit must initially penetrate the rock through the action of vertical force and later cause shear-cum-tensile failure by the horizontal force. In the case of drag bits the penetration of the tooth causes very little formation fracture. It is the horizontal force that mainly contributes in the formation fracture by shear. Diamond bits, though not much used in blasthole drilling, are a type of drag bit. In the case of old style diamond bits where the diamonds jut outward, failure surfaces are almost wholly formed by shear stress as shown in Figure 9.17 because the horizontal force on the diamond tends to push the formation within it. In the case of PDC bits the diamond compact is placed somewhat like a plow, as shown in Figure 9.18. Since in this case the horizontal force tries to move the formation outward, the major cause behind generation of failure surface is tensile stress. As most of the rocks are very weak in tension, PDC bits give better penetration rates than the old style diamond bits under the same operating conditions.

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Bit Matrix

Embedded Diamond Formation

Second Chip

First Chip

Figure 9.17 Formation fracture by old style diamond bit.

Bit Matrix

PDC i.e. Polycrystalline Diamond Compact

Formation

First Chip Second Chip

Figure 9.18 Formation fracture by PDC bit.

In a tricone bit the cones of the bit rotate around their respective axes as the bit rotates around the axis of the drill string. The teeth on the cone come in contact with the rock mass and later move away from it. Therefore the load exerted on a tooth increases initially and then decreases. Of the many stages in which this happens, five stages are shown in Figure 9.19. As shown in Figure 9.19a, a tooth of the bit touches and penetrates the rock surface at one stage with low feed force because most of the feed force is borne by the adjacent tooth that is in the vertical direction. The volume of crushed zone created by the tooth is very small. Due to rotation of the cone, the feed force on the tooth increases while the feed force on the earlier adjacent tooth is being released. As a result of this, penetration of the tooth increases and the volume of crushed zone also increases initially. Later shear cracks are formed in the rock as shown in Figure 9.19b. At this stage the length of the cracks in the direction of feed force, which is still slightly off the vertical, is highest. As the cone rotates further, due to lateral shift of the tooth, which in this case is presumed to be towards the right hand side of the picture, the cracks in the right hand direction propagate to a considerable distance. The cracks in the left hand direction do not propagate, or propagate to a very short distance. This can be seen in Figure 9.19c. Stresses generated in the rock due to further rotation create chips as shown in Figure 9.19d. Laterally loosened fine particles in the crushed zone are removed at this stage by flushing air. Finally due to lateral shift and rotation of the tooth the chips are separated from the rock mass and move away with flushing air as shown in Figure 9.19e.

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Feed Force Wedge

Feed Force Wedge

Feed Force Wedge Tooth Height

Crushed Zone

Rock

Rock

Rock Initial Cracks a

Tooth Penetration b

Feed Force Wedge

Propogating Cracks

c

Wedge

Rock

Rock Chip

Separated Chip d

e

Figure 9.19 Formation fracture underneath the tooth of a tricone bit.

Due to the shifting position of the tooth and consequent rock fragmentation, as explained above and shown in Figure 9.19, very high torque is needed for rotating tricone bits. Tricone bits meant for drilling in very hard formation have lesser offset. Such offset imparts only limited shift to the tooth. This is done so because, if the offset is kept large, rather than rock fracture, the teeth of the bit tend to break or wear out rapidly. For this reason the shape of teeth meant for hard rock drilling is with a larger wedge angle in the case of steel tooth bits, or nearly hemispherical in the case of TC inserted bits. In other words this means that for bits meant to drill in very hard formations, the height of the teeth is less, and for those meant to drill in soft formations the height of insert is greater. Technical common sense tells us that as long as the penetration of the tooth in the formation does not exceed the height of the tooth, the following trends are likely to occur. 1

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Since the teeth of the bit will penetrate in the formation to a larger depth, the penetration rate is likely to increase with increasing feed force.

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2

277

If the speed of rotation of the bit increases, the teeth of the bit will penetrate into the formation more frequently. Hence the penetration rate is likely to increase with increasing rotary speed.

In actual drilling with tricone bits, there are four factors that largely affect the process of drilling. Mathematically it can be expressed as P = f(W, N, C, Q) where P W N C Q

= = = = =

Penetration rate Feed force on the bit Rotary speed of the bit Height of tooth of the bit Compressed air flow volume used.

Of the above factors the volume of compressed air i.e. Q does not take part in the formation fracturing process but is very important from the viewpoint of removing the cuttings from the bottom of the hole. Therefore it is elaborated in a separate chapter. George Shiveley took into consideration a large spectrum of blasthole drilling data obtained in field practice. Drilling was carried out by tricone drill bits in limestone quarries. In all the cases the tooth penetration in the formation was less than the tooth height. Some plots and the conclusions drawn by him are described below. A plot shown in Figure 9.20 clearly indicates that for a constant rotary speed the penetration rate varies as per the feed force exerted on the tricone bit. Mathematically it can be stated that, P∝W The plot in Figure 9.21 is for penetration rates at 100 and 200 rpm rotary speed for different feed forces on the bit. The data used to draw this plot were from a coal mine. This plot also reconfirms that rotary speed remaining the same, the penetration rate varies as per the feed force exerted on the bit. More importantly the plot also indicates that by doubling the rotary speed (i.e. from 100 rpm to 200 rpm) the penetration rates also doubles. Yet another plot is shown in Figure 9.22. It represents data from a copper mine. Here it can be seen that till the penetration rate was about 2.4 ft/min (which corresponded with feed force of 27000 lb) the curve was a straight line, but when higher and higher feed force was exerted on the drill bit while keeping the same rotary speed, the penetration rate became successively lower than the one predicted by the straight line. Here, it may be noted that the bit used for drilling had a tooth height of 0.37 inches and was rotated at 79 rpm. This means that while drilling at the penetration rate of 2.4 ft/min the penetration of the tooth was equal to the tooth height.

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5 N = 100 RPM N = 155 RPM

Penetration Rate ft/min

4 N = 220 RPM N = 285 RPM 3

2

1

0 0

10

20

30

Weight on the Bit in lb/1000 Figure 9.20 Plot of penetration rate vs. feed force on the bit at various rotary speeds.

5 N = 100 RPM N = 200 RPM


4

3

2

1

0 0

10

20

30

Weight on the Bit in lb/1000 Figure 9.21 Plot of penetration rate vs. feed force on the bit at 100 and 200 RPM rotary speed.

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5 N = 79 RPM


4

3

2

1

0 0

20

40

60

80

Weight on the Bit in lb/1000 Figure 9.22 Plot of penetration rate vs. feed force on the bit at rotary speeds of 79 RPM.

A tricone bit is said to have bogged down when the teeth of the bit penetrate in the ground to the fullest extent and additional feed force on the bit does not increase the bit penetration in the formation. Theoretically, in certain cases, such increase in feed force on the bit can actually reduce the penetration rate because the process of formation fracturing gets hampered. Insufficient flushing can occur even before bogging down of a bit. In such cases the cuttings formed are not removed from the hole bottom and secondary crushing is required before they are lifted off the hole bottom. The formation fragmenting action of a tricone bit can be divided into three different phases. If the weight on the bit is very low the penetration of teeth is very small, and whatever fragmentation takes place is without forming cracks in the formation as shown in Figure 9.19a. Such fragmentation takes place mainly due to the abrasion between the teeth and the formation. At higher feed force cracks are formed, as shown in Figure 9.19b and 9.19c. With successive crack formation the material gets removed but without any lateral displacement. Such fragmentation is called fatigue fragmentation. When the bit has a large cone offset and the load applied on the bit is significantly high the formation fracturing takes place due to the laterally applied loads as illustrated by Figure 9.19d and 9.19e. This type of fragmentation is called scraping. While using appropriate tricone bits in drilling soft, medium-hard and hard formations all the three types of fragmentation modes are present but the contributory proportion changes as shown in Figure 9.23. In soft formations much of the

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280

Soft Formations

Medium Hard Formations

Hard Formations

Abrasive Fatigue Scraping Bit Bogging

Increasing Feed Force on the Bit Figure 9.23 Contributory proportions of different modes for formation fracture under tricone bits.

fragmentation is due to scraping action but as the formations drilled are harder the contribution of abrasive and fatigue type fragmentation becomes significantly higher. In tricone bits the limitations on feed force on the bit are mostly due to the bearings in the bit. With excessive feed force the balls or rollers make a dent in the cone or bit leg surface with which they are in contact. Such dents result in vibrations that accelerate the process of dent-making and the bit fails prematurely. Small diameter tricone bits have less space for accommodating the bearings and therefore, the bearings of small diameter bits are proportionally smaller than those of large diameter bits. For this reason the maximum recommended weights on small diameter bits are also less than those for large diameter bits.

9.4.4

Specific energy

As a matter of fact, drilling is an excavation process. The difference between excavation, as it is normally visualized, and drilling is that drilling takes place in a particular cross sectional area in the rock to create a hole. The specific energy of rocks can be defined as the work done in the excavation of a unit volume of rock. Naturally, specific energy is dependent upon the properties of the rock to a very great extent. In rotary blasthole drilling the excavation is carried out within the boundary of the blasthole and for this purpose energy in the form of thrust and rotation of the drill bit is applied by the blasthole drill. An equation that correlates various

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parameters of energy transfer through a drill bit in rotary drilling was set forth by Teale as follows: S = (F/A) + (2 * π/A) * (NT/P) where S F A N T P

= = = = = =

Specific energy Thrust applied on the drill bit Area of blasthole cross section Rotary speed Torque applied to the drill bit Penetration rate

As can be seen the equation has two components. The first component viz. (F/A) does not contribute to the ‘excavation’ process in a significant manner. This is because the energy used for the penetration of the teeth of the bit into the formation is spent in crushing the rock underneath the teeth and the formation of cracks. The second component viz. (2 * π/A) * (NT/P), however, causes formation fracture to a very large extent because it loosens and displaces the fragments made by the cracks. The cross sectional area of the blasthole is almost constant as it is being drilled hence the first term, (2 * π/A), is a constant. In the initial stage of drilling when the thrust on the bit is low, hardly any penetration takes place. In such circumstances the specific energy is very high. But when thrust is increased the torque required for rotating the drill bit also increases. This increases the penetration rate. Thus, the specific energy decreases. The minimum value of specific energy can be considered as an indicator of drillability. However, such correlation has to be used very cautiously.

9.5

DRILLABILITY OF ROCKS

Drillability of rock means the real or projected rate of penetration in a given rock by a particular type of drilling bit. Sometimes it is considered as a property of rock but actually it depends not only upon the type of rock and its characteristics, but also upon many other factors. Factors that affect drillability of rock are of four categories. Factors Related to the Rock Factors Related to the Blasthole Drill Factors Related to the Drill Bit Factors Related to the Rock Mass From the above description of the process of rock fracture under a drill bit, it can be easily surmised that the fracture process is very complicated. It is possible to establish a qualitative relation of drillability with many factors as has been done in Table 9.2. However, as the factors are interrelated with each other in a complex manner, it is very

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Table 9.2 Correlation of drillability with factors in drilling. Category

Factor

Dependence of drillability with factor

Rock Factors

UCS Tensile Strength Density Abrasiveness Toughness Shear Strength Particle Cohesion

Decreases with increasing UCS Decreases with increasing Tensile Strength Decreases with increasing Density Decreases with increasing Abrasiveness Decreases with increasing Toughness Decreases with increasing Shear Strength Decreases with increasing Particle Cohesion

Machine Factors

Feed Force Rotary Speed Torque Vibrations Air Flow Air Pressure

Increases with increasing Feed Force Increases with increasing Rotary Speed Increases with increasing Torque Decreases with increasing Vibrations Increases with increasing Air Flow Increases with increasing Air Pressure

Bit Factors

Teeth Height Teeth Apex Angle Number of Teeth Hardness Abrasiveness Diameter

Increases with increasing Tooth Height Decreases with increasing Tooth Apex Angle Increases with increasing Number of Teeth Increases with increasing Hardness Increases with increasing Abrasiveness Decreases with increasing Diameter

Rock Mass Factors

Joint Spacing Joint Orientation Homo Water Contents

Decreases with increasing Joint Spacing Difficult to describe qualitatively Difficult to describe qualitatively Decreases with increasing Water Contents

difficult to develop a mathematical formula that can take into account all the factors mentioned.

9.6

ESTIMATION OF PENETRATION RATE

Many terms such as penetration rate, net penetration rate, gross penetration rate, drilling rate, drill performance etc. are used in connection with the progress of blasthole drilling. Since the terminology is not standardized, often these terms are intermixed. Here in this book such terms are used with specific meanings specified in Table 9.3. Many researchers have attempted to correlate penetration rate with one or more easily measurable properties of rock. Some of these are elaborated in following sections.

9.6.1

Protodyakonov approach

One of the earliest attempts of correlating penetration rate with rock properties was by Protodyakonov, who in 1926 proposed a toughness index of rock based on a simple test.

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Table 9.3 Definitions of penetration rate category factor. Category

Factor

Penetration Rate

Measured in m/min. Calculated by dividing the length of the blasthole drilled by the time interval during which the formation fracturing activity through application of feed force and bit rotation are continuous.

Instantaneous Penetration Rate

Measured in m/min. Calculated in the same manner as penetration rate mentioned above but over a very short interval of time.

Net Penetration Rate

Same as penetration rate mentioned above.

Gross Penetration Rate

Measured in m/min. Calculated by dividing the length of complete blasthole drilled by time taken for all the activities required for reaching completion of drilling the hole. Here time spent for moving the drill to the location of the hole and positioning it is not included but time spent in adding and removing the drill rods is included. Drilling Rate Measured in m/h. Calculated by dividing the length of complete blasthole drilled by time taken for all the activities required for reaching completion of drilling the hole. Here time spent for moving the drill to the location of the hole and positioning as well as time spent in adding and removing the drill rods is included.

Drill Performance

Measured in m/shift. Calculated by dividing the total length of all the blastholes drilled in one shift by the duration of shift i.e. 8 hours. It includes time taken for all the activities required for drilling the blastholes during the shift but does not include pauses for non availability of the power etc. In this book this term is used as a cost effectiveness indicator as stated in the beginning of this chapter.

In actual test, samples of the rock measuring about 15 to 20 mm were kept in the pan of the test apparatus shown in Figure 9.24, so that they filled the pan up to a height of 40 mm. A 2.4 kg weight, was dropped on these rock specimens by removing the retaining pin so that the drop height was 600 mm. After 10 drops (or more if required) of the weight the sample was removed and sieved through a wire mesh having square openings of size 0.5 mm. Crushed material passing through sieve was collected and poured in the measuring apparatus shown in Figure 9.24. After lightly tapping the cylinder, a piston was inserted in the cylinder to measure the height of column of crushed material in the measuring cylinder. The Protodyakonov number, that represented the dynamic strength of the rock, was found by using an empirical formula as under, f = 20(n/h) where n was the number of drops of the weight and h was the height of crushed rock, measured in mm, in the cylinder. It was realized later that Protodyakonov number correlated with properties of rock as f = 1.887 * (σc2/(2 * E))

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All Dimensions in mm.

Figure 9.24 Apparatus for measuring protodyakonov index.

where σc = UCS in kg/cm2 E = Modulus of elasticity in kg/cm2 Based on the tests described above Protodyakonov proposed a classification of rocks as shown in Table 9.4. The indices presented give indications of likely penetration rates only on relative basis rather than absolute basis. Therefore, they are not of much use in actual estimation.

9.6.2

Paone and Bruce approach

An equation set forth by Paone and Bruce for penetration rates in diamond core drilling is δ = 2 * π * (T – μ * F * r)/(S * A – F) where δ T μ F r S A

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= = = = = = =

Depth of penetration per revolution of the bit in m Torque applied to drilling bit in Nm Coefficient of friction between bit and rock Feed force applied on the bit in N Mean radius of the drilling kerf in m Resistance of rock to drilling Cross sectional area under bit in m2

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Table 9.4 Rock classification based on protodyakonov indices. Category

Hardness level

I

Hardest

The hardest, toughest and most dense quartzite and basalt, other exceptionally hard rocks

20

II

Very hard

Very hard granitic rocks, quartz porphyry, very hard granite, siliceous schist, weaker quartzite, hardest sandstone and limestone

15

III

Hard

Granite (dense) and granitic rocks, very hard sandstone and limestone, quartz bearing veins, hard conglomerate, very hard iron ore

10

IIIa

Hard

Limestone (hard), weaker granites, hard sandstone, hard marble, dolomite and pyrites

8

Description of rock

Rock hardness f

IV

Rather hard

Ordinary sandstone, iron ore

6

IVa

Rather hard

Sandy shale, schist, schistose sandstone

5

V

Moderate

Hard argillaceous shale, non-hard sandstone and limestone, soft conglomerates

4

Va

Moderate

Various types of shale (non-hard), dense marl

3

VI

Rather soft

Soft shale, very soft limestone, chalk, rock-salt, gypsum, frozen soil, anthracite, ordinary marl, weathered sandstone, cemented shingle and gravel, rocky soil

2

Vla

Rather soft

Detritus soil, weathered shales, compressed shingles and rock debris, hard bituminous coal, hardened clay

1.5

VII

Soft

Clay (dense), soft bituminous coal, hard alluvium

1.0

Vlla

Soft

Soft sandy clay, loess, gravel

0.8

VIII

Earthy

Vegetable earth, peat, soft loam, damp sand

0.6

XI

Dry Substances

Sand, talus, soft gravel, piled up earth, extracted coal

0.5

X

Flowing

Shifting sands, swampy soil, saturated loess and other saturated soils

0.3

This equation applies to diamond core bits that have small and uniform size diamonds studded in a sintered matrix. While calculating the depth of penetration per revolution by using above equation, all the other parameters, except S, are easily known from the bit geometry and machine characteristics. Figure 9.25 shows the experimental and theoretical results for penetration rates for diamond core drilling. In theoretical calculations the coefficient of friction for the rock, i.e. μ, was presumed to be 0.4 and resistance to rock drilling i.e. S, was presumed to be equal to the compressive strength of rock. In tests the three quantities viz. penetration rate, coefficient of friction and compressive strength were measured. Figure 9.25 indicates very good correlation between experimental and theoretical results.

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0.060

Penetration Per Revolution in inches

0.055 0.050 0.045

Experimental Curve Drawn From Test Results

0.040 0.035 0.030 0.025

Theoretical Curve Drawn On the Basis of Equation

0.020 0.015 0.010 0.005 0.000

0

10

20

30

40

50

60

70

Compressive Strength in 103 psi Figure 9.25 Penetration rate for diamond drilling.

For many reasons, explained elsewhere in this book, rotary drilling with diamond bits was not practiced in the blasthole drilling industry. But soon, when bits with new type of diamonds will be available, diamond drilling may be introduced for blasthole drilling. Fracture process in diamond bit and drag bit drilling is similar hence the equation can also be used for predicting penetration rates of drag bit drilling.

9.6.3

Bauer, Calder and Workman approach

In the early 1960s rotary blasthole drilling became well established in blasthole drilling activities. Very heavy blasthole drills were developed for drilling in extremely hard formations, met in the iron ore mines, with the tungsten carbide inserted tricone bits. Bauer and Calder studied the subject of working performance of TC inserted tricone bits in hard iron ores of Mesabi Range in USA and developed the following empirical equation in 1967 for prediction of the penetration rate. log(P/6) = K log(12W/SC) where P = Penetration rate in ft/hr

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SC = UCS in thousands of lb/in2 W = Feed force exerted on the bit in terms of lb/in of bit diameter K = A factor that depended on the rock and varied between 1.4 and 1.75 for rocks of UCS from 15000 lb/in2 to 50000 lb/in2 In 1971 Bauer modified the equation by adding one more factor i.e. rotary speed. His equation was P = (61 − 28 * log SC)(W/φ)(N/300) where P SC W N φ

= = = = =

Penetration rate in ft/hr UCS in thousands of lb/in2 Feed force exerted on the bit in terms of thousands of lb Rotary speed of the bit in RPM Diameter of the bit in inches

In hard iron ores the values of penetration rates projected by using this equation tallied well with the actual measurements taken in the field but gave erroneous results when the iron ores were of less uniaxial compressive strength. To overcome these shortcomings investigations were done by Calder and Workman. They used data obtained in blasthole drilling in formations of lower uniaxial compressive strength and proposed a modified equation in 1994 as under, P = 5.7 * 10−5 * (RF–28 * log(0.145 * SC))W * N where P RF SC W N

= = = = =

Penetration rate in m/hr Rock Penetration Factor UCS in MPa Feed force exerted on the bit in terms of kg/mm Rotary speed of the bit in RPM

The rock penetration factor RF can be found from Table 9.5. If the UCS of the formation lies somewhere in between the limits of each class appropriate value from the class should be taken from the curve shown in Figure 9.26.

Table 9.5 Rock penetration factors.

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Rock description

Uniaxial comp. strength psi

Uniaxial comp. strength MPa

Rock factor

Very Hard Hard Moderate Soft Very Soft Extremely Soft

More than 30000 15000–30000 10000–15000 5000–10000 1000–5000 Less than

More than 207 103–207 69–103 34–69 7–34 1000 Less than 7

61 180 100 135 200 300

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The rock factor RF was determined by using the penetration rate data shown in Figure 9.27. In the use of this equation over the past decade it has been found that this equation gives acceptably correct estimate of penetration rate to be expected in blasthole drilling. In all the above equations it has been presumed that the bailing velocity of the compressed air is sufficiently high and the blasthole drilling process is not hampered due to insufficient flushing.

Rock Penetration Factor

350 300 250 200 150 100 50 0 0

25

50

75 100 125 150 175 200 225 250 275

Uniaxial Compressive Strength in MPa Figure 9.26 Rock penetration Factor vs. UCS data curve used for interpolating rock factor. 100 90

Penetration Rate in m/hr

80 70 279 mm dia 311 mm dia

60 50 40

251 mm dia

381 mm dia

200 mm dia

30 20 10 0 1000

10000 Formation UCS in lbs

100000

Figure 9.27 Penetration rate in rock mass of different UCS and different blasthole diameters.

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It is also presumed that the formation in which drilling is carried out is dry and the dry dust collection method as described in chapter 7 is used for the dust suppression. EXAMPLE

Blasthole drilling is to be carried out in a rock formation having UCS of 200 MPa by using TC inserted tricone bit of 311 mm diameter with a feed force including the drill head weight of 30,000 kg at a rotary speed of 80 rpm. What penetration rate can be expected? SOLUTION

As per chart in Figure 9.26, for the rock UCS of 200 MPa the rock factor works out to 65. The feed force exerted on the bit is 30000/311 i.e. 96.46 kg/mm. These factors give P = 5.7 * 10−5 * (65 − 28 * log(0.145 * 200)) * 94.46 * 80 Evaluating the above equation gives P = 10.36 m/hr.

9.6.4

Specific energy approach

Following equation has been suggested for calculating penetration rate based on specific energy of the rock. P = 48 * R * C/(π * D2 * Es) where P R C D Es

= = = = =

Penetration rate in in/m Drill head rotary power in in.lb/min Energy transmission factor Blasthole diameter in inches Specific energy in in.lb/in3

The most common values of C i.e. energy transmission factor lie between 0.6 to 0.8. The value of Es can be determined by the following equation. Es = 13900 * CRS + 15500 where Es = Specific energy in in.lb/in3 CRS = Coefficient of rock strength (From Table 9.6.)

9.6.5

Microbit test approach

In this test a large size rock sample is taken from the worksite to a testing laboratory. A hole is drilled in the sample by a two or three cone drill bit of small size.

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Table 9.6 Coefficient of rock strength for some rocks. Rock name

CRS

Rock name

CRS

Humboldt Iron Silicate Hornblende Schist Granite Pegmatite Wausau Quartzite Wausau Argillite Mankato Stone New Ulm Quartzite Jaspar Quartzite Rockville Granite Charcoal Granite Diamond Gray Granite Dresser Basalt

2.39 1.64 0.77 0.78 2.28 0.45 0.75 1.01 0.84 1.21 0.82 2.86

Shelly Limestone Mount Iron Taconite Aurora Taconite Babbitt Taconite Sabbitt Diabase Ely Gabbro Trap Rock Anorthosite Duluth Basalt Marble Primax Gabbro Iron Ore

0.57 1.47 2.62 2.84 2.44 1.21 0.64 0.73 2.11 0.98 1.02 1.23

The penetration rate attained in such tests is carefully measured and then the results are extrapolated to the actual drilling conditions at the worksite. It is necessary to carefully prepare the rock sample for the test, and drilling parameters such as weight on the bit, rotary speed etc. have to be suitably adjusted so that the extrapolations, to be carried out later on, give correct results. The test is repeated 5 to 10 times with a new bit for each of the tests. Mean penetration rate measured in the tests is used for interpolation. Interpolation is done on the basis of a chart similar to the one shown in Figure 9.28. The chart in Figure 9.28 is presented for illustrating the method. Tricone bits, on the basis of which the chart was designed, are old types and have now been largely replaced by more efficient bits. The microbit drilling approach was usually adopted by tricone bit manufacturers since they had very precise knowledge of the physical and geometrical parameters of their bits and manufacturing a suitable scaled down version of the bit was much easier for them. At first glance the microbit drilling test appears to be better placed for giving more accurate results for penetration rates but actually it is not so because the scale factor is not very easy to establish. Further, the test is also very costly as it involves use of specially made small drill bits rather than readily available simple tools. In many cases a simple masonry drill bit can be used for drilling a hole in the sample but it is more akin to drag bit or diamond bit drilling. Manufacturers of tricone drill bits usually have a large database of the use of tricone bits in various types of rock formations drilled in several places. From such databases they have prepared graphs of tricone bit penetration rates as shown in Figure 9.29. In most cases these graphs give very reliable values of penetration rates that can be expected in the field.

9.6.6

Indenter test approach

This test is also called as stamp test because the indenter makes a stamp on the rock.

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100 90 80 70 60 50 40

A - Carthage Marble, Sandstone, Limestone, Gypsum B - Dolomite Lime, Dolomite, Hard Shale C - Granite, Chert, Quartzite

30

M Pe icro ne bi tr t at io n

Ra

te

in

Ft

/h

r

10.0 9.0 8.0 7.0 6.0 5.0 4.0 3.0

4.0

2.0

3.5 3.0

1.0 0.9 0.8 0.7 0.6 0.5 0.4

0.2

0.2

0.1 10

2

3

4

5 6 7 8 9

2

3

4

0.1 0.04

0.06 0.02

0.4

0.3

0.6

0.8

1.0

1.5

2.0

2.5

Penetration Rate in Ft/hr

20

5 6 7 8 9

100 Weight on the Bit in lb/inch of Bit Diameter 1000

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Figure 9.28 Penetration rates based on microbit test.

2

3

4

5 6 7 8 9 10000

292


100 8 6

ne 20 Sh 00 ale ps 35 i 00 ps Do i lom Ma rb ite le 89 70 00 00 psi ps i

ds to

2 Sa n

Penetration Rate in ft/hr

4

0 8 6 4

Sandstone 21000 psi 2

Quartzite 60000 psi

Basalt 46000 psi

Granite 51000 psi 1 102

2

4

6 8 103

2

4

6 8 104

Weight on the bit in lb/inch

Figure 9.29 Penetration rates based on database with a manufacturer (at 60 RPM).

This test is used for predicting drillability of rocks in rotary percussion as well as rotary drilling. Since a small tungsten carbide insert is used as an indentor rather than a tricone bit, the test can be carried out in any material testing laboratory. In their mode, penetration of an indenter into rock sample and the tooth of a tricone drill bit into in situ rock mass are identical to each other. An equation that gives the relation between several factors encountered in the process of rotary drilling is as under: V = 0.159 * K * (P/S)–1(W *N /(R * B)) where V W N K P/S R B

= = = = = = =

Penetration rate Weight on the bit Rotary speed in rpm A parameter related to rock Parameter related to penetration of indenter into the rock Radius of the load bearing tooth of the tricone bit Number of teeth on a bit in contact with the rock and bear the load

The above equation can be derived from fundamental principles. Of the above parameters K and S are determined by tests carried out in the laboratory. Parameters B and R can be known from the details of the bit provided by the manufacturer or actual physical observation and counting of the bit teeth.

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Parameters W and N are governed in actual drilling by using the controls provided on the blasthole drill. The sample to be used for the stamp test need not be a core sample but can be of any shape. The size of the sample should be such that the least dimension is more than 100 mm. If the rock is soft, larger sample sizes of the order of 200 mm are desirable. In fact a larger sample always proves ideal. Care should be taken in selecting an intact sample without cracks. For testing, the sample is embedded in concrete and the block is then cut by circular diamond saw so the sample surface is fresh and undisturbed. In the actual test a tungsten carbide stamp of circular cross section is pressed in the rock sample as shown schematically in Figure 9.30. The stamp is chosen in such a way that it has the same contact area as that of the insert on the drill bit to be used in actual drilling. Initially the stamp is in contact with the rock surface with little force. During test the force on the stamp is increased and the penetration of the indenter in the rock is continuously measured. At one juncture a crater is formed and the load on the stamp gets suddenly released. This is one cycle of the stamp test in which the stamp makes one indenture in the rock sample. The test is repeated at four or more positions on the sample. These positions are selected in such a way that the craters formed in earlier tests remain and are unaffected by the new stamp penetration. Apart from force and penetration measurements taken during the tests, the volume of material removed in each of the indentures is also measured. For this, weight of the material fragmented in the indenture is precisely measured and its volume is calculated on the basis of density of the sample. Graphs of force vs. penetration are drawn to determine the maximum stamp force for the penetration. One such graph is shown in Figure 9.31. Once the average values of stamp force FST, stamp penetration XST and crater volume VST are known, the stamp strength is calculated by the formula σST = FST/(4 * π * B2) In actual practice the test is repeated several times and the average values of FST, XST and VST are taken for calculations. Surface Carefully Prepared for the Test

Force Exerted on the Indenter

Tungsten Carbide Indenter

Rock Sample

Concrete Embedding the Rock Sample

Figure 9.30 Schematics of stamp test.

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Fst

Xst Figure 9.31 Force penetration curves from a stamp test.

The stamp strength is found to be much higher than the uniaxial compressive strength because in the stamp test the rock sample is virtually in semi infinite state. Test results show that stamp strength lies between 5 to 15 times the uniaxial compressive strength. In the case of tricone bits each of the three cones have many teeth in the form of milled points or tungsten carbide inserts. During rotation of the bit some of the inserts are in contact with the formation. All the weight on the bit is used by these teeth for penetration into the rock. In one rotation of the bit many (not necessarily all) teeth on the cones come in contact with the formation. The weight exerted on the bit has to be sufficiently high to exceed the value FST on each of the inserts in contact with the formation. Suppose the bit used for drilling has diameter D mm and the number of teeth of the bit coming in contact with the formation in one revolution is T rpm and the rotary speed is N, then the volume of material removed per revolution will be N * VST. Thus in one minute the volume removal will be T * N * VST. Since the cross sectional area of the hole is o * D2/4 the penetration rate P in terms of mm/min can be calculated by P = 4 * T * N * VST/π * D2 where P = Penetration rate in mm/min VST = Crater volume in mm3 Nowadays this test is becoming popular for the following reasons. 1 2 3 4

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It gives more reliable estimates of the penetration rates. It needs only one or two tungsten carbide indenters and a universal testing machine. It can be completed at short notice and can be carried out at virtually any material testing laboratory. Since it does not require any specially machined micro bit the test is normally far more inexpensive compared to a microbit test.

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Chapter 10

Flushing the blasthole

10.1

INTRODUCTION

A blasthole, or for that matter any type of hole in the ground, cannot be drilled without taking out the fragmented material from the hole. Flushing is a component of drilling activity, where a flowing medium is taken to the bottom of the hole by exerting pressure on it so that it returns to the surface of the ground and brings along cuttings with it. As mentioned in chapter 2 of this book, flushing is one of the essential components of the drilling process. In rotary blasthole drilling the flushing medium happens to be compressed air. In fact the power required to be spent in compressing the air to be used for flushing is even more than that required for any other essential drilling activity. For efficient blasthole drilling, the flushing has to be carried out with a well calculated plan. With the possibility of varying the volume of compressed air generated by the compressor, a driller can accomplish significant power saving if he is well acquainted with the process of flushing. This chapter is devoted to giving essential technical knowledge of the compressed air flushing process as used in rotary blasthole drilling.

10.2 WHY COMPRESSED AIR? Way back in 1946, i.e. in the beginning of rotary blasthole drilling, initial attempts were made to drill blastholes with water and mud as the flushing fluids. Later, when compressed air was used for flushing, it was noticed that compressed air cleaned the holes very efficiently and the size of cuttings ejected from the blasthole was far larger than that possible with water or mud circulation. This resulted in a far higher penetration rate. For this reason compressed air has been invariably used as a flushing medium in rotary blasthole drilling from the year 1950. The other reason for the preference for compressed air flushing was due to the fact that it gave a dry hole in which explosives could be filled directly i.e. without need for dewatering them.

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10.3


SCHEMATICS OF FLUSHING

When a blasthole is flushed with compressed air, the compressed air is generated at the compressor located on the main frame of the blasthole drill. It then goes to the drill head from where it enters the rotating drill string through the swivel of the drill head. This has been well illustrated in Figure 8.3. The passage of compressed air in a blasthole is shown in Figure 10.1. When the compressed air comes out of the nozzles of the drill bit and impinges on the formation, the cuttings formed by the interaction between the drill bit and formation get lifted from the bottom and may start moving upward, depending upon the forces acting on them. Three forces acting on the cutting are shown in the inset in the Figure 10.1. These three forces are gravitational force, buoyant force and drag force.

Compressed Air Moving Down through the Central Bore of the Drill Drill Pipe Compressed Air Moving Up through the Annular Space Between Hole Walls and Drill Pipes

Drag Force Buoyant Force

Gravitational Force

Drill Bit

Figure 10.1 Flushing schematics.

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10.3.1

297

Gravitational force

Gravitational force is due to the gravitational field of the earth. It is proportional to the mass of the particle and is constant throughout the fall of the particle. For a spherical particle it is given by Fg

= 4 * π * r3 * dm g/3

where Fg = Gravitational force r = Radius of the falling spherical particle dm = Density of the material of particle g = Gravitational constant. Gravitational force always acts in a vertically downward direction.

10.3.2

Buoyant force

Buoyant force is due to the displacement of the medium in which the particle is falling. This displacement is due to the volume of the falling body and can be written as Fb = 4 * π * r3 * df g/3 where Fb = Buoyant force r = Radius of the falling spherical particle df = Density of the material of particle g = Gravitational constant. Buoyant force always acts in a vertically upward direction.

10.3.3

Drag force

When a particle is suspended in a stream of moving fluid it is always being dragged in the direction of fluid flow. The drag force is generated due to the friction between the particle surface and the medium. Besides, the particle has also to displace the medium and make room for it to move. Both these components of the drag force are dependent upon the velocity at which the relative movement between the particle and medium takes place. The component of the drag force arising out of friction can be called Fdf and the component of drag force arising out of medium displacement can be called Fdd. With this we have the total drag force Fd as Fd = Fdf + Fdd Fdf and Fdd are equated as

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Fdf = Cdf * A * df * V2/2 and Fdd = Cdd * A * df * V2/2 where Cdf = Coefficient of drag due to the friction between the particle and medium Cdd = Coefficient of drag due to the need for displacing the medium df = Density of the material of particle V = Velocity at which the particle moves A = Area of cross section of the particle in the direction perpendicular to the direction of its movement. It can also be written as Fd = (Cdf + Cdd) * A * df * V2/2 or Fd = Cd * A * df * V2/2 When the velocity of the particle is low the value of Fdf makes a much larger contribution to Fd, but at very high velocity of the particle the contribution of Fdd to Fd is far higher than that of Fdf.

10.3.4 Terminal velocity Balancing these three forces acting on the particle an equation can be written as Gravitational force = Drag force + Buoyant force or Net Force on the particle = Gravitational force − (Drag force + Buoyant force) When the particle is falling in the medium its velocity keeps on increasing due the action of the gravity. The gravitational and buoyant force remain almost the same. The drag force keeps on increasing due to increase in the velocity of the particle. At one juncture the drag force increases and together with the buoyant force balances the gravitational force. At this juncture the net force on the particle becomes zero and the acceleration of the particle also becomes zero. After this the falling particle continues downward movement at a constant velocity, which is called the ‘terminal velocity’. In chemical engineering, where sedimentation of small particles is considered, it is also called settling velocity. The equation for settling velocity for spherical particles can be derived as follows: 4 * π * r3 * dm g/3 − 4 * π * r3 * df g/3 = Cd * π * r2 * df * V2/2 4 * r * dm g/3 − 4 * r * df g/3 = Cd * df * V2/2 V = ((8/(3 * Cd)) * r * g * ((dm − df)/df))0.5

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where r = Radius of the falling spherical particle dm = Density of the material of particle df = Density of the fluid Cd = Coefficient of drag g = Gravitational constant equal to 9.80665 m/s2. V = Velocity of the falling spherical particle.

10.3.5

Laws for settlement of particles in fluid

While considering the problem of settling of small particles in fluids, Stoke formulated that the drag force acting upon a particle moving in the fluid is given by the equation Fd = 6 * π * μ * r * V Later, he went on to derive the equation for settling velocity as follows: 4 * π * r3 * dm g/3 − 4 * π * r3 * df g/3 = 6 * π * μ * r * V i.e. 2 * (dm − df) * r2 = 9 * μ * V i.e. V = 2 * r2 * (dm − df)/(9 * μ) where r = Radius of the falling spherical particle dm = Density of the material of particle df = Density of the fluid μ = Dynamic viscosity of the medium g = Gravitational constant V = Velocity of the falling spherical particle. The above equation is known as Stoke’s Law. Stoke’s Law is valid when the fluid flow is in the region of Reynold’s Number from 0.0001 and 2. When the fluid flow is in the region of Reynold’s Number from 2 to 500 the settling velocity is given by an equation called the Intermediate Law. This equation is V = (0.306 * g0.71 * r1.14 * (dm − df)0.70)/(df0.29 * μ0.43)) where r = Radius of the falling spherical particle dm = Density of the material of particle df = Density of the fluid Cd = Coefficient of drag

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g = Gravitational constant equal to 9.80665 m/s2 V = Velocity of the falling spherical particle. Finally for fluid flow in the region of Reynold’s Number from 500 to 200,000 the settling velocity is given by an equation called Newton’s Law. In this region the coefficient of drag is fairly constant and is given by Cd = 0.44. Substituting this value into the equation for settling velocity we get, V = ((8/(3 * 0.44)) * r * g * ((dm − df)/df))0.5 = (6.06 * r * g * ((dm − df)/df))0.5 where r = Radius of the falling spherical particle dm = Density of the material of particle df = Density of the fluid g = Gravitational constant V = Velocity of the falling spherical particle. The important assumptions made in arriving at the above formula are: 1 2 3 4 5

The shape of the falling body is perfectly round so the surface area is at a minimum for the given volume. The surface of the falling body is very smooth. The falling bodies are located far apart from each other so that the presence of one body does not affect the fall of other falling bodies. The air through which the body is falling is still. The extent of air column through which the body is falling is several times larger than the diameter of the falling body.

Due consideration has been given to the deviations from these assumptions made in the following treatment of the subject. For a typical rock cutting formed in the formation fracturing process, we have a particle density of 2700 kg/m3, particle diameter 5 mm or 0.005 m and air density of 1.225 kg/m3. Substituting the values we get the terminal velocity as 18.09 m/s i.e. 1085.5 m/min or 3561 ft/min. It can be said that if the spherical particle is in a large annular space, when the upward velocity Va of the air in the annulus equals the terminal velocity Vt, the particle gets suspended in the air and will not fall down. Similarly, if the upward air velocity Va is more than the terminal velocity of the body Vt then the particle will be thrown up by the air at a speed of Va − Vt. In blasthole drilling practice, compressed air forced through the drill pipes is jetted out at the bottom of the hole through the bit nozzles. This air then drags the cuttings with it through the annular space between the hole walls and drill pipe. The

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upward velocity of air in the annular space is called ‘bailing velocity’ because at that velocity of the compressed air the particles are bailed out of the blasthole. Bailing velocity can easily be calculated by dividing the volume of compressed air by the area of the annular space. The equation can be expressed as Vb = 4 * 106 * Q/(π * (D2 − d2)) where Vb = Bailing velocity in m/min Q = Discharge of compressor in m3/min D = Diameter of the drill hole i.e. drill bit in mm d = Outer diameter of the drill pipe in mm Discharge volumes of compressed air are always given as free air discharge, meaning that the air will occupy an equivalent volume when the pressure of the air comes down to that of standard atmospheric air. For this purpose the 101.325 kPa is considered as standard atmospheric pressure and the discharge is also called SCFM when it is being expressed in imperial units. When drilling is carried out with water or mud circulation, the velocity of fluid moving upward is nearly constant through the depth of the hole. In contrast, in the case of compressed air circulation used in a blasthole drilling, the bailing velocity varies continuously because as the pressure drops, the air continuously expands. It is highest at the mouth of the hole.

10.4

FORMULATION OF DESIRED BAILING VELOCITY

How much bailing velocity is appropriate for particular conditions of the blasthole is a debatable matter. Under certain conditions in a blasthole, cuttings formed may be very large but the quantum of such cuttings may be very low, i.e. less than 0.5%. In such cases it is inappropriate to have a very large compressor working all the time. It is certainly more economical to get the chips re-fragmented and taken out of the blasthole. An appropriate value of the desired annular bailing velocity Vd for particular drilling conditions can be equated as Vd = 1.1α Vt where α = A factor depending upon many conditions described hereunder. Vt = A value for terminal velocity calculated on certain assumptions. Let us say that Vt is equal to 1100 m/min as calculated earlier for 5 mm spherical particles of density 2700 kg/m3 at mean sea level with air density of 1.225 kg/m3. The

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multiplier 1.1 is included so that the desired bailing velocity is 10% higher than the terminal velocity, thereby enabling the particles to move up with a velocity equal to 10% of the terminal velocity. The equation, therefore, turns into: Vb = 1210α The value of the correlation factor α must account for all the deviations from the assumed conditions used for calculating the value of 1100 m/min on the basis of the equation for terminal velocity. Aspects to be considered while arriving at an appropriate value of the correlation factor are: 1 2 3 4 5 6 7 8

Size of Fragments Density of Fragments Roundness of Fragments Roughness of Fragments Rate of Fragmentation Annular Space Inclination of the Blasthole Amount of Water Injection (if any) In other words the correlation factor α can be equated as α = αrαdαrouαrndαfrαaαiαw

where αr = Component depending upon the size of the fragments. αd = Component depending upon the density of fragments i.e. rock. αrou = Component depending upon the roundness of fragments. αrnd = Component depending upon the roughness of fragments. αfr = Component depending upon the rate of fragmentation of fragments. αa = Component depending upon the available annular space. αi = Component depending upon the inclination of the blasthole. αw = Component depending upon presence of water injection. In formulating the above equation it has been presumed that all these components of the correlation factor are independent of each other. Values of each of the above correlation factors can be determined from the conditions actually met in drilling on the basis of the elaboration given in the following paragraphs.

10.4.1

Size of fragments

Rock chips are formed at the bottom of the hole. In order that these chips are smoothly removed from the hole, they must essentially be smaller than about 90% of the size of the smallest passage through which they have to pass. If rock chips are larger than this size, the chips have to reduce in size by undergoing re-fragmentation.

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In almost any type of drilling with any type of bit, the smallest passage through which the rock chip has to pass is the flushing groove of the bit as shown in Figure 10.2. Studies of the geometry of different types of drill bits have suggested that most of the bits have 2 to 4 flushing grooves, each with a mean dimension of about 7 to 8% of the bit diameter. It is therefore appropriate to assume that the size of largest cutting is about one fifteenth of the bit diameter. For small and medium size blastholes field observations match very well with this correlation. In the case of large blastholes of diameter 300 mm or so, the largest cuttings were found to be of 19 mm size. Since the desired bailing velocity of 1100 m/s has been calculated for the particles of 5 mm size, which is equivalent to a hole diameter of 75 mm, the correlation factor αr will have a value of 1 for this blasthole diameter. Further, as per the equation for terminal velocity, the effect of particle size will be proportional to the square root of the diameter. With this we can determine one component of the correlation factor, viz. αr, on account of size of fragment as αr = (D/75)0.5 where D is the diameter of hole in mm.

Blasthole wall

One of the three grooves of tricone bit for cutting removal

Figure 10.2 Grooves around a tricone bit.

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10.4.2

Density of fragments

The density of rock being drilled, assumed while arriving at the value of 1100 m/min, is 2.7 (2700 kg/m3). In practice, the density of rock varies between about 2 for shells to 5 for magnetite ore. The correlation factor αd on account of particle density can be written as αd = (ρ/2700)0.5 where ρ is the density of the particles in kg/m3.

10.4.3

Roundness of fragments

While arriving at the desired bailing velocity of 1100 m/s the particles were assumed to be perfectly spherical. Roundness of fragments is very important. In almost all types of rocks the chips are far from spherical. In igneous or metamorphic rocks the chips are closer to tetrahedron or parallelepiped shapes, whereas in the case of sedimentary rocks they are close to flat disks. Coefficient of drag for non spherical chips is much higher, and therefore the bailing velocity required for lifting off such particles is much less than that required for spherical particles. Roundness or sphericity of a particle can be defined as the ratio of the surface area of a sphere of the same volume as the non spherical particle to the surface area of the non spherical particle. Table 10.1 presents sphericity coefficients of typical shapes. From the values presented in the table, it can be concluded that for particles of igneous and metamorphic rocks the value of sphericity of the cuttings will be about 0.75 and for sedimentary rocks it will be about 0.6. Very roughly, the drag coefficient can be equated to the sphericity. Therefore, the value of αrnd will be either 0.75 or 0.6.

10.4.4

Roughness of fragments

Roughness of the surface of particles does affect the terminal velocity. When the surface is rough the terminal velocity is less. As the size of the grains which form rock is coarser, the surface of rock particles becomes more rough because fracture generally takes place at the joint between two particles. However, since the particles formed in the drilling process are small, their roughness does not affect the terminal velocity

Table 10.1 Sphericity of shapes.

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Particle shape

Sphericity

Particle shape

Sphericity

Sphere Octahedron Cube Prism L × L × 2L Prism L × 2L × 3L

1 0.85 0.81 0.77 0.73

Cylinder H = R/15 Cylinder H = R/3 Cylinder H = R Cylinder H = 2 Cylinder H = 20R

0.25 0.59 0.83 0.87 0.58

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305

greatly. In case of large size chips, the reduction in terminal velocity on account of roughness of the particle surface can be as much as 10%. An estimated correlation factor on account of particle roughness can be written as αrou = 1/(1 + (D/1000)2) where D is the diameter of the hole.

10.4.5

Rate of fragmentation

Rate of fragmentation is directly dependent upon penetration rate. A higher penetration rate means more fragments. If the volume of fragments to be removed from the blasthole is large, a higher bailing velocity is needed. It is proposed that if the penetration rate is more than 15 m/h then for every 1 m/h increase in the penetration rate, the bailing velocity should be increased by 1%. An estimated correlation factor on account of the penetration rate of fragmentation can be written as αfr = (1 + 0.01 * (Pr − 15)) where Pr is the penetration rate in m/h.

10.4.6

Annular space

From field observations it has been noticed that for effective removal of rock cuttings the annular space should be about 17% of the cross sectional area of the blasthole. It is proposed that if the percentage of annular space is less than 17%, then for every 1% reduction in this percentage the bailing velocity needs to be increased by 2%. An estimated correlation factor on account of insufficient annular space can be written as αa = (1 + 0.02 * (17 − A)) where A is the percentage of annular space for the combination of the bits and drill rods to be used in actual drilling.

10.4.7

Inclination of the blasthole

If a hole is inclined the tendency of the fragments to dash against the wall of the hole is much greater. In such circumstances a higher bailing velocity is necessary for effective cutting removal. It is proposed that for every 1° inclination of the blasthole, the bailing velocity be increased by 0.5%.

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An estimated correlation factor on account of blasthole inclination can be written as αi = (1 + 0.005 * (I)) where I is the angle of inclination with the vertical in degrees.

10.4.8

Quantity of water injection

In many blasthole drills water is injected into the compressed air to minimize dust pollution. This method of dust suppression is not ideal but it is an inexpensive method. When water is injected into the compressed air a mist is formed. This misty air is then circulated in the hole. Lifting formation fragments in such misty air is rather difficult because particles – particularly the sand and dust size – tend to stick together and form somewhat larger chunks to be removed by the circulating compressed air. It is proposed that if water injection is used while drilling the bailing velocity should be increased by 10%. Thus, the correlation factor on account of the presence of water injection needs to be equal to 1 when water injection is not used and be equal to 1.1 when water injection is used. With all the reasoning and correlation factors mentioned above the final equation for bailing velocity takes the form as under, Vd = 1210αrαdαrouαrndαfrαaαiαw The above is the final equation for desirable bailing velocity of flushing air. It is particularly valid in blasthole drilling as the drilling depth is usually less than 30 m and almost never exceeds 50 m. In actual practice, a bailing velocity of about 900 m/min is considered as the bare minimum required, and about 1000 m/min is desirable. While drilling large diameter blastholes in soft rocks, bailing velocities of about 2000 m/min have given very satisfactory results. In heavy iron ores, bailing velocities of the order of 2500 m/min have given good results. In very abrasive formations bailing velocities higher than 2000 m/min cause very heavy abrasive wear on the drill string components. Therefore, bailing velocities beyond 2400 m/min are rarely used. This restriction has very little effect on the penetration rate. To look into the correctness of the above hypothesis and formula, bailing velocities for three typical cases have been calculated in Table 10.2. These values for desired bailing velocities, calculated by using the empirical formula for the three random cases of blastholes, tally well with results obtained in standard field practice. It thus appears that the empirical formula for bailing velocity can be used under many drilling conditions. Compressed air is far more efficient in lifting formation fragments from the hole bottom as compared to water or water-based mud used in oilwell drilling. The basic

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Table 10.2 Desired bailing velocity in m/min calculated after taking all the correlation components into consideration. Case 1

Case 2

Case 3

254 mm hole inclined at 20° in granite of density 2.6, sphericity 0.75, pen. rate 22 m/hr, annular space 18%, with water injection

406 mm vertical hole in sandstone of density 2.6, sphericity 0.6, pen. rate 35 m/hr, annular space 15%, without water injection

Factor

Aspect considered

311 mm vertical hole in iron ore of density 5.2, sphericity 0.75, pen. rate 18 m/hr, annular space 17%, without water injection

αr

Size of Fragments

2.03633658

1.840289832

2.326657116

αd

Density of Fragments

1.387777333

0.981306763

0.981306763

αrnd

Roundness of Fragments 0.75

0.75

0.6

αrou

Roughness of Fragments 0.911808929

0.939394053

0.858489951

αfr

Rate of Fragmentation

1.03

1.07

1.2

αa

Annular Space

1

0.98

1.04

αI

Hole Inclination

1

1.1

1

αw

Water Injection

1

1.1

1

2408.557682

1953.352769

1775.92082

Aspect for correlation factor

Desired Bailing Velocity m/min

reason for the efficiency of compressed air in lifting the cuttings of the bottom of the hole is that when compressed air escapes from the drill bit, it expands and the fragments are lifted from the bottom due to the drag force. Further the drag force on the particle keeps on acting for a long time due to continuous expansion of the compressed air that takes place till the particle reaches the mouth of the hole and escapes to the atmospheric air. This gives very high velocity to the fragments. As the depth of the blasthole increases, the requirement of compressed air pressure also increases. It is very important to keep the bailing velocity at the optimum as calculated in Table 10.2. Lower bailing velocity results in the following: 1 2 3 4 5

Regrinding of the chips under the drill bit. Reduced penetration rate. More chances of premature drill bit damage. Higher bit cost and hence higher drilling cost. Higher volume of fines and hence more load on the filters of the dust collector. Higher bailing velocity results in the following:

1 2

Chapter 10.indd 307

Higher power consumption and hence higher drilling cost. High wear and tear on drill string components.

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10.4.9

Other approaches

Many writers have given deep consideration to the formulation of the desired bailing velocity required in rotary blasthole drilling. Equations put forth by them are empirical in nature and are as follows: Vd = 573 * (ρr/(ρr + 1)) * dp0.6 Bauer and Crosby have put forth an equation for balancing velocity i.e. velocity for keeping rock particles floating in the compressed air flow: Vd = 250 * ρ0.5 d0.5 r * p In both the above equations, Vd = Desired bailing velocity in m/min ρr = Density of rock chips in g/cm3 dp = Mean diameter of rock chips in mm. An equation proposed for the discharge capacity of the compressor on the blasthole drill is Qd = 224 * D0.5 where Qd = Air flow in m3/min. D = Diameter of Blasthole in m.

10.5

FORMULATION OF COMPRESSED AIR PRESSURE

What volume is adequate has been arrived at on the basis of various aspects of flushing discussed in the earlier part of this chapter. Here, in this part, the subject of discussion is how much pressure is needed. Compressed air must have a certain built-in pressure to ensure that an adequate volume of air flows through the flow passage. As stated earlier and shown in Figure 8.3, the components through which compressed air flows can be grouped into four sections. The number and type of these components changes from one blasthole drill to another. These four sections are described, and their general composition in terms of components, is given in Table 10.3. Loss in compressed air pressure occurs in each of these four sections. It is possible to calculate the pressure lost in these sections for every blasthole drill, but the calculations are very tedious and more importantly the theoretical treatment of such calculations is far from perfect. Hence it is necessary to carry out air tests. Manufacturers of rotary blasthole drills often carry out air tests on their blasthole drill models. For such tests the apparatus used, and the test procedure, has been described in chapter 8.

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Table 10.3 Different sections in compressed air flow path in a rotary blasthole drill. Section Flow section no. location

Components likely used for the flow path in the section Comments on diameter and length

1

From Compressor The flow path comprises of to the Drill Dead metal tubing, one gate valve, eight 90° bends, one 180° bend, one piece of short hose pipe, one piece of long hose pipe.

The diameter of the metal tubes and hose pipes equals to about half the diameter of the largest blasthole. Length of both the hoses together equals about 1.5 times the length of longest drill pipe. Length of metal tubing also equals to about 1.5 times the length of longest drill pipe.

2

From Drill Dead to the Bit

Apart from its length each drill pipe has one gradual enlargement and one gradual contraction.

3

From Drill Pipe The drill bit, three nozzles Side of the Bit to and the clearance kept in Blasthole Side of the bearings of the drill bit. the Bit

The diameter of the nozzles is to be selected as per procedure explained later in this chapter. The underlying principle is that 30% of the compressed air should flow through bit bearings.

4

Blasthole Bottom This section is formed to its Opening at outside the drill pipes and the Ground Level the drill bit. Hence, no components are involved.

A small part of this section is below and on the side of the drill bit. It is very irregular. Grooves are also formed around the stabilizer. Apart from these the full section is uniform and annular in cross section. The annulus is formed by the diameter of blasthole and the outer diameter of the drill pipes.

One to five drill pipes. One Stabilizer that may be roller type.

Giving due consideration to the components in the four sections and considering average values of compressed air flow, diameters and lengths of metal pipe and hose, blasthole diameters, drill pipe inner and outer diameters etc., the values of compressed air pressure loss in sections 1 + 2 + 4 have been arrived at in Table 10.4. If we consider that smaller diameter blasthole bits have smaller API pin connection and much smaller clearance in the bit bearings, the smaller bits must ensure higher pressure loss through the nozzles so as to ensure that about 30% of the compressed air flows through the bearings. On the basis of this logic, the desired pressure loss values when 70% of compressed air flows through the nozzles have been given in Table 10.4. Field observations roughly tally with the values of desired pressure loss given in the second last row, and the actual pressure provided on the compressors of the blasthole drills is given in the last row in the respective categories in Table 10.4.

10.5.1

Oil injection

As stated earlier, oil is often introduced into the compressed air flow in the form of mist through an oil injection system. The purpose of this process is bit lubrication.

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Table 10.4 Estimation of pressure loss in pa in four sections of the flow path in a rotary blasthole drill. Pressure loss for different blasthole drills having following recommended range of blasthole diameters in mm Section and component Pressure Loss in Section 1 (Pa) Metal Tubing Valves 9 × 90° Bends 1 × 180° Bend Hose Pipe

150 to 200

200 to 279

279 to 381

381 to 445

40707

28451

21948

18180

16282.8

11380.4

8779.2

7272

Pressure Loss in Section 2 (Pa) 5 Drill Pipes

6800

6938

3346

819

Pressure Loss in Section 3 (Pa) Around Bit Annulus

1500 22500

1200 21800

900 13200

600 9600

Total Pr Loss Section 1 + 2 + 4 Pa

114989.8

97521.4

61557.2

39747

Desired Pressure Loss at Bit Pa

550000

400000

280000

210000

Min. Desired Comp. Pressure in Pa in psi

664989.8 96.5

497521.4 72.2

341557.2 49.83

249747 36.2

125 to 100

110 to 100

85 to 50

65 to 50

Common Pressure Ratings of the Compressors in psi

With the introduction of oil, the resistance to flow of compressed air increases. The increase, however, is very small. Quite likely the fine film of oil formed on the walls of the flow path reduces friction factor.

10.6

CHOOSING A DRILL PIPE

The diameter of blastholes is chosen on the basis of many factors related to production requirements, blasting requirements, and parameters pertaining to loading and hauling equipment used in the mine etc. They are elaborated elsewhere in this book. When a drill pipe is to be chosen for drilling blastholes of a particular diameter, the following factors are required to be given due consideration: 1 2 3 4 5

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Drill Pipe Dimensional Parameters Drill Pipe Surface Treatment Size and Shape of Drill Cuttings Bailing Velocity in the Blasthole Drill Pipe Wall Thickness.

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The following elaboration of these factors will give ample guidelines for selection of a drill pipe. Later in this section, an example is given to clarify a systematic procedure for drill pipe selection.

10.6.1

Drill pipe dimensional parameters

As far as the dimensional parameters are concerned, drill pipes are available in many different external diameters, wall thicknesses, lengths etc. Dimensional parameters of commonly available drill pipes are given in Table 10.5. In blasthole drills, drill pipes are handled through many mechanical devices. For example, they are stored in pipe racks or carousel type pipe changers, handled by centralizer, tightened by wrenches etc. All this makes it necessary to have perfect dimensions of a drill pipe in terms of pipe length, length of threaded portion, position and dimensions of wrench flats etc., so they do not pose any problems while being handled by these devices provided in blasthole drills. Most of the suppliers of drill pipes ask for the make, model and some other relevant details of the blasthole drill with which the drill pipes are to be used. Giving such data to the suppliers is not sufficient by itself. Purchasers should ask for the external dimension drawings for the drill pipe from the suppliers, and verify that the drill pipe will fit in the blasthole drill being used by them, before finalizing an order.

10.6.2

Drill pipe surface treatment

Surface treatment given to blasthole drill pipes is of two types viz. surface hardening and anti-corrosion. In oilwell or waterwell drilling, water and mud are used as circulating fluid. The annular velocity of such fluids is much lower and the wear of drill pipes due to friction with the drill cuttings is very small. In blasthole drilling compressed air is used as a circulating fluid. The velocity of air flow and the drill cuttings is very high. This results in heavy wear on the drill pipes due to such continuous sandblasting of the external surface of the drill pipes. For this reason it is always desirable to choose surface-hardened drill pipes in blasthole drilling. This need is far more pronounced when blastholes are to be drilled in hard and abrasive rocks such as granite, quartzite etc. than in soft and non-abrasive rocks such as slate, limestone etc. In many rock masses a small quantity of groundwater seeps into the blastholes. Such groundwater may not pose any problem for drilling but generates mild acids or other chemicals by getting mixed with some minerals. When the drill pipes come in contact with such acids they start corroding. In such conditions anti-corrosion treatment is important. Many manufacturers make drill pipes for water well drilling to be carried out with DTH hammers. For waterwell drilling smaller wall thickness is desirable, so longer drill strings used in deep wells can be lifted with ease. Feed force and torque required for DTH drilling are also much lower than those in rotary drilling with tricone bits.

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Table 10.5 Dimensional parameters of commonly available drill pipes. Drill pipe OD Inch 3.5

mm

Wall thickness

Cross section area

Body weight

in

mm

in2

lb/ft

mm2

kg/m

88.9

0.75

19.05

6.48

4180.3

22.05

32.84

4

101.6

0.5

12.7

5.498

3546.9

18.71

27.87

4.5

114.3

0.375

4.86

3135.3

16.54

24.63

4.5

114.3

0.5

12.7

6.283

4053.7

21.38

31.85

5

127

0.5

12.7

7.069

4560.4

24.05

35.83

5

127

0.75

19.05

10.014

6460.5

34.08

50.76

5.5

139.7

0.5

12.7

7.854

5067.1

26.73

39.81

5.5

139.7

0.75

19.05

11.192

7220.6

38.08

56.73

6

152.4

0.75

19.05

12.37

7980.6

42.09

62.7

6.25

158.75

0.5

12.7

9.032

5827.1

30.73

45.78

6.25

158.75

0.75

19.05

12.959

8360.7

44.1

65.69

6.25

158.75

1

25.4

16.493

10640.9

56.12

83.6

6.5

165.1

0.75

19.05

13.548

8740.7

46.1

68.67

6.625

168.28

0.864

21.946

15.637

10088.5

53.21

79.26

6.75

171.45

0.75

19.05

14.137

9120.7

48.11

71.66

7

177.8

0.5

12.7

10.21

6587.2

34.74

51.75

7

177.8

0.75

19.05

14.726

9500.8

50.11

74.64

7

177.8

1

25.4

18.85

12161

64.14

95.55

7.5

190.5

1

25.4

20.42

13174.4

69.49

103.51

7.625

193.68

0.75

19.05

16.199

10450.8

55.12

82.11

7.625

193.68

0.875

22.225

18.555

11971

63.14

94.05

7.625

193.68

1

25.4

20.813

13427.7

70.82

105.5

8.625

219.08

0.906

23.012

21.97

14174.5

74.76

111.37

8.625

219.08

1

25.4

23.955

15454.6

81.51

121.42

8.625

219.08

1.5

38.1

33.576

21661.7

114.25

170.19

9.25

234.95

0.75

19.05

20.028

12921

68.15

101.52

9.25

234.95

1

25.4

25.918

16721.3

88.19

131.38

9.25

234.95

1.5

38.1

36.521

23561.9

124.27

185.12

9.25

234.95

1.5

38.1

36.521

23561.9

124.27

185.12

9.625

244.48

1.5

38.1

38.288

24702

130.29

194.08

10.75

273.05

1

25.4

30.631

19761.6

104.23

155.26

10.75

273.05

1.25

31.75

37.306

24068.6

126.94

189.1

10.75

273.05

1.5

38.1

43.59

28122.2

148.33

220.95

10.75

273.05

2

50.8

54.978

35469.5

187.08

278.68

11.75

298.45

1.25

31.75

41.233

26602.1

140.31

209.01

9.525

(Continued)

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Table 10.5 (Continued) Drill pipe OD

Wall thickness

Cross section area 2

2

Body weight

Inch

mm

in

mm

in

mm

lb/ft

kg/m

12.25

311.15

1

25.4

35.343

22801.8

120.26

179.15

12.75

323.85

1

25.4

36.914

23815.2

125.61

187.11

13.375

339.73

1

25.4

38.877

25082

132.29

197.06

13.375

339.73

1.25

31.75

47.615

30719.1

162.02

241.35

13.375

339.73

1.5

38.1

55.96

36102.9

190.42

283.65

14

355.6

1.5

38.1

58.905

38003

200.44

298.58

14

355.6

2

50.8

75.398

48643.9

256.56

382.18

14.375

365.13

1.5

38.1

60.672

39143.1

206.45

307.54

14.375

365.13

2

50.8

77.754

50164

264.58

394.13

15

381

1.25

31.75

53.996

34836.1

183.74

273.7

This makes the thin-walled drill pipes acceptable in waterwell drilling. Further, in water well drilling dimensional inaccuracies of the drill pipes are tolerable to a greater extent, as drill pipes are handled manually and stored on the ground and not in a pipe changer. For blasthole drilling such drill pipes do not have adequate wall thickness and are without any surface hardening treatment so they should be avoided.

10.6.3

Size and shape of drill cuttings

In the formulation of bailing velocity in the earlier part of this chapter, the following two assumptions were made in respect of the largest drill cuttings that are formed while blasthole drilling. 1 2

While drilling in igneous and metamorphic rocks the size of the largest cuttings is about 1/15 times the diameter of the blasthole. While drilling sedimentary rocks the size of the largest cuttings is about 1/12 times the diameter of the blasthole.

Reasoning behind these assumptions is that drill cuttings in sedimentary rocks are flaky and have sphericity of 0.6 as against 0.75 for the cuttings in igneous and metamorphic rocks. Observations about fluid flow patterns around objects of different shapes reveal that if a cutting has to pass smoothly through the annulus, the annulus should have thickness equal to 1.25 times the dimension of the cutting size. Thus the desired thickness T for the annulus can be equated as T ≥ D/12 for igneous or metamorphic rocks T ≥ D/9.6 for sedimentary rocks.

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With this hypothesis, the optimum outside diameter d for the drill pipe can be equated as d = D − 2 * (D/12) = 5 * D/6 = 0.833 * D for igneous and metamorphic rocks or d = D − 2 * (D/9.6) = 3.8 * D/4.8 = 0.792 * D for sedimentary rocks. In actual practice it may not be possible to get a drill pipe with outside diameter calculated as above. In such an event a drill pipe having the nearest, and preferably slightly smaller, OD should be chosen for blasthole drilling. Dimensional parameters for standard blasthole drill pipes are given in Table 10.4. An appendix at the end of this book gives more details.

10.6.4

Bailing velocity

In order to clean the hole effectively the compressed air flow in the annulus must have sufficient bailing velocity. This topic has already been discussed earlier in this chapter. Accordingly the desirable bailing velocity Vd can be calculated. Once the outside diameter of a drill pipe is determined, and a standard drill pipe is chosen as stated in previous subsection, it is necessary to check if the compressor on the blasthole drill has adequate discharge capacity to give the desired bailing velocity in actual operation. The desired capacity for the compressor can be calculated by the following equation: Qd = (π * (D2 − d2) * Vd)/(4 * 106) where Qd is the desired discharge capacity of the compressor on the blasthole drill in m3/min, D is the diameter of blasthole in mm, d is the diameter of drill pipe in mm and Vd is the desired bailing velocity in m/min. In almost all the cases it will be found that the compressor on the blasthole drill has more than adequate discharge capacity because compressors on blasthole drills are most usually chosen to drill blastholes of largest rated diameter.

10.6.5

Drill pipe wall thickness

Drill pipes with larger wall thickness are always advantageous in blasthole drilling. The reasons are: 1

2

3

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Drill pipes with larger thickness are heavier and their heavy weight adds to the feed force exerted by the feed mechanism to give even higher feed force on the drill bit. Unlike other purposes such as oilwell or waterwell drilling, in blasthole drilling the drill string is very short, so the heavier weight of the drill pipes does not pose a limitation on lifting the drill string. Drill pipes in oilwell or water drilling are always in tension, whereas those in blasthole drilling are in compression. Thus, blasthole drill pipes are susceptible to buckling. When a blasthole drill string is exerted with a feed force just above its

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315

critical buckling load, the drill string suddenly starts deflecting in the center. Such deflection substantially increases the possibilities of blasthole deviation. Drill pipes with greater wall thickness have a much higher section modulus and have a much higher critical load against buckling. Thus they do not buckle and contribute to hole deviation and other undesirable factors such as rubbing against the wall of the blasthole. A drill pipe with large wall thickness may be good in blasthole drilling for long life and higher bit loading, but a drill string composed of such drill pipes is frequently required to be lifted by using power. Consumption of power for lifting heavier drill strings is always higher. In this regard the following observations are quite self explanatory. 1

2

3

4

Very heavy drill collars are essential in the oilwell drilling industry as the drills normally do not have any mechanisms for exerting positive feed force on the bit. Bit weight is solely attained by use of heavy drill collars. Integral drill pipes are the same as drill collars with some dimensional difference. Integral drill pipes have very large wall thickness – in most of the cases far higher than an equivalent fabricated drill pipe, as can be seen from details given in one of the appendices at the end of this book. If larger wall thickness was the only criteria for selecting the best drill pipe for blasthole drilling, everyone would have chosen integral drill pipes rather than fabricated drill pipes. Actually the trend is otherwise. Manufacturers who solely deal in blasthole drilling products rarely attempt to manufacture integral drill pipes. Integral drill pipes are manufactured only by those who manufacture drill collars for oilwell drilling.

The above elaboration sufficiently indicates that in addition to the outside diameter due consideration must also be given to appropriate wall thickness of the drill pipe to be used for blasthole drilling. One important factor to be kept in mind while choosing a drill pipe for blasthole drilling is that the drill pipes are stored in pipe racks, which are vertically placed in the mast. If the drill pipes are very heavy the stability of the blasthole drill is seriously affected. When a drill pipe with a specific outside diameter is to be chosen from the table of available drill pipes given in one of the appendices at the end of this book, there may be many options of wall thickness for such drill pipes. Further, even if such an option does not exist, it is necessary to check the susceptibility of the available drill pipe to buckling. Critical buckling load can be defined as the axial compressive load at which the drill string starts deflecting somewhere in the midst of its length. In the case of a drill string with uniform cross section, the point at which deflection starts happens to be at the midpoint of the drill string. As per Euler’s formula, critical buckling load can be equated to properties of the drill string as under. P = 1000 * C * π2 * E * I/(1000 * L)2

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where P is the critical buckling load in newtons, C is the coefficient dependent upon the end fixing conditions for the drill string, E is the modulus of elasticity for the material of the drill string in GPa, I is the moment of inertia for the drill string in mm4 and L is the length of the drill string in m. Since components of the drill string are made from either low carbon or alloy steel tubes, E is equal to 200 GPa. Since most of the drill string length consists of drill pipes, the moment of inertia for the drill pipes can be used in calculations. For a drill pipe with outside diameter D in mm and inside diameter d in mm, the moment of inertia I in mm4 is given by the equation: I = (π/64) * (D4 − d4) If both the end points of the drill string are purely pin type, as shown in Figure 10.3A, then the value of C equals to 1. If both the drill string ends are rigidly fixed, as shown in Figure 10.3B, then the value of C equals to 4. In actual practice the drill string is restrained from buckling at the top by the deck bushing, and at the bottom it is usually restrained from buckling by the combination of drill bit and stabilizer. For such conditions the value of C is estimated at nearly 3.7. While using the length of drill string in Euler’s formula, the following points can be made. 1

When drilling is to be carried out by using only the base drill pipe, the drill string length is equal to the length of base drill pipe plus an additional length of about 2 m due to the drill bit, stabilizer, saver sub at the upper end and some length of

B

A

C

Column Length

Both Ends Pinned

Both Ends Fixed

Both Ends Semi Fixed

C=1

C=4

C = Variable (1 to 4)

Figure 10.3 End fixing conditions of drill string.

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2

317

the shock absorber. Therefore, the length L to be used for calculations works out to L = Lp + 2, where Lp is length of base drill pipe. When drilling is to be carried out by using many drill pipes, and the maximum depth of the blasthole, the length L to be used for calculations works out to L = Lh − 1, where Lh is depth of the blasthole. Deduction of 1 m is done because a stabilizer effectively restrains the drill string from buckling.

In actual calculations the larger of the two values should be used. Once the buckling load for the chosen drill pipe is calculated by using Euler’s formula it must be verified that the drill pipe can withstand the maximum rated feed force of the blasthole drill (and not the maximum feed force likely to be exerted on the drill string) with a factor of safety more than 1.25 to allow for the vibrations etc. Drill pipes affect the stability of a blasthole drill considerably. When the diameter and wall thickness etc. of the chosen drill pipe differs from the original, it is essential to verify if the stability of the blasthole drill will remain unaffected. This can be done by checking with the blasthole drill manufacturer.

EXAMPLE

An example of systematic selection of a drill pipe for rotary blasthole drilling operations is illustrated through Table 10.6. In the table, cell A1 states what is in the cell C1. Cell B1 specifies whether contents of cell C1 is data input or the results of the calculation. If the contents are data input, cell B1 mentions “data input”. If the contents are a formula cell B1 gives the formula used for calculations. Repetitions with the next higher wall thickness may be necessary if pipe thickness in cell C35 is found to be inadequate. In the example shown in Table 10.6 the depth of blasthole was taken as 20 m. This was because the drill was to be used in a copper mine where the primary loading equipment was an electric shovel and hauling was to be carried out by dumper. Had the drill to be used in a gigantic coal mine where a dragline was the primary equipment for removal of overburden, the blasthole depth would have been considerably greater, say about 50 m. In such a case the length for calculating the critical buckling load would have increased from 22 m to 49 m. As a result of this the critical buckling load would have reduced considerably.

10.7

DISCARDING A DRILL PIPE

As drilling is carried out by using a drill pipe the drill pipe goes on wearing; In particular its outer diameter reduces considerably. As the outer diameter reduces, the annular area increases and the bailing velocity decreases. In actual practice this decrease can be compensated for by increasing the volume flow from the compressor. Most of the modern screw compressors have the facility of increasing or decreasing volume flow by a simple control. Obviously, increasing the volume flow is possible up to the maximum discharge capacity of the compressor.

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Table 10.6 Choice of suitable drill pipe. Aspect

Formula used (if any)

Input/output

Cell ref.

Make and Model of the Blasthole Drill Used

Data Input

Bucyrus 49RIII

C1

Method of Drilling

Data Input

Rotary

C2

Maximum Rated Pulldown of the Blasthole Drill in kg

Data Input

54800

C3

Maximum Discharge of Compressor in m3/min

Data Input

103

C4

Pressure oh the Compressed Air in kpa

Data Input

448

C5

Maximum Rated Depth Capacity in m

Data Input

85.5

C6

Single Pass Length in m

Data Input

21.3

C7

Type of Formation to be Drilled

Data Input

Granitic Copper Ore

C8

Nature of Formation Igneous or Metamorphic or Sedimentary

Data Input

Igneous

C9

Feed Force Required on the Bit for Drilling the Formation in kg/mm

Data Input

90

C10

Blasthole Diameter in mm

Data Input

406

C11

Actual Feed Force on the Drill String in kg

C12 = C10 * C11

36540

C12

Can the Blasthole Drill Exert Necessary Feed Force?

C13 = IF(C12 < C3, "Yes", "No")

Yes

C13

Depth of Blasthole in m

Data Input

20

C14

0

C15

Inclination of the Blasthole in Degrees Data Input 3

Data Input

2700

C16

Data Input

12

C17

Annular Space Datum in %

Data Input

20

C18

Presence of Water Injection

Data Input

No

C19

Desired Drill Pipe Diameter in mm

338.198 C20 = IF(C9 = "Sedimentary", 0.792 * C11,0.833 * C11) Data input 339.72

C20

Length of Drill String for Calculating Critical Buckling Load in m

C22 = IF(C14 > (C7 + 2), C14 − 1,C7 + 2)

23.3

C22

Factor for Size of Fragment

C23 = (C11/75)^0.5

2.326657116

C23

Factor for Formation Density

C24 = (C16/2700)^0.5

1

C24

0.75

C25

0.858489951

C26

Density of the Formation in kg/m

Expected Penetration Rate in m/hr

Nearest Available Drill Pipe OD in mm

Factor for Roundness of Drill Cuttings C25 = IF(C9 = "Sedimentary", 0.6, 0.75) Factor for Roughness of Drill Cuttings

C26 = 1/(1 + (C11/1000)^2)

C21

(Continued)

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319

Table 10.6 (continued) Aspect

Formula used (if any)

Input/output

Factor for Rate of Fragmentation

C27 = 1 + 0.01 * (C17 − 15) C28 = 1 + 0.02 * (17 − C18)

1.07

C27

0.94

C28

C29 = 1 + 0.005 * C15 C30 = IF(C19 = “No”, 1, 1.1)

1

C29

1

C30

Calculated Desired Bailing Velocity in m/min

C31 = 1210 * C23 * C24 * C25 * C26 * C27 * C28 * C29 * C30

1823.164542

C31

Discharge Required from Compressor in m3/min

C32 = (PI/4) * (C11^2 − C21^2) * 70.77398847 C31/10^6

C32

Does the Compressor on the Drill Give Sufficient Air Discharge?

C33 = IF(C32 < C4, “Yes”, “No”) Yes

C33

Type of Drill Pipe (Alloy Steel or Mild Steel)

Data Input

Alloy Steel

C34

Available Thickness of the Chosen Drill Pipe in mm

Data Input

25.4

C35

Factor for Annular Space Factor for Hole Inclination Factor for Water Injection

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Cell ref.

ID of the Drill Pipe in mm

C36 = C21 − 2 * C35

288.92

C36

Moment of Inertia for the Drill Pipe in mm4

C37 = (PI/64) * (C21^4 − C36^4)

311771256.3

C37

Critical Buckling Load in kg

C38 = (3.7 * 1000 * PI^2 * 200 * C37/(1000 * C22)^2)/9.80665

427695.1119

C38

Factor of Safety Against Buckling

C39 = C38/C3

11.70484707

C39

Is Chosen Drill Pipe Appropriate for Blasthole Drilling ?

C40 = IF(C39 > 1.25, “Yes”, “No”)

Yes

C40

Actual Area Ratio (Annular Space Area/Hole Area)

C41 = (C11^2 − C21^2)/C11^2

29.98514985

C41

Outside Diameter of the Chosen Drill Pipe in mm

C42 = C21

339.72

C42

ID of the Drill Pipe in mm

C43 = C35

288.92

C43

Maximum Bit Loading in Blasthole Drilling in kg

C44 = C12

36540

C44

Minimum Desired Factor of Safety Against Buckling

Data Input

3

C45

Critical Buckling Load to Which the Drill Pipe Must Withstand in kg

C46 = C44 * C43

109620

C46

Minimum Acceptable Diameter of the Drill Pipe After Abrasive Erosion on the Basis of Buckling Load Criteria

C47 = ((64/PI) * C12 * C45 * (1000 * C22)^2 * 9.80665/(3700 * PI^2 * 200) + C36^4)^0.25

304.489999

C47

Minimum Acceptable Diameter of the Drill Pipe After Abrasive Erosion on the Basis of Desired Bailing Velocity By Using Max. Compressor Discharge

C48 = (c11^2 − (4 * c4 * 1000^3/(3.14159 * c31 * 1000)))^0.5

304.8016539

C48

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As the diameter of the drill pipe reduces the moment of inertia decreases, and the possibility of buckling of the drill pipe also increases because of decreased critical load. It is possible to find out what is the minimum outside diameter to which the selected drill pipe can be allowed to wear down, on each of the above considerations. This has been done in rows 41 to 48, where calculations reveal the following. Minimum acceptable diameter of the drill pipe after abrasive erosion on the basis of buckling load criteria = 304.49 mm. Minimum acceptable diameter of the drill pipe after abrasive erosion on the basis of desired bailing velocity by using maximum compressor discharge = 304.80 mm. It is somewhat strange that the minimum acceptable diameters are almost the same in both cases. This happens rarely. When the diameters differ the higher value should be chosen so both the criteria, i.e. adequate bailing velocity and safety against drill pipe buckling, are satisfied. Therefore the drill pipe can be allowed to wear from the original 339.72 mm diameter to 305 mm diameter in actual drilling. Once the diameter of any part on its length becomes equal to 305 mm, it should be discarded. Particular attention must be given to the desired factor of safety against buckling i.e. data input in cell C45. About 2 to 3 can be considered as adequate.

10.8

CHOOSING NOZZLES FOR TRICONE BITS

A tricone bit releases compressed air through three nozzles as illustrated in Figure 10.4. The nozzles of the bit are separate from the bit and have different bore diameters. As shown in Figure 8.10 they are shaped so as to efficiently convert air pressure into kinetic energy without turbulence loss. Nozzles can be fixed into the bit by means of a circlip like the one shown in Figure 10.4, or by means of screwing into the threads provided, or by means of a pin as shown in Figure 10.5. The advantage of use of the nozzle is that when the air jet is directed towards the blasthole bottom, the cuttings (formed by the interaction of the bit teeth with the formation) are separated from the formation very rapidly and are sent to the annulus for bailing from the blasthole. The main reason for the quick separation is that when air goes in the cracks formed in the formation it exerts drag forces on the cutting formed by the cracks developed on all the sides of the bit. Similarly the small quantity of air that fills the cracks starts expanding. It has been experienced that use of appropriate bit nozzles gives a faster penetration rate than is possible by directing the air flow towards the blasthole bottom through the central hole, shown in Figure 5.10. It is also true that the penetration rate increases with the increase in velocity of the air emerging from bit nozzles. Nozzles for a tricone bit must be selected meticulously to get the fastest penetration rate but at the same time they must also give long life to the tricone bits. This can be achieved by keeping the rotary speed as well as bit weight within the limits stated in

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Nozzle

321


Air Jet Emerging from Nozzle

Figure 10.4 Jetting of the compressed air by bit nozzles.

Figure 10.5 Fitting a nozzle in tricone bit by a pin.

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322


chapter 5 and ensuring that 20%–30% of the compressed air volume passes through the bearing of the tricone bit. In chapter 8 the equation for the flow of compressed air was given as Q = (60/1.225) * α * ψ * p1 * A * [2/(R * T1)]0.5 where Q = Volume flow rate in m3/min α = Nozzle efficiency ψ = Flow coefficient p1 = Upstream absolute pressure in Pa T1 = Upstream absolute temperature in K A = Area of nozzle opening in m2 R = Gas constant in J/(kg ⋅ K). This equation can be taken as the basis for nozzle selection in a tricone bit. Common sizes of nozzles available for tricone bits are given in Table 10.7. If a compressor has a discharge capacity Qc at a pressure rating of pc then the compressor keeps on giving compressed air to the air flow conduits so long as the pressure loss in the conduits is less than or equal to the pressure rating. Now let us assume that for a particular diameter of blasthole and the diameter of the drill pipes used for blasthole drilling the required volume flow is Qr and the compressor provided on the rotary blasthole drill is adequate. This means Qc > Qr. Since modern screw compressors have discharge controls to control the output of the compressor, we also assume that the compressor is maintaining a discharge of only Qr to the flushing system. Therefore each nozzle should be of such size that it passes a discharge of 0.70 * Qr through it. Table 10.7 Commonly available nozzle sizes. Nozzle Size No.

7

8

9

10

11

12

13

Nozzle Dia (in.)

7/32

1/4

9/32

5/16

11/32

3/8

13/32

Nozzle Dia (mm)

5.55625

6.35

7.14375

7.9375

8.73125

9.525

10.31875

Nozzle Size No.

14

15

16

17

18

19

20

Nozzle Dia (in.)

7/16

15/32

1/2

17/32

9/16

19/32

5/8

Nozzle Dia (mm)

11.1125

11.90625 12.7

13.49375 14.2875

15.08125 15.875

Nozzle Size No.

21

22

23

24

25

26

28

Nozzle Dia (in.)

21/32

11/16

23/16

3/4

25/32

13/16

7/8

Nozzle Dia (mm)

16.66875 17.4625

18.25625 19.05

19.84375 20.6375

Nozzle Size No.

30

32

34

36

38

40

Nozzle Dia (in.) Nozzle Dia (mm)

15/16 23.8125

1 25.4

11/16 26.9875

11/8 28.575

13/16 30.1625

11/4 31.75

22.225

Some manufacturers make nozzles for their tricone bits in sizes ranging from 5 mm to 28 mm in steps of 1 mm.

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323

To fully utilize the pressure developed by the compressor, a pressure loss to be achieved while this volume of compressed air flows through the flushing system, can be found from following equation: pr = pc − (pd + pa) where pr = Pressure loss to be achieved through use of nozzles pc = Pressure of compressed air at air compressor pd = Pressure lost while compressed air flows to the tricone drill bit pa = Pressure required at the blasthole bottom to ensure that cutting laden air passes through the annulus. It can be said that under most circumstances (pd + pa) is less than 100 kPa. Table 10.8 gives the diameter of each of the three nozzles that together will pass the compressed air flow given in column 1. Nozzle diameter values in this table are calculated for a temperature of 35°C. This table can be used for choosing nozzles of appropriate size. The following example will clarify the method of choosing nozzles.

EXAMPLE

A rotary blasthole drill is being used for drilling blastholes of diameter 311.15 mm with a 228.6 mm drill pipe. The formations being drilled are such that bailing velocity should be about 2200 m/min for efficient hole cleaning. The compressor on the blasthole drill is capable of giving discharge up to 85 m3/min at a governed pressure of 448 kPa. Find the diameter of each of the nozzles to be used with the tricone bit. Atmospheric temperature and pressure at the worksite are 35°C and 101.325 kPa.

SOLUTION

From Table 3.1A in appendix 3 it can be seen that for getting a bailing velocity of 2200 m/min in a blasthole of diameter 311.15 being drilled with drill pipes of diameter 228.6 mm, the volume of compressed air required will be 76.99 m3/min. It is, therefore, presumed that the operator will govern the discharge of the compressor to 77 m3/min. Of this flow volume it is necessary that 70% or 53.9 m3/min flow must pass through the nozzles. From Table 10.8, for a compressor discharge of 78 m3/min and compressor pressure of 448 kPa, the appropriate diameter of nozzle is 21.97 mm. The nearest size nozzle available for the bit is 20.6375 or 22.225 mm. In almost all the cases it is better to select the nozzle with smaller diameter than larger, because a smaller nozzle builds up higher pressure but discharges less air through the nozzles and ensures more compressed air flows through the bearings of the tricone bit. Verification that the nozzle does give about 30% of air flow through bearings is needed.

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Table 10.8 Nozzle selection aid. Pressure pc

1034

862

758

689

448

414

379

350

Pressure pd + pa

100

100

100

100

100

100

100

100

Pressure pc − (pd + pa )

934

762

658

589

348

314

279

250

Volume flow

Nozzle sizes for dissipating above pressure pc − (pd + pa ), when the compressed air flow volume is as shown on left cells

m3/min

cfm

20

706.29

7.32

8.02

8.55

8.97

11.12

11.57

12.09

12.58

22

776.92

7.68

8.41

8.97

9.41

11.67

12.14

12.68

13.20

24

847.55

8.02

8.79

9.37

9.83

12.19

12.68

13.25

13.79

26

918.18

8.35

9.15

9.75

10.23

12.68

13.19

13.79

14.35

28

988.81

8.66

9.49

10.12

10.61

13.16

13.69

14.31

14.89

30

1059.44

8.97

9.82

10.48

10.99

13.62

14.17

14.81

15.41

32

1130.07

9.26

10.15

10.82

11.35

14.07

14.64

15.30

15.92

34

1200.70

9.54

10.46

11.15

11.70

14.50

15.09

15.77

16.41

36

1271.33

9.82

10.76

11.48

12.04

14.92

15.53

16.23

16.88

38

1341.96

10.09

11.06

11.79

12.37

15.33

15.95

16.67

17.35

40

1412.59

10.35

11.34

12.10

12.69

15.73

16.36

17.10

17.80

42

1483.22

10.61

11.62

12.40

13.00

16.12

16.77

17.53

18.24

44

1553.85

10.86

11.90

12.69

13.31

16.50

17.16

17.94

18.67

46

1624.47

11.10

12.16

12.97

13.61

16.87

17.55

18.34

19.09

48

1695.10

11.34

12.43

13.25

13.90

17.23

17.93

18.74

19.50

50

1765.73

11.57

12.68

13.52

14.18

17.59

18.30

19.12

19.90

52

1836.36

11.80

12.93

13.79

14.47

17.94

18.66

19.50

20.29

54

1906.99

12.03

13.18

14.05

14.74

18.28

19.01

19.87

20.68

56

1977.62

12.25

13.42

14.31

15.01

18.61

19.36

20.24

21.06

58

2048.25

12.47

13.66

14.57

15.28

18.94

19.71

20.59

21.43

60

2118.88

12.68

13.89

14.81

15.54

19.27

20.04

20.95

21.80

62

2189.51

12.89

14.12

15.06

15.80

19.59

20.37

21.29

22.16

64

2260.14

13.09

14.35

15.30

16.05

19.90

20.70

21.63

22.51

66

2330.77

13.30

14.57

15.54

16.30

20.21

21.02

21.97

22.86

68

2401.40

13.50

14.79

15.77

16.54

20.51

21.34

22.30

23.20

70

2472.03

13.70

15.01

16.00

16.78

20.81

21.65

22.63

23.54

72

2542.66

13.89

15.22

16.23

17.02

21.11

21.96

22.95

23.88

74

2613.29

14.08

15.43

16.45

17.26

21.40

22.26

23.26

24.21

76

2683.91

14.27

15.64

16.67

17.49

21.69

22.56

23.58

24.53

78

2754.54

14.46

15.84

16.89

17.72

21.97

22.85

23.88

24.85

80

2825.17

14.64

16.04

17.11

17.94

22.25

23.14

24.19

25.17

(Continued)

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325

Table 10.8 (Continued) Pressure pc

1034

862

758

689

448

414

379

350

Pressure pd + pa

100

100

100

100

100

100

100

100

Pressure pc − (pd + pa )

934

762

658

589

348

314

279

250

Volume Flow cfm

Nozzle sizes for dissipating above pressure pc − (pd + pa ), when the compressed air flow volume is as shown on left cells

82

2895.80

14.82

16.24

17.32

18.17

22.52

23.43

24.49

25.48

84

2966.43

15.00

16.44

17.53

18.39

22.80

23.71

24.78

25.79

86

3037.06

15.18

16.63

17.74

18.60

23.07

24.00

25.08

26.10

88

3107.69

15.36

16.83

17.94

18.82

23.33

24.27

25.37

26.40

90

3178.32

15.53

17.02

18.14

19.03

23.60

24.55

25.65

26.70

92

3248.95

15.70

17.20

18.34

19.24

23.86

24.82

25.94

26.99

94

3319.58

15.87

17.39

18.54

19.45

24.12

25.09

26.22

27.28

96

3390.21

16.04

17.57

18.74

19.66

24.37

25.35

26.50

27.57

m3/min

98

3460.84

16.20

17.76

18.93

19.86

24.62

25.61

26.77

27.86

100

3531.47

16.37

17.94

19.13

20.06

24.87

25.88

27.04

28.14

103

3637.41

16.61

18.20

19.41

20.36

25.24

26.26

27.45

28.56

106

3743.35

16.85

18.47

19.69

20.65

25.61

26.64

27.84

28.97

110

3884.61

17.17

18.81

20.06

21.04

26.09

27.14

28.36

29.51

For this purpose the nozzle flow equation given in chapter 8 has to be used. In this case we have p = 348 kPa α = 0.95 ψ = 0.4841 R = 287.1 J/kg ⋅ K p1 = 348000 + 101325 = 449325 Pa T1 = 15 + 273.16 = 288.16°K A = π * (22.25/1000)2/4 = 0.00033451 m2 Hence M = 0.95 * 0.00033451 * 0.4841 * 449325 * 2/(287.1 * 288.16)]0.5 or M = 0.33986 kg/s or Q = 16.6464 m3/min Therefore, discharge from three nozzles will be 49.93 m3/min and balance 77 − 49.93 = 27.07 m3/min or 27.07/77 = 35.15% air will flow through the bearings. If under the same conditions, nozzles of diameter 22.225 were chosen then the discharge through each nozzle would have been 19.3058 m3/min and the three nozzles would have discharged 57.92 m3/min. Under such conditions the flow through the bearings would have been 77 − 57.92 = 19.08 m3/min or 19.08/77 = 24.78%.

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For ensuring 30% discharge through the bearings it would have been necessary to control the discharge from the compressor at 57.92/0.7 = 82.75 m3/min. This would have resulted in a large increase in the bailing velocity to about 2380 m/min. As stated earlier, drill manufacturers carry out air tests to verify that bearings of the tricone bit are getting adequate air supply. These tests and the method of selecting nozzles adopted by them is elaborated in an appendix.

10.9

CHOOSING THE RIGHT COMPRESSOR

The compressor is an integral part of a rotary blasthole drill. It cannot be changed without incurring very heavy expenses. It must, therefore be appropriately chosen while purchasing the blasthole drill itself. All the manufacturers provide the most appropriate compressor on their blasthole drills for normal working conditions. But when the blasthole drill is to operate at altitude or for drilling in a very heavy mineral mass, such standard compressors prove insufficient and an optional compressor has to be chosen. As this topic is more related to altitude operation it is discussed in the next chapter that deals with operation of the blasthole drill at high altitude.

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Chapter 11

Effect of altitude and severe weather

11.1

INTRODUCTION

Blasthole drilling is carried out by using a blasthole drill and many other accessories described in earlier chapters of this book. The working of some of these components does depend upon the atmospheric parameters and other environmental conditions that change with altitude. The height of a workplace from mean sea level is called altitude. As many mines are located above 1500 m altitude and some are as high as 4500 m altitude, it is important to know the effects on such components. There is a well established relation between most of the environmental parameters and the altitude of the place. Parameters such as temperature, pressure and density of the atmospheric air do change with altitude. Humidity is not dependent upon altitude but the surroundings. These changes affect the working of many items used in the drilling process. In a very strictest sense, even the process of rock fracture must also be affected due to the changes of the atmospheric parameters mentioned above. However, the effect is so negligible that it is not even researched. Apart from the above changes, the worksite often experiences severe operating conditions such as stormy winds, extreme cold, extreme heat, heavy rainfall, heavy snowfall etc. These too affect working of the drill and other components. This chapter is dedicated to giving the essential knowledge about the effect of altitude and severe weather on the blasthole drilling process.

11.2

ATMOSPHERE

The atmosphere surrounds the earth and extends to about 3000 km in height. The mass of all the matter in this zone is estimated to be about 151.7 × 106 kg. The matter in the atmosphere is collectively called air and consists of many gases as shown in Table 11.1. The percentage of component gases marked with an asterisk in this table varies to some extent due to several reasons. In addition, there are organic and inorganic impurities in the form of dust, water vapor etc. The proportion of impurities and water vapor greatly depends upon the surroundings of the place. The atmosphere itself is subdivided into five zones viz. troposphere, stratosphere, mesosphere, ionosphere and exosphere as shown in Figure 11.1. Of these, the troposphere extends up to an altitude of 11 km. As all the drilling sites in the world are located in this zone we need be concerned with this zone only.

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328


Table 11.1 Composition of air in the atmosphere. Component

% by vol

% by mass

Nitrogen

78.08

75.71

Oxygen

20.95

23.15

Argon

0.93

1.28

Carbon Dioxide*

0.03

0.05

Neon

0.0018

0.00125

Helium

0.00052

0.00007

1.5 * 10

−04

9.4 * 10−05

Krypton

1.1 * 10

−04

2.9 * 10−04

Carbon Monoxide*

1.0 * 10−04

2.0 * 10−04

Nitrous Oxide*

5.0 * 10

−05

8.0 * 10−05

Hydrogen*

5.0 * 10−05

3.5 * 10−05

Ozone*

4.0 * 10

−05

7.0 * 10−06

Xenon

8.0 * 10−06

3.6 * 10−06

Nitrogen Dioxide

1.0 * 10

−07

2.0 * 10−07

Iodine

2.0 * 10−11

1.0 * 10−10

Radon

−18

5.0 * 10−17

Methane

6.0 * 10

Exosphere - From 410 to nearly 3000 km

Ionosphere - From 90 to 410 km

Mesosphere - From 35 to 90 km Stratosphere - From 11 to 35 km

Troposphere - From 0 to 11 km

Figure 11.1 Different zones of atmosphere.

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11.2.1

329

Effects of altitude on atmosphere

Pressure, temperature, as well as density of the atmospheric air, slowly reduces with increasing altitude. Changes in these parameters with respect to altitude are given in one of the appendices at the end of this book. The data presented in the appendix is applicable for latitude of 40°. The atmospheric pressure and air density remain virtually unaffected by latitude but the temperature is affected considerably. Usually precise indication of atmospheric temperature is not required for evaluation of any factor in blasthole drilling, however, and the following rule of thumb can help in this context. To estimate mean annual temperature of a place located between latitude 0° (i.e. equator) and latitude 75° (generally the northernmost or the southernmost inhabited place), add 0.7°C in the 40° latitude temperature for decrease of each 1° latitude or subtract 0.7°C from the 40° latitude temperature for increase of each 1° latitude.

11.2.2

Effects of humidity on atmosphere

Up to a certain limit, water vapor can get mixed with air and behave like a gas i.e. without condensation. The presence of water vapor in the air is called humidity. The amount of vapor in the air is most simply expressed as its weight per unit volume. There is a limit to the amount of water vapor that a given volume of air can hold. This limiting amount of water vapor makes the air saturated and any additional water vapor in the air, if present, gets condensed immediately. At higher temperatures air can hold more water vapor than at lower temperatures. Atmospheric pressure does not have any bearing on the saturation of air so long as the temperature is constant. Absolute humidity is measured as grams of water vapor per cubic meter of air. Measurements of humidity in this absolute manner, however, are very difficult hence the term relative humidity is more commonly used. For different temperatures, the absolute humidity required for saturation of the air with water vapor, are well established by means of laboratory tests. In general, air is not saturated. It contains only a fraction of the possible water vapor. This fraction, expressed in percentage, is called relative humidity. In reality the above is an oversimplified version of absolute and relative humidity, and hence a little imprecise, because these quantities are measured on the basis of pressure rather than weight. The vapor pressure of saturated air for different temperatures is given in Table 11.2. The actual water content of moist air at different temperatures and relative humidity is given in Figure 11.2. Table 11.2 Vapor pressure of saturated air for different temperatures.

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Temp in °C

Vapor pr. in kPa

Temp in °C

Vapor pr. in kPa

−10 −5 0 5 10 15 20

0.260 0.402 0.611 0.872 1.227 1.705 2.337

25 30 35 40 45 50

3.167 4.263 5.624 7.378 9.580 12.300

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11.3

EFFECT OF SCARCE AIR IN DRILLING

In rotary blasthole drilling atmospheric air is required for three functions as follows: 1 2 3

The atmospheric air gives sufficient oxygen to the burning processes inside the cylinders of a diesel engine. The atmospheric air absorbs heat from various components that generate or carry generated heat during the drilling process. The atmospheric air, in compressed form, is circulated through the blasthole for the purpose of cutting removal by flushing.

In the following sections the influence of changes in atmospheric parameters on various components used on drilling machines and in the drilling process is elaborated.

11.3.1

Diesel engines

A diesel engine generates power from the combustion of fuel within its cylinders. The combustion requires oxygen that is obtained by the engine from the atmospheric air. If, for any reasons, the oxygen available from the atmosphere is not sufficient then less quantity of fuel is burnt and the diesel engine generates less power. Every diesel engine has a rated power. The rated power is based on the tests carried out on the diesel engine on a test bed in a laboratory. Temperature and humidity in the laboratory are controlled within certain limits. Since the laboratory is located at a known altitude, the other atmospheric parameters are well established. Diesel engine manufacturers rate their engines for some well defined conditions as specified by one of the various standards. The standards also define the factors of deration to be applied to the engine power output when the engine is used under different atmospheric conditions. The best way, therefore, is to use the factors stipulated by the standards for the purpose of engine derating. Table 11.3 gives the standard conditions used for the purpose of engine derating by some of the well known standards. There are two distinct types of diesel engines viz. naturally aspirated and turbocharged. In a naturally aspirated engine the atmospheric air is simply sucked in through the engine valves. The pressure of the air, before it is sucked, is the same as the atmospheric pressure. In turbo-charged engines, however, there is a turbo charger which, by its fan, sucks the air to the engine suction mechanism. Naturally this fan acts as a very low pressure compressor, and therefore the pressure of the air before it is sucked in by the engine suction mechanism is somewhat higher than the atmospheric pressure. This results in more air being sucked into a turbo-charged engine as compared to an identical naturally aspirated engine. More often, the power output of a turbo-charged engine is purposely restricted by the amount of fuel pumped into the engine, rather than the volume of air sucked into it. Some engines can be equipped with special turbo chargers that enable engine operations in excess of 3000 m altitude without power deration. Due to these factors, a turbo charged engine is required to be derated to a much lesser extent than a equivalent naturally aspirated diesel engine.

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Relative Humidity %

10

50

20

30

40

60

80

100

40

Temperature in °C

30

20

10

0

−10

−20

1

2

4

6

8

10

20

40

3

Water Contents in g/m

Figure 11.2 Water content of air for different atmospheric temperature and relative humidity values.

60

80

100

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Table 11.3 Standard conditions for engine power output as defined by different standards. Standard

Pressure kPa

Temperature °C

Rel. humidity %

SAE (American) BS (British) DIN (German) ISO (International) IS (Indian)

100 100 100 100 100

25 27 27 27 27

0 60 60 60 60

Table 11.4 Rules of thumb for engine power deration.

Engine type Naturally Aspirated Turbo charged

Altitude in m up to which no deration is required 500 m Variable between 2000 m to 3000 m

Deration in % for each 100 m increase in altitude

Deration in % for each 10

1 1

3.6 5.4

A rule of thumb for engine power deration is given in Table 11.4. It is to be specifically noted that turbo-charged engines need not be derated up to an altitude of 2000 m, or 3000 m in some cases. As far as naturally aspirated engines are concerned a more precise formula for engine power deration, developed on a technological basis, is as follows: R = [(PO − RO ∗ PHO)/(PT − RT ∗ PHT)] ∗ [(TT + 273.15)/(TO + 273.15)]0.5 where the variables are: a

Values used for engine power rating PT = Absolute pressure in kPa. RT = Relative humidity in percent. TT = Atmospheric temperature in °C. PHT = Vapor pressure at saturation coexistent with temperature TT, in kPa.

b

Values for workplace of engine operation PO = Atmospheric pressure in kPa. RO = Relative humidity in percent. TO = Atmospheric temperature in °C. PHO = Vapor pressure at saturation coexistent to temperature TO, in kPa.

Values of PT and PO can be found from the appendix that gives properties of atmospheric air and PHT and PHO from Table 11.2.

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EXAMPLE

A naturally-aspirated diesel engine has a continuous rating of 400 kW as per a standard that specifies ratings to be given at sea level under 20°C air temperature and 60% relative humidity. Find the rated power of the engine when it is operated at an altitude of 1400 m where prevailing temperature and relative humidity are 35°C and 80% respectively. SOLUTION

From the data of engine rating conditions it can be concluded that PT = 101.325 kPa RT = 0.6 as given in the example data TT = 20°C. as given in the example data PHT = 2.337 kPa (from Table 11.2) From the details of the operating conditions it is evident that PO = 85.6 kPa. RO = 0.8 as given in the example data TO = 35°C. as given in the example data PHO = 5.624 kPa (from Table 11.2) The above data yields the deration factor to be R = [(85.6 − 0.8 ∗ 5.624)/(101.325 − 0.6 ∗ 2.337)] ∗ [(20 + 273.150/(35 + 273.15)]0.5 or R = 0.7914 From the deration factor we get the power output at the altitude and the other conditions as P = 0.7914 ∗ 400 = 316.56 kW

11.3.2

Electric motors and transformers

Electric motors generate heat while converting the electrical energy to mechanical energy. In a similar manner, a transformer also generates heat while the power supply is transformed from one type to the other. The dissipation of heat from such units to the surroundings is quite a complex process, as all the three modes of heat transfer viz. conduction, convection and radiation are involved. Electric motors and transformers are rated on the basis of certain temperature rise permitted in the machine over the ambient temperature, i.e. temperature of the surrounding atmospheric air, measured by a thermometer at a distance of 2 m or more from the machine. Distance and temperature rises differ for different standards.

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In actual practice the motors and transformers dissipate about 60% of the heat through convection and the balance by conduction and radiation. At higher altitude, where the density of air is less than that at sea level, the dissipation of heat through convection is considerably affected, whereas there is no effect on the conduction and radiation of heat. The deration of electric motors and transformers depends upon many factors. Hence, for a particular set of environmental conditions, the best way of knowing the deration is to refer the problem to the motor or transformer manufacturers because only they know the values of various parameters needed for the solution. In cases where the problem is required to be solved without referring to the manufacturer, for whatever reason, the deration on account of altitude should be obtained from Figure 11.3 which illustrates the deration factors for various altitudes. The graph in Figure 11.3 is based on cooling air temperature of 40°C or less. Occasionally mines are located at places where the temperature of cooling air itself is more than 40°C. In such cases the deration in the deration factor, of 1% for each 1°C difference between the air temperature and 40°C datum, may be taken and the new deration factor thus obtained may be applied in addition to the deration factor on account of altitude. The need for having both the deration factors, viz. one on account of altitude and the other on account of temperature, arises only under unlikely conditions because at altitudes beyond 1000 m, air temperature hardly ever rises beyond 40°C. The following example, therefore, is only for the purpose of illustration and understanding. EXAMPLE

A 500 kW electric motor, designed for use up to an altitude of 1000 m, is intended to be used at a place having 1900 m altitude where environmental temperatures occasionally rise to about 48°C. Find the power output of this motor at such a place. SOLUTION

From the graph in Figure 11.3 the deration factor for 1900 m altitude works out to 0.955. Since the temperature rises up to 48°C, this deration factor should be reduced further to an extent of (100 − 8)/100 = 0.92.

Deration Factor

1.00 0.95 0.90 0.85 0.80

0

1000

2000 Altitude in m

3000

4000

Figure 11.3 Deration factor for electric motors and transformers.

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335

This gives the corrected deration factor to be 0.955 * 0.92 = 0.8786 and the derated power of the motor works out to 500 * 0.8786 = 439.3 kW. Whatever has been said above for electric motors also applies to electric transformers because the cooling of transformers also is in the same pattern as that of an electric motor.

11.3.3

Hydraulic system

As has been seen earlier, the hydraulic system of a blasthole drill consists of many components such as hydraulic pumps, hydraulic motors, hydraulic cylinders, hose pipes, valves etc. At higher altitude the barometric pressure is less than that at sea level. If the hydraulic system of the drill is a closed loop system then there is no effect as far as the power transmission capability of the system is concerned, but the system is more susceptible to bubble formation. This can be taken care of by providing a somewhat larger bleeder tank, if necessary. If the hydraulic system is open type then the positive suction pressure, which is generally present at the hydraulic pumps, reduces. This makes it necessary to provide a pressurized hydraulic tank or place the tank at higher level with respect to the hydraulic pumps. It may be noted that up to 1500 m altitude the effects mentioned are not very critical, and hence in many cases no remedial measures may be necessary up to this altitude. Air also takes away the heat generated by the hydraulic system through the heat exchanger. Since the density of the air is less at higher altitude, the dissipation of the heat is also slower. But side by side the air temperature is also less. As these two counter-balance each other, no remedial measures are required unless the altitude of operation exceeds 1500 m. The remedial measures are in terms of providing an hydraulic oil cooler of larger size. The best course of action to be taken by a user is to consult the manufacturer prior to the purchase of the blasthole drill since only they are aware of various factors considered while designing the heat exchanger of the hydraulic system. One has to take into account the increase in viscosity of the hydraulic fluid at higher altitude due to the decreased temperatures. For this the solution is very simple. Just change the hydraulic fluid with a thinner fluid that flows easily. Usually this information is readily available in the service manuals of the drill. In absence of such information the manufacturer should be consulted. Oil companies make such information available through their product literature. In fact, most of the drill manufacturers just repeat this information to their customers.

11.3.4

Air compressors

Low atmospheric pressures and low temperatures experienced at high altitudes affect the compressors on rotary blasthole drills to a very great extent. A compressor delivers less compressed air as the altitude becomes higher. Actually the power required for the compressor also decreases as per altitude. However, the deration of engine power output is at a higher rate than the reduced power rate for compression. Hence it is necessary to adopt an engine with higher power.

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As stated earlier the compressors used on rotary blasthole drills are positive displacement type, i.e. screw or vane compressors. Single stage compressors that develop sufficiently high pressure for rotary drilling are usually chosen for rotary blasthole drills. Table 11.5 gives values for discharge correction factors and percentage reduction in power requirements for compressing air to 689 kPa (100 psi) at different altitudes for oil-flooded single stage screw compressors as well as oil-free two stage screw compressors. The table is in imperial units but it contains factors for ACFM, SCFM, power reduction etc. Hence, the values of these factors can easily be interpolated either by Newton’s Method for Interpolation (preferred) or the Linear Interpolation Method. Here again, the best guides for tackling the situations are the compressor manufacturers who have all the design data for their compressors. In practice, when a user requires a blasthole drill to operate at a certain high altitude and informs the manufacturer accordingly, the manufacturer usually proposes some changes in the configuration of the blasthole drill. As far as the compressor is concerned these changes are: 1 2

Using a larger size compressor instead of the one that is required to achieve the necessary flushing parameters at sea level. Providing an engine or an electric motor with a larger power output as a compressor drive. An elaborate example in this regard is given elsewhere in this book.

11.3.5

Blasthole flushing

For flushing the blasthole, compressed air moves from the compressor discharge port to the drill bit through metal pipes on the blasthole drill and then the drill string components up to the bit where it is discharged in the blasthole. The pressure loss experienced by the compressed air while it flows through the accessories does depend upon the compressor discharge pressure, but does not quite depend upon the atmospheric parameters. Therefore, the pressure of the compressed air just before the drill bit is the same so long as the compressor discharge pressure is unaltered. This is true for any atmospheric pressure. Now, consider that pressures of the compressed air prevailing in the flushing path are as given in Table 11.6. The above table gives the pressure values for otherwise identical conditions of blasthole drill, drill string components, formation, drill bit, blasthole depth etc. As the pressure P0 is set at the compressor output valve, it does not change. Therefore, p0 = P0. Since the operation at sea level and at altitude does not change the components in the flow path on the blasthole drill or the components in the drill string, the pressure loss will remain the same. Therefore, (P0 − P1) = (p0 − p1). But since P0 = p0, p1 = P1, we also have p3 < P3. The pressure loss experienced in the annulus will also remain same, as the flow path is unchanged i.e. P2 − P3 = p2 − p3. Therefore, p2 < P2. The above means P1 − P2, i.e. pressure loss in the nozzles of the bit, increases because P1 − P2 < p1 − p2.

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Table 11.5 Effect of altitude on discharge capacity and power requirement of screw compressors.

Atmospheric Altitude ft

Pressure psia

Temperature deg F

Relative humidity %

0 1000 2000 3000 4000 5000 6000 7000 8000 9000 10000

14.7 14.18 13.66 13.17 12.69 12.23 11.78 11.34 10.90 10.49 10.10

68 64 61 57 54 50 47 43 39 32 32

36 36 36 36 36 36 36 36 36 36 36

Oil flooded single stage rotary screw compressors

Oil free two stage rotary screw compressor

Correction factor

Correction factor

Inlet pressure psia

ACFM

SCFM

0.3389 0.2995 0.2641 0.2325 0.2042 0.1791 0.1567 0.1368 0.1191 0.1035 0.0898

1.0000 0.9984 0.9969 0.9954 0.9940 0.9926 0.9912 0.9899 0.9887 0.9874 0.9862

1.0000 0.9718 0.9438 0.9164 0.8899 0.8636 0.8383 0.8133 0.7878 0.7638 0.7408

Total

Reduction in power %

ACFM

SCFM

Total

Reduction in power %

1.0000 0.9703 0.9409 0.9122 0.8845 0.8572 0.8309 0.8124 0.7789 0.7542 0.7305

0.0 1.8 3.5 5.2 6.9 8.5 10.1 11.6 13.1 14.6 16.1

1.0000 0.9998 0.9995 0.9991 0.9988 0.9985 0.9982 0.9979 0.9976 0.9974 0.9971

1.0000 0.9718 0.9438 0.9164 0.8899 0.8636 0.8383 0.8133 0.7878 0.7638 0.7408

1.0000 0.9716 0.9433 0.9156 0.8888 0.8623 0.8368 0.8116 0.7860 0.7618 0.7386

0.0 1.1 2.9 4.7 6.3 8.0 9.5 11.1 12.6 14.0 15.4

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Table 11.6 Pressure locations in the compressed air flow path. Location of air pressure

At sea level

At altitude

Pressure at the Compressor Outlet Before the Tricone Bit After the Tricone Bit i.e. At Blasthole Bottom After the Blasthole i.e. Atmospheric Pressure

P0 P1 P2 P3

p0 p1 p2 p3

The increase in pressure loss is due to the fact that air is ejected through the flow nozzles at a higher speed. This actually means that at higher altitude the blasthole flushing improves. Besides the above, when the compressed air passes through the flow path, it cools to a lower temperature at higher altitude than at sea level because the atmospheric air temperature is lower at higher altitudes. To whatever little extent it may be, this low temperature compressed air must be improving the heat dissipation at the bit bearing and thus resulting in the increased bit life.

11.3.6

Drill lubrication

Temperature and pressure reduce at high altitude. In blasthole drills the pumps used for either oil injection into compressed air for bit lubrication, or for supplying lubricant to different parts, do not have much suction head. Therefore, lower atmospheric pressure at high altitudes does not affect the pumping process to a significant extent. However, in colder temperatures at high altitude lubricating oils become thicker and do not flow easily. Many multi-viscosity synthetic lubricants that suit the temperatures year-round are available from lubricant manufacturers.

11.4

EFFECT OF SEVERE WEATHER

Severe weather conditions are often experienced at a blasthole drilling worksite even when it is not located at very high altitudes. Typically the conditions are: stormy winds, heavy rainfall, heavy snowfall and extremely cold or hot weather. There are no rigorous definitions for distinguishing the conditions as normal or severe, but some rough data presented in Table 11.7 will serve the purpose. The following elaboration on how the high altitude or severe weather affects the drill and drilling process are intended for imparting preliminary knowledge. If a blasthole drill is expected to be operated at a certain location, the user must find out the altitude and weather conditions likely to be experienced at the worksite throughout the year. Such information should be provided to the drill manufacturer right when the drill is being selected for a purchase. A manufacturer is best equipped to suggest the best alternatives to combat the weather conditions.

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Table 11.7 Severe weather criteria. Aspect

Normal conditions

Severe conditions

Winds

Velocity <50 km/h

Velocity >50 km/h

Rainfall

<25 mm/h of a short period of an hour of <100 mm/day

>25 mm/h of a short period of an hour of >100 mm/day

Cold Weather

Temperature >0°C

Temperature <0°C

Hot Weather

Temperature <40°C

Temperature >40°C

Snowfall

<300 mm/h of a short period of an hour or <1000 mm/day

>300 mm/h of a short period of an hour or >1000 mm/day

11.4.1

Stormy winds

Stormy winds affect the working of a blasthole drill in four different ways. 1

2 3 4

A rotary blasthole drill is equipped with a tall mast. In high velocity winds the drag pressure generated on the mast gives rise to high stresses in the mast structure, its mounting, the main frame and leveling jacks. As large forces are acting at a height above the ground level, the stability of the blasthole drill is seriously affected. A quantum of pebbles, sand and dust blown in the atmosphere settle down on the drill. Visibility becomes very poor due to dust and clay mixed with the atmospheric weather.

It is believed that most blasthole drills can be operated in winds with a velocity up to 50 km/h from the viewpoint of strength and stability, or at least they should be designed to do so. Design engineers of the drill manufacturers are better equipped to give guidelines in this regard. With the modern equipment, weather forecasts have become quite reliable, particularly about storms. Warning of storms is often available 24 hours in advance. Further, a storm rarely starts suddenly. Wind velocity goes on increasing over a time interval of 15 minutes to even 2 hours. Blasthole drilling operations can be planned by keeping in mind the possibility of a storm. In this connection recommendations made by the drill manufacturers must be adhered to. In extreme situations the drilling has to be stopped and the mast has to be lowered. If the area of blasthole drilling is likely to meet stormy weather very often, the drill should be equipped with a machinery house to prevent sand or dust settling on the components mounted on the main frame. Poor visibility can be tackled by torch lights so long as the helper can be comfortable in the storm.

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11.4.2

Heavy rainfall

The nuisance caused by very heavy rainfall is much less severe on the blasthole drill, particularly when it is equipped with a machinery house, but is quite troublesome as far as the drilling process is concerned. In heavy rain, water can flow or seep into a blasthole. Drilling may have to be stopped in situations where further drilling is expected to be impossible. Water collected in the blastholes may have to be removed by using air-operated dewatering pumps. As far as a blasthole drill is concerned, the possibility of electric current leakage due to rain water and consequent short circuits should not be ignored. In rainy weather the electric batteries lose power quickly due to humidity and need attention. A dry dust collection system is particularly affected by water in the hole because the filters in the dust collectors do not get cleaned by the vibrations imparted by the mechanism in the dust collector. In heavy rain the ground surface also becomes damp and slippery. If the surface is waterlogged it can become very difficult to place the blasthole drill on the correct collaring position of a blasthole.

11.4.3

Heavy snowfall

Heavy snowfall is a bird of the same feather as heavy rainfall. The severity of effects of heavy snowfall is, however, less than for heavy rainfall. Measures and precautions suggested in case of heavy rainfall are also applicable in the case of heavy snowfall.

11.4.4

Extreme hot weather

No special measures are necessary to combat hot weather when the ambient temperatures are below 30°C. Even in these temperatures it is advisable to have a drill cab equipped with an air conditioner so the driller operates the drill comfortably. Up to a temperature of 40°C no basic changes in the machine are required so long as lubricants recommended for these temperatures are selected. A machinery house proves very beneficial when the drill has to operate at temperatures up to 40°C. It can be kept cool by an air conditioner. Even when temperature is likely to shoot up to 55°C, use of larger radiators for cooling the blasthole drill components and an air conditioner for the machinery house prove adequate. A double-walled machinery house can help in such circumstances. At high temperatures a water sprinkling system on the roof of the machinery house can also be beneficial.

11.4.5

Extreme cold weather

Blasthole drills have been used in places where the temperatures dipped to even −70°C. Under such conditions, if wind blows, the temperature can drop down to even −100°C. Ill effects of cold weather are summarized hereunder.

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1 2 3 4 5 6 7 8

341

Personal discomforts to driller/assistant. Increase in viscosity of all fluids. Freezing of the water used on drills. Condensation of atmospheric vapor. Troubles in starting the diesel engine. Segregation of components in the lubricants. Increase in brittleness of solid materials. Reduction in pressure ratings of the hoses.

The severeness of the ill effects increases rapidly as the temperatures decrease. Usually a rotary blasthole drill is designed in such a way that it can operate satisfactorily in all extreme weather conditions except extreme cold. Apart from use of appropriate oils and lubricants recommended for cold weather conditions, no specific device is required to be incorporated in the blasthole drill. However, for use of a blasthole drill in extreme cold weather, many additional devices are required to be included in the drill when it is manufactured. Details of such devices have already been given in Chapter 7 on rotary blasthole drills.

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Chapter 12

Computers in rotary blasthole drilling

12.1

INTRODUCTION

The word “computer” as it is interpreted today may have a specific meaning, but genetically every piece of an equipment that performs mathematical and or logical operations on the data received, can be considered as a computer. Even the computer industry accepts that abacus, slide rule, calculating machines etc., were the forerunners of present day computers for they had the ability to perform numerical manipulation. Similarly, this book recognizes that all the devices or systems used on blasthole drills, whether analog or digital in nature, whether for gaining data about the formation or for introducing some automation in the process of drilling, or for positioning the blasthole drill, are computers on rotary blasthole drills. The digital computers have one advantage over the earlier ‘computers’. They can store data, do logical and mathematical analysis of the data received and come to a decision in a far more reliable manner with higher accuracy in a very short time. This is not a book on electronics and computers to explain “why and how it is so”, but one must recognize that if analog data are transmitted there is always a loss of accuracy. This loss of accuracy is primarily due to the loss of signal strength experienced in the transmission medium, such as loss of electric voltage or loss of hydraulic pressure etc. However, when digital data are transmitted, it is binary in its form and is always in terms of existence or non-existence of a particular property of the transmitting and storing medium, e.g. presence or absence of magnetic field at certain point on a floppy disk, or existence (or non-existence) of voltage of certain intensity in the conductor wire etc. Therefore, even if the intensity of the magnetism or voltage of the data changes, its interpretation remains the same. Under the worst circumstances the data may not be read due to a faulty reading device, or may not be transmitted at all by a faulty transmitting device, and so may have to be retransmitted or re-read but it will never change its meaning. This chapter covers the modes of computerization of rotary blasthole drills and their benefits. 12.2

DRILL OPERATION

Blasthole drilling is carried out in repetition of four main activities. These are, 1 2

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Tramming Leveling

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3 4


Drilling Pipe changing

These four activities are the part of a cycle that repeats for drilling each of the thousands of blastholes.

12.2.1 Tramming For tramming activity the clock starts when a blasthole is complete, all the drill rods have been withdrawn from the blasthole and the drill is ready to move to the next blasthole. In this activity the leveling jacks are withdrawn and then the drill is actually trammed to the location of the next blasthole. In some cases back and forth movements of the drills may have to be made to ensure that the blasthole will be precisely drilled at its intended location and in precise alignment. If the blastholes to be drilled are inclined, the judgment of the operator must be very sharp. In some blasthole drills it is necessary to remove all the drill rods from the drill string, store them in the pipe rack and then move the rotary head to the lowermost position in the mast before tramming the drill. Such drills have sufficient stability only when the drill head is in the lowermost position. A remote propel option, described in chapter 7, is of great help in precisely positioning the blasthole drill on the position of the new blasthole. Faster tramming speed of a drill is of moderate importance since the distance to be trammed is of the order of 10 to 12 m only. In most cases the tramming activity needs 2 to 3 minutes.

12.2.2

Leveling

Once the blasthole drill is positioned precisely on the intended location, it has to be leveled so it gets a firm base for the drilling operation. Leveling is done by means of lowering the leveling jacks. The clock for the leveling activity starts once it is time to start extending i.e. lowering the leveling jacks. In older blasthole drills the leveling was done by the driller. Many modern blasthole drills have an automatic leveling system for the job. Leveling activity is carried out by the driller with guidance received from the levels and indicators placed in the operators cab and the signals given by the helper who stands outside. Usually leveling activity requires about 1 minute.

12.2.3

Drilling and Pipe Handling

Drilling and pipe handling is carried out in tandem, with pauses in between as required. Initially the drilling activity begins with low feed force, slow rotary speed and lesser supply of compressed air. With this, the blasthole starts in precise alignment and the likelihood of its deviation from the intended alignment is greatly reduced. This phase of drilling the first two meters or so is called collaring.

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Full-fledged drilling with most appropriate feed force, feed speed and rotary speed starts after the collaring phase. When the blasthole is drilled to the maximum depth attainable with the first drill pipe, and it is necessary to drill deeper, drilling is paused for adding a drill pipe. Thus the drilling clock is paused and the pipe handling clock begins. For adding a new drill pipe to the drill string, at first the existing drill pipe is gripped by the spanner at the bottom of the mast (Figure 7.36) and then the drill stem wrench i.e. hydraulic breakout wrench (Figure 6.21) is used to loosen the sub attached to the bottom of the drill head spindle or the shock absorber attached below the drill head. Further loosening of the drill pipe is carried out with the drill head itself after deactivation of the breakout wrench. When the drill head separates from the drill pipe the drill head is moved to the top of the mast. Activation of the pipe rack ensures that the new drill pipe comes in exact alignment of the drill string. Then the drill head is rotated to attach the drill pipe to it, and later to the first drill pipe held by the spanner. Retraction of the spanner allows the recommencing of drilling activity. The activities described above may have to be repeated 3 or 4 times till the desired depth of the blasthole is reached. In the final pipe handling, drill pipes are withdrawn, detached from the drill string and stored in the pipe rack till the drill bit comes above the ground level. The time required for adding a new drill pipe, including the drill head movements, is typically 1.25 to 2 minutes. Loosening the drill pipe, storing it back to the pipe rack, moving down the drill head and starting to withdraw the lower drill pipe takes about 1.5 to 2.5 minutes. Time spent in drilling activity is greatly variable and can be anywhere from 15 minutes to even 2 hrs. The drill can be trammed to the next blasthole only after it is lowered on the ground by retracting the hydraulic jacks and is standing on its crawler tracks. Lowering the blasthole drill by lowering the leveling jacks requires about 1 minute. Total time required for one cycle is about 0.25 to 1.5 hr.

12.3

MODES OF BLASTHOLE DRILL COMPUTERIZATION

Full-fledged computerization of a rotary blasthole drill usually consists of four subsystems as follows: 1 2 3 4

Drilling Knowledge System Drill Automation System Drill Navigation System Integrated Mining System

Each of these subsystems has a specific objective and is incorporated in a blasthole drill by using specific components. With exception of the integrated mining

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system, the other three subsystems are almost exclusive of each other in the sense that only one of the three can be present on a rotary blasthole drill. The objective of drilling knowledge systems, in short, is to get data about the drilling operation and the material being drilled. The objective of a drill automation system is to introduce automation in various activities to be carried out in a drilling cycle. Means of introducing automation in rod handling, leveling and drilling have been in existence for a long time. A drill navigation system aims at automating the tramming activity of the drilling cycle. Finally, the integrated mining system has a the ultimate objective of controlling all the operations of all the equipment used in the mine. To achieve this, it is essential to channel some output from the drilling activities to the activities of other mining equipment. When any of the above subsystems has to be built in a rotary blasthole drill, it is essential to incorporate some components in the drill. These have been mentioned in Table 12.1. To work properly and achieve the objective, there has to be communication between various components used for computerization. If only the blasthole drills are to be computerized to optimize drilling and blasting, the communication between the components is like the one shown through schematics in Figure 12.1. Table 12.1 Items needed in different types of computerized drilling systems. If needed for following system Item Transducer for Measurement of Position of Drill Head Rotary Speed Rotary Torque Compressed Air Pressure Compressed Air Flow Vertical Vibrations in Drill String Horizontal Vibrations in Drill String Vertical Vibrations in Mast Horizontal Vibrations in Mast Drill Level Mast Angle with Vertical System Unit Equipped for Display of Drilling Knowledge Data Additional Auto Drill Software Additional Drill Position Software Display Monitor Device for Manual Input to System Control Unit Cabinet Radio Transmitter Computer and Network Cabling GPS Transmitters GPS Ground Stations

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Drilling knowledge

Auto drill

Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes

Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes

Drill position

Yes Yes Yes Yes

Yes Yes Yes

Yes Yes Yes Yes Yes Yes Yes Yes

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Signals from Different Transducers

Input and Storage Device

Rotation Control

Drill Knowledge System Unit

Radio Transceiver Blasthole Drill

Feed Control

Drill Level

Propel Control

GPS Transmitter

Auto Drill System Unit

Drill Positioning System Unit

Screen for Parameter Display

Blasthole Position Information

Satellites in the Sky

Ground Antennas

Data Flow Lines

Main Office

Personal Computer for Blast Design

Cable Transmission

Radio Transceiver

Wireless Transmission Mainframe Computer for Complete Mine Management

Figure 12.1 Schematics of data flow in rotary blasthole drill equipped with drill knowledge, auto drill and GPS positioning system.

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For computerization of the entire mining operations through an integrated mining system, many other pieces of production equipment in the mine, such as shovels, dumpers, wheel loaders, draglines, dozers and other blasthole drills must be equipped with systems similar to those elaborated through Table 12.1. Very specialized software must be used to coordinate and control all this equipment. This software needs improvements and modification from time to time.

12.4

DRILLING KNOWLEDGE SYSTEMS

With blasthole drilling being a part of the production process, attempts at introducing speed and automation were made right from the early days. The first measure adopted was the use of compressed air as a flushing medium rather than water or mud. The next measure was of introducing a pipe changer in a rotary blasthole drill. With this, the operation of adding and removing the drill pipes to the drill string was made semi-automatic, easy and quick. Right from the inception of drilling, the importance of knowledge about the formations encountered while drilling was recognized. Drillers always carefully watched the cuttings ejected with the flushing fluid and correlated the observations with the properties of the formation according to their experience. At the same time they also monitored the condition of the drill and drilling operations in general and controlled the feed force, rotary speed etc., on the basis of conditions visualized by them. Very few instruments such as rotary speed gage, ampere meters, air pressure gage, feed force indication gage etc. were available to them.

12.4.1

Formation logging units

Way back in 1911, Conrad Schlumberger used the electric resistivity principle for logging the formation properties in one of the oilwells. Since then many improvements have taken place in techniques. Logging units that work on different principles have been developed. Some of these techniques are electric resistivity, neutral gamma, neutron, sonic, temperature, optical, caliper etc. These techniques are used for gathering rock mass data at great depths in the oilwells. Occasionally they have also been used for blasthole logging. Till about the mid 1980s information about the formations encountered in a blasthole did not have the great importance it has today because: 1

2 3 4

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In blastholes the formations encountered were more or less uniform. Their thickness, levels etc. were already defined to the desired accuracy through diamond core drilling. The large and fresh cuttings ejected from the blasthole gave adequate information about the rock mass in which blasthole drilling was being carried out. The techniques of interpretation of such data were not developed to the desired extent. The usefulness of all such data was not known as the blasting technology had not developed to the desired extent.

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5

349

It was thought that finally every cubic meter of the formation will eventually be blasted and the fractured formation will be seen while being loaded in the hauling equipment

As will be explained later, these concepts have now vanished and the importance of knowing formation characteristics through recorders is very aptly understood.

12.4.2

Analog Drilling Recorders

Analog drilling recorders also have their origin in the oilwell drilling industry. In the late 1920s the oilwell drilling industry saw the advantage of weight indicators. Pressure gages were introduced in the early 1930s. Analog recorders for plotting on paper the data received from the various transducers, i.e. sensing devices like the weight indicators, pressure gages etc., were used in the oilwell drilling industry in the early 1960s. Since all the drilling recorders measure different parameters while drilling is being carried out, they are also known as MWD (i.e. Measurement While Drilling) Recorders. As blasthole drilling was a production operation, speed of drilling was considered very important. In the 1950s blasthole drills did have separate gages to show the rotary speed, feed force, compressed air pressure etc. but no instrument to record the information automatically as drilling progressed. In the mid 1960s the recorders developed for oilwell operations were modified for top drive rotary blasthole drills. These blasthole drilling recorders recorded information only about drill operations. The recorders were pen type. They received data from transducers. The arms of the recorder moved as per the magnitudes of the data sensed by a transducer and plotted it on continuous paper that was moved at a constant speed. The data were usually converted into meaningful units shown on the continuous paper before being plotted. For example, hydraulic pressure in the feed force circuit was sensed but the plot showed feed force in terms of thousand pounds. One such early recorder is shown in Figure 12.2.

Figure 12.2 Analog drilling recorder with six arms.

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A part of one such plot is shown in Figure 12.3. With such plots drill supervisors could not only keep control of the drilling operations but could also advise the drillers about optimizing the drilling operations. Penetration rate gave some indication about blastability of the formation.

12.4.3

Electronic drilling recorders

With the advent of microprocessors in the late 1970s electronic devices made a quantum leap. The data gathered by drilling recorders could easily be converted into digital signals with the help of analog-to-digital converters and microprocessors. The advantage of such conversion was that data could be stored in digital form on magnetic storage devices like disks and could be viewed by rendering it on a display device like a monitor or could be plotted on paper. It could also be sent to a remote place by wired or wireless communication at any time without losing any accuracy. One of the first digital drilling recorders was installed on a rotary blasthole drill in 1982. It stored data in the memory of the device. Since then digital drill recording devices have improved tremendously. Many manufacturers now make digital drill recording devices. They are also known as ‘Drill Monitors’ or ‘Drill Efficiency Indicators’. A typical log recorded by a digital drill recorder is shown in Figure 12.4. Once these types of drilling recorders were found to be very satisfactory and the data retrieved from them was very helpful in planning the project very efficiently, it was visualized in the early 1990s that complete mining operations could be made automatic. For this purpose, systems and devices to achieve similar objectives were developed for some other machines used in mining, e.g. shovels, draglines, dozers etc. It was found that the tramming operation of all such equipment could be made precise and automatic by use of GPS that was developed for rocket launching and monitoring. Automatic leveling systems, through printed logic circuits, were already in use in blasthole drills. Drilling and even adding or removing drill pipes could be made automatic by incorporating certain devices in the blasthole drill.

Rate of Rotary Speed Penetration

Weight on the Bit

Drill Movement

Figure 12.3 Part of a plot from blasthole drilling recorder.

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Instantaneous Penetration Rate in m/h

Compressed Air Pressure in bars

Rotary Speed in rev/min

351

Rate of Drill Head Advance in m/s

Figure 12.4 Drilling parameter record by digital recorder.

12.4.4

Parameters of drilling knowledge data

To introduce automation in the blasthole drilling process, many parameters are required to be known about the working of the blasthole drill. These parameters can be divided into two categories viz. parameters associated with the material being drilled and parameters related to the position of the blasthole. Table 12.2 lists the parameters associated with the material being drilled. Table 12.3 lists the parameters related to the position of the blasthole. Measuring these parameters is not enough. To achieve the ultimate objective of automatic control of the mining process, lot of calculations are required to be done on such data.

12.5

POSITIONING BY GPS

GPS stands for Global Positioning System. GPS works by combining three segments viz. satellites in space, control stations on the ground that ensure precise positioning of the satellites with respect to the center of the earth, and the information fed by the user. The satellites available for positioning an object like blasthole drills are either from a US system called GPS or a Russian system called GLONASS i.e. Global Orbiting Navigation Satellite System. By using specialized components it is also possible to simultaneously use both the satellite and the constellations and get better accuracy.

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Table 12.2 Drilling parameters required to be known. Drilling data Item

Symbol

Time Counter Desired Blasthole Depth Actual Drilled Depth

T L L1

Depth Remaining to be Drilled Cumulative Depth Drilled in a Shift.

L2 LC

Instantaneous Penetration Rate Overall Penetration Rate for the Blasthole Overall Penetration Rate for the Shift

Pi Po

Feed Force Exerted

F

Weight on the Drill Bit

W

Rotary Speed

R

Rotary Torque

T

Hor. Vibration Level at the Drill Head

VH

Vert. Vibration Level at the Drill Head

VV

PS

Pressure of Comp. Air at the Drill Head PRC Pressure of Comp. Air at the Drill Bit

PRB

Dependence and ways of determination of the parameter Computer’s Time Counter Predefined Length Distance from a Predefined Position of the Rotary Head + Length of Drill Pipes Added to the Drill String L2 = L − L1 Sum of final values of L1 for all the blastholes drilled in the Shift. Pi = ΔL1/ΔT Po = L1/TH, Where TH is the Time Required for Drilling the Length of the Blasthole PS = LC/TS,Where TS is the Time Required for Drilling the Length of all the Blastholes in the Shift Function of Pressure in the Feed force Hydraulic Circuit or the Current Drawn by the Electric Feed Motors or a Mechanical Force Measuring Device W = F + WH, Where WH is the Weight of the Drill Head and the Drill String Components Can be Measured by a Revolution Counter and Computer’s Time Counter Function of the Hydraulic Pressure in the Rotary Hydraulic Circuit or the Current Drawn by the Electric Feed Motors Can be Measured by Using Transducers in Horizontal Direction in Rotary Drill Head. Can be Measured by Using Transducers in Vertical Direction in the Rotary Drill Head. Can be Measured by Using a Pressure Transducers in Compressed Air Flow Path PRB = PRH − PRS, Where PRS is the Precalculated Pressure Loss in the Drill String

Table 12.3 Drill positioning parameters required to be known.

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Drilling data

Dependence on

Time Counter Blasthole Identification Number X ordinate of the Blasthole Position Y ordinate of the Blasthole Position Blasthole Inclination With Vertical Has the Drill been Appropriately Positioned on the Blasthole and Leveled? Has the Drill String Been Completely Withdrawn from the Blasthole and Can it Move Safely? No of Holes to be Drilled

Computer’s Time Counter Predefined Predefined Predefined Predefined The Measurements by the GPS System The Measurements by the GPS System Predefined

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353

The GPS system has 24 satellites in orbits at a distance 20200 km from center of the earth. Each of these satellite is capable of receiving and transmitting signals at different frequencies of radio waves for different communication channels. Barring exceptions, generally at least three satellites of GPS are visible from any point on the earth’s surface. GLONASS visibility is even better than this. Ground control stations continuously monitor the position of the satellites at specific points with respect to the center of earth in the following manner. 1 2 3

Each satellite always remains exactly at the same distance from the center of earth. Each satellite moves around the earth in a very specific direction at the same rotational speed as that of the earth From any given spot on the surface of earth a minimum of four to six satellites are visible so long as the spot is open to sky without any manmade or natural cover

With this, the position of each of the satellites remains unique and unchanged. In other words, if with respect to a point on the ground surface of the earth, a satellite is on a line that is at an angle 57° 31’ 27.819” above the horizon in a vertical direction, at an horizontal angle of 22° 5’ 7.331” measured from east in northerly direction and at a distance 22783.782 m from the point, it will always remain precisely at these angular parameters and distance with respect to that point.

12.5.1

Location recognition by GPS

If the signal sent from a blasthole drill to one satellite indicates that its distance from that satellite is X1, it means that the drill is somewhere on an imaginary surface of a sphere of radius X1 with its center at that satellite. If at exactly the same time the distance of the drill from another satellite is determined to be X2, then the drill is somewhere on a circle that is formed by the intersection of two spherical surfaces, one with radius X1 and the other with radius X2. If at exactly the same time the distance of the drill from a third satellite is determined to be X3, then the drill is at one of the two points where the surface of the sphere of radius X3 with its center at the third satellite intersects the circle cited in last paragraph. Which of the two points is the correct position of the drill can be determined in the same manner by using a distance measurement from a fourth satellite or even by knowing the position of a fixed tower in the vicinity of the drill because the actual position of the one of the two points will usually be at a very short or long distance from such tower. Once the position of the drill can be precisely and unambiguously determined from each of the three or four satellites in the above manner, it can be transformed into latitude, longitude and altitude of the drill by using trigonometrical formulas because the position of each of the satellites is firmly fixed with respect to any point on the surface of the earth.

12.5.2

Movement of drill by using GPS

On a mining bench the position of each of the blastholes is precisely predefined. If such positions of the blastholes with their identifications are fed to the computer placed on the drill, it can be displayed on the screen of the computer. Similarly the position of the blasthole drill itself, as determined by GPS, can also be superimposed on the screen. Both these can be seen on a screen display shown in Figure 12.5.

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Figure 12.5 Blasthole drill navigation aided by GPS.

The sequence of drill movements from position of one blasthole to the other is also well defined. Thus, a blasthole drill can be moved by the driller to the next blasthole position by looking at the screen in the operators cab rather than relying a flag mark made on the ground. As the drill moves the image of the drill on the screen also moves because the satellites are tracking the movement of the drill continuously. When the drill approaches the new blasthole, the screen image automatically gets magnified so the driller can position the drill more precisely. Finally when the drill has moved to the new blasthole within the tolerance parameters predefined for positioning, a signal is generated by the computer informing the driller that the drill has moved to the new blasthole precisely and he can start the next operation in the drilling cycle i.e. leveling the drill. Depending upon the type of equipment used, the accuracy possible with GPS can be as high as 5 cm meaning that the drill can be positioned within 5 cm of the precise position.

12.6

COMPUTERIZED DRILL SYSTEMS

Many computerized drill systems are available from many manufacturers. Components, specifications and accuracy of the digital drill recording system differs from manufacturer to manufacturer.

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355

A complete computerized drill system used on a blasthole drill comes with various degrees of sophistication. By and large the complete computerization of a rotary blasthole drill can comprise a drilling knowledge system, an automatic leveling system, an automatic drilling system and GPS positioning system. Further the complete system can be integrated with an integrated mining system, which not only controls the blasthole drill operations but also the shovel and truck or dragline control system. Items needed for computerization of the blasthole drill to various degrees of sophistication have already been listed in Table 12.1. The position of the items needed for a drilling knowledge system is given in Table 12.4. The main computer needed for running the programs of drill knowledge, drill automation, GPS and integrated mining system is placed in the operator’s cab of the rotary blasthole drill and works in collaboration with the main computer usually located in the mine office. The position of the GPS components that go in the drill is shown in Figure 12.6.

12.6.1

Measurements while drilling

Since the data required by a drilling knowledge system is acquired while the drilling operation is being carried out, it is often called ‘Measurement While Drilling’ and abbreviated as MWD. The MWD module can be incorporated in a blasthole drill to store the data measured while blastholes are being drilled.

Table 12.4 Tabular form of MWD data. Drill no.

Date and time start

Blasthole Rot. Weight on Rate of Air Vibration Blastability depth speed the bit Torque penetration pressure amplitude index

22

8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs 8/25/2006 15.24 Hrs

0.2

27.4

136.37

19.2

17.2

192.2

0.46

29.8

0.4

32.3

45.15

48.9

17.7

315.4

0.67

18.9

0.6

49.1

41.11

53.3

17.4

316.5

1.04

26.8

0.8

46.6

79.03

74.2

12.5

318

1.04

47

1.0

46.4

122.5

88.3

6

318.9

1.1

83.8

1.2

45.9

137.26

84.7

3.5

319.4

0.91

176

1.4

45.6

145.99

73.8

2

319.9

1.84

278.9

1.6

46.3

140.05

72.1

2.9

319.6

0.9

183.1

1.8

52.8

147.02

79.6

4

319.3

0.8

112.8

2.0

63.8

381.48

112.5

10.4

319.1

1.09

45.5

22 22 22 22 22 22 22 22 22

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GPS Antenna No 2

Display Screen and Handheld Input Terminal in front of the Operator

GPS Antenna No 1

Drill Head Position Sensor Radio Antenna

GPS Receiver, Radio Transreceiver and Drill Level Indicator behind the Operator

Figure 12.6 Position of GPS components in a blasthole drill.

A few parameters in the drilling knowledge system are basic in nature. They are either natural or are operator-controlled and are called independent parameters. Some other parameters are calculated from these independent parameters. 12.6.1.1

Independent parameters

The independent parameters are listed in Table 12.1 and elaborated below. 12.6.1.1.1 Time This basic element is essential in most of the computations. In drilling knowledge systems, the accuracy of the clock need not be more than a hundredth of a second.

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Hence, usually the clock of the computer used in the system is sufficient. Measurement of time naturally includes date, as date is a designation given to a certain time interval of which second is the basic unit of measurement. 12.6.1.1.2

Depth

In drill knowledge systems, the depth is always known by measuring the travel of the drill head. Naturally, it is the length of the blasthole and not vertical depth. In most of the cases the depth measurement accuracy of 1 in 100 is sufficient. To actually measure the drilled movement, an optical encoder is fixed near the sheave, over which the chain or the rope that shifts the drill head travels. The optical encoder senses the rotational movement of the sheaves or shaft and converts it into the distance traveled by the drill head. In the drills, where the movement of the drill head is achieved by the rack and pinion system, a separate cable is attached to the drill head with an additional sheave at the top of the mast. The other end of the cable is in the reel with auto-winding arrangements in it. By reading the angle through which the reel has rotated and the winding radius of the reel, exact measurement of drill head travel can be taken. 12.6.1.1.3

Rotary speed

The rotary speed of the drill string can be accurately measured by an optical encoder fixed on the drill head or by measuring the rotary speed of the hydraulic or electric motor used to drive the drill head and using the gear reduction ratio of the drill head in the calculations. Almost every rotary blasthole drill has a built-in system to measure rotary speed without any drill knowledge system. The relevant measurement signals are simply grabbed by the drill knowledge system and used. Under most circumstances the accuracy of 1 in 200 is sufficient in rotary speed measurement. 12.6.1.1.4 Weight on the bit Depending upon the system of exerting feed force on the drill head, every blasthole drill has a facility of measuring weight on the bit. In computerized systems these measurements can be collected and used for knowing the weight on the bit more precisely by adding the weight of the drill head and other drill string components in the drill string. Under most circumstances the accuracy of 1 in 1000 is sufficient in bit load measurement. 12.6.1.1.5 Torque Torque is the force required for rotating the drill bit and the drill string. Torque is a good indicator of rate of penetration and hardness of rock. 12.6.1.1.6 Vibrations Vibrations experienced by the drill string are eventually transmitted to the drill mast through the drill head. A vibration sensor mounted on the drill mast can easily give

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an idea of the very rough treatment experienced by the mast. A vibration sensor is required to measure both the amplitude and frequency of the vibrations in lateral, longitudinal as well as vertical directions. Vibration sensors are of particular importance from the viewpoint of automated drilling systems for the reason that vibrations reduce the fatigue life of the drill head and mast. Theoretically, in an extremely rare situation, the vibrational frequency of the drill string can match with that of the mast, and resonance can occur. In such an event the amplitude of the vibration increases to such an extent that it instantly causes heavy damage to the mast. Therefore, in automated drill systems the rotary speed and feed force are adjusted to reduce the amplitude and frequency of the vibration when these parameters start exceeding a set limit. 12.6.1.1.7

Compressed air pressure

This parameter is invariably measured and shown to the operator by a pressure gage in the operator’s cab. For introducing compressed air pressure in a drilling knowledge system, a suitable pressure sensor is place in the flow conduit. 12.6.1.2

Calculated parameters

Some parameters that have great relevance to the productivity and economics of rotary blasthole drilling and subsequent mining operations, have to be calculated form the independent parameters. 12.6.1.2.1

Rate of penetration

Rate of penetration is probably the most important parameter from the viewpoint of rotary blasthole drilling. Actually it is calculated by dividing the distance that the drill head travels while drilling is carried out, by the time spent for travelling the distance. As stated in Table 12.2, the instantaneous penetration rate is Pi = ΔL1/ΔT. In most of the drilling knowledge systems the distance interval ΔL1 can be set to a fixed distance and the average penetration rate for that small interval is calculated. 12.6.1.2.2

Blastability index

Blastability index is the most important parameter from the viewpoint of post-drilling operations such as blasting, loading, hauling, crushing etc. Unlike rate of penetration, blastability index cannot be directly calculated from the data acquired through the drilling knowledge system alone, but requires great calibration. Some diamond core drill holes have to be drilled at the worksite and the cores tested to ascertain the UCS of the formation. The UCS values so determined, and the presence of joints and cracks etc., are initially used for correlating the data acquired in drilling knowledge with a realistic blastability index. The methods and the mathematical formulation involved in establishing the correlation are often proprietary.

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As stated earlier the data retrieved through MWD systems can be viewed graphically or in a tabular form. The tabular form is shown in Table 12.4.

12.6.2

Automation systems

A rotary blasthole drill can be equipped with none, one or more of the four automation systems. They are: 1 2 3 4

Automatic fire extinguishing system Automatic lubrication system Automatic leveling system Automatic drill system

Of these, the first two are totally independent of drill computerization and have been sufficiently elaborated in chapter 7. Even the automatic leveling system can be singularly installed on a rotary blasthole drill. When the drill is computerized the automatic leveling system is included in it. The following can be said about an automatic drilling system. During the use of the drill every assembly in the drill undergoes some normal dysfunctioning. Similarly, every component in the drill string undergoes some normal wear and tear. Almost invariably the overall cost of drilling is minimum when such dysfunction is within a normal range. Three basic parameters that control the rotary blasthole drilling operation are: weight on the bit, rotary speed and compressed air flow through the blasthole. Circumstances that may prove harmful to the blasthole drill or a component in the drill string, or which may stop the process of drilling, can occur when the above three parameters exceed their normal limits. The main objective behind any automatic drilling system is that it should not allow the rotary blasthole drill to work in a manner which may prove harmful to any part of the blasthole drill or any component in the drill string, or stop the drilling process itself. Rotary blasthole drills can be used for drilling in a wide range of parameters and circumstances. The variations can be in respect of: 1 2 3 4 5 6 7

Blasthole diameter Blasthole inclination Drill pipe diameter Depth of blasthole Properties of rock Altitude of operation Water injection etc

Therefore, there cannot be a generalized unique set of ranges for the three controlling parameters. For establishing the normal range, every automatic drilling system has to be calibrated for the particular blasthole drill and the worksite where the drill is expected to work. When calibration is made once in the beginning, it is usually not necessary to do the calibration again at the same site unless the properties of the formation differ drastically.

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Start Engine / Tram to Blasthole Point Start Auto Level System and Level the Blasthole Drill Select the Function Set Up Module Input Torque Limit

Torque Limit Set ?

No

Yes Input Rotary Speed Limit

Rotary Speed Limit Set ?

No

Yes Input Maximum Feed Force

Maximum Feed Force Set ?

No

Yes Input Bit Diameter

Bit Diameter Set ?

No

Yes Input Blasthole Depth

Blasthole Depth Set ?

No

Yes Input Max. Vibration Limit

Max. Vibration Limit Set ?

No

Yes A

Figure 12.7 Flow chart for rotary blasthole drill automation –1.

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A

Input SPD and Torque for Vibration Limit

No

SPD and Torque for Vibration Limit Set ? Yes Input Feed Force and SPD for Collar Mode

No

Feed Force and SPD for Collar Mode Set ? Yes Input All the Remaining Parameters

Have All Other Parameters Been Set ?

No

Yes Start Drilling / Activate New Hole Key Establish Zero Reference Depth Enter ID Information

Activate Drilling Key

Correct Conditions

Monitor Interlocks

Check User Display

Are All Interlocks OK?

No

Actual Drilling Started Monitor All Desired Parameters Display Parameters at User's Request

Yes

Desired Depth Reached? No Has User Stopped Drilling?

No

End Drilling and Carry Out All Processes to Proceed to Next Blast hole

Figure 12.7 Flow chart for rotary blasthole drill automation – 2.

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A process flow chart that controls the blasthole drill automation is shown in Figure 12.7. The flow chart is for a blasthole drill having a capability of drilling vertical blastholes with only one drill pipe. These flow charts are based on the US Patent assigned to Reedrill Inc. USA. Wherever required, more sophisticated programs can be made and used.

12.7

HARDWARE FOR DRILL COMPUTERIZATION

The hardware items required on a blasthole drill for its computerization have been shown in Figure 12.1. To make the matter simple, the following is a list. 1 2 3 4 5 6 7 8 9 10

System Unit Display Monitor Input Device Radio Transmitter GPS Antenna GPS Receiver Amplifier Different Types of Transducers Power Supply Unit Software Details of some of the above devices are given below.

12.7.1

System unit

In its architecture and operation the system unit used for drill computerization is very similar to the system unit of a present day personal computer. In fact it is certainly possible to use the system unit of a personal computer for the computerization of a rotary blasthole drill but it is not done so for commercial as well as technical reasons. The system units used by different manufacturers differ considerably and their composition also changes from time to time as new technologies emerge in the field of computers. The following is a rough idea. The system unit has a specialized motherboard and one or more special purpose processors on it. ROM is provided on the motherboard for storing permanent operating instructions and RAM is provided for storing temporary data. A hard disk is also used for mass storage of the data. The motherboard has several communication ports and a system bus to accept data from or to send data to various other components like electro-hydraulic control valves, radio transreceiver, GPS antenna, transducers. The processor in the system unit has to be as fast as possible because it has to work in real time. It means that the processor should be capable of analyzing all the signals and finish all the calculations without making any device wait for giving the data to the processor. A typical system unit and display monitor of the computerized drill system looks like the one shown in Figure 12.8.

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Figure 12.8 System unit and display monitor of a computer used for blasthole drill computerization.

12.7.2

Display monitor

Since the display screen in the operator’s cab has to be clearly visible during bright daylight, it is a color plasma display type. Such a unit also occupies much less space in the operator’s cab and hence can be placed at a convenient position. Since it is meant to be seen only by the operator in the cab, its size is small. Usually 200 × 250 mm is sufficient. Some manufacturers provide touch screens for display. These screens allow the selection of a choice or a region on the screen by finger touch rather than a mouse click as is used in commonly seen personal computers. Display units are made robust to withstand the harsh environmental conditions prevailing on the site. The actual parameters to be seen on the display are chosen by the drill operator and depend upon the operation being carried out. For example, while changing a component in the drill string, the display chosen by the operator will be as shown in Figure 12.9, or while actual drilling operation is being carried out the display chosen by the operator will be as shown in Figure 12.10.

12.7.3

Input device

At many instances data have to be input to the computerized drill system e.g. coordinates of each blasthole, blasthole number, drill string component number, desired blasthole depth etc.

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Figure 12.9 Display while changing a drill string component.

Since the display monitors chosen by manufacturers are touch screen type, they can also act as input devices. However, in the absence of such touch screens an input device much like a keyboard of a personal computer becomes essential. Handheld devices, however, are rapidly vanishing because they need additional storage space and in any case add to the number of components in the system.

12.7.4

Software

The computerized system on a blasthole drill cannot work without software. The computers located in the mine office have software running on commonly used operating systems like MS Windows. But as the components used on rotary blasthole drills are special purpose, the commonly available operating system such as Microsoft Windows or Linux etc., cannot be used. The system unit in the operator’s cab is equipped with tailor made special software. Since the modality of tasks to be carried out is very limited the software does

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Figure 12.10 Display while actual drilling is in progress.

not have two components e.g. an operating system and software to run with the operating system. The composite software is usually stored in some permanent storage device such as ROM or hard disk. The software allows: 1 2 3 4 5

Receiving input from touch screens Receiving input from transducers of the drilling knowledge system. Communication with display unit Communication through radio transreceiver Communication through GPS receiver and GPS antenna Figure 12.1 clearly shows all these aspects.

12.8

ADVANTAGES OF DRILL COMPUTERIZATION

Drill computerization has been described from various angles in the forgoing part of this chapter. It is now necessary to take a look at the advantages offered by computerization. Before proceeding with the topic it must be clearly understood that a drill computerization system, or any component thereof, is not made for replacing respective controlling persons but to enable them to be more efficient and cost-effective in their work.

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12.8.1

Advantages of drilling knowledge system

This system enables retrieval of vital data about the formation while drilling is carried out. Further, the data are stored in the system. Review of the stored data can be beneficial for the reasons given below. A drilling knowledge system keeps an indirect watch on the driller and helper and their work. Therefore, the drilling supervisor needs to pay less attention to the drilling operation. Obviously, a supervisor can control more blasthole drills or other machines depending upon his aptitude. Situations where the skill of the driller and or helper is less than desired, can be picked up easily by reviewing the data stored. Proper advice can be given to the driller and/or helper to be better at their job. Other reasons of low productivity can be easily known and corrective action can be taken. Review of the stored data by a researcher can give many clues about the correlation of different parameters in rotary drilling operations. For example Figure 12.11 shows the correlation between penetration rate, torque and vibration when the bit load and rotary speed were kept constant. Such studies prove very useful in designing equipment and more importantly in designing a blasting program.

Torque N.cm

Vibration Nm/s

Blasthole Depth m

Penetration Rate cm/min

Rotary Speed RPM

Blasthole Depth m

Bit Load (N/100)

Figure 12.11 Correlation between drilling parameters.

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The penetration rate achieved in drilling is a very important parameter. If other drilling parameters like bit load and torque used for drilling are not changed, the penetration rate achieved is proportional to the strength of the formation. On the basis of instantaneous penetration rate recorded by the computerized system, inferences can be drawn from the variations of strength of the formation. Figure 12.12 shows the presence of a hard rock layer and soft bed from the penetration rate. Since the penetration rate, torque, bit load and rotary speed are continuously measured by the computerized system while drilling is being carried out by a blasthole drill, it can also give a graph of indices that are based on these measurements. Three indices as stated below are commonly recognized: 1 2 3

Index of Rotary Energy Formation Alternation Index Ground Strength Index

These indices are calculated by simple formulae. The formula for the index of rotary energy is: IE = (T*N)/P

Medium Rock Hard Rock

Medium Rock

Soft Sand Layer

Medium Rock

0 10 20 30 40 50 60 70

Rock Profile in Blasthole

Penetration Rate m/h

Figure 12.12 Correlation between penetration rate and UCS.

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where IE T N P

= = = =

Index of rotary energy Torque used for drilling Rotary speed used for drilling Penetration rate achieved in drilling

The formula for formation alternation index is: IA = 1 + (E/Ea) − (P/Pa) where IA E P Ea

= = = =

Index of rotary energy Bit load used for drilling Penetration rate achieved in drilling Maximum bit load used during drilling of the blasthole or perhaps all the blastholes in a bench Pa = Maximum penetration rate achieved in drilling of the blasthole or perhaps all the blastholes in a bench

The formula for the ground strength index is: IR = (E*N)/P where IR E N P

= = = =

Index of ground strength Bit load used for drilling Rotary speed used for drilling Penetration rate achieved in drilling

If the properties of the overburden and ore are drastically different, the penetration rate or other similar parameters change at the dividing plane. Where the situation warrants drilling activity can be easily stopped at such a plane by the driller. Since the drilling parameters are visible on the screen in the operator’s cab, drilling operations can be continued even if the mast and drill pipe are not visible because all the inferences that a driller draws by viewing the mast, drill pipe etc. are displayed automatically and precisely on a display monitor in the cab. Since a drill knowledge system also keeps records of the utilization of accessories, the procurement of accessories can be planned in a better way. This results in reduced inventory and associated cost savings. A proper video shooting of the drilling operation from the driller’s cab along with precise sound recording and simultaneous shooting of the screen of the drill knowledge system can be very useful in imparting training to newly-recruited operators. Even if the blasthole drill is equipped with only a drill knowledge system, the information gathered by it is very useful in designing the blasting plan to achieve optimum fragmentation, because the explosive quantity and its placement can be determined for each blasthole individually.

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12.8.2

369

Advantages of drill automation system

It has been experienced that a drill automation system carries out the drilling operations faster than most of the operators while remaining nearer to the best quality of operation. Naturally, with the drill automation system, production in terms of m/day is usually more so the cost is reduced. A drill automation system reduces downtime to a recognizable extent. The resulting better utilization gives higher production in the same time and naturally ends in reducing the cost of operation. Since an automation system ensures that the drill is used within an acceptable range of parameters, the wear and tear of the drill itself is less and the drill life is enhanced. The direct impact of this is in terms of reduction in maintenance cost. Indirectly this ends in reducing owning cost. For the same reason given in the above paragraph, the components in the drill string are also used in an appropriate manner to give longer life in terms of meterage (i.e. footage). This factor, too, results in reduced cost. Nowadays, with the proliferation of computers and information technology, it has become difficult to get persons of high aptitude for an unattractive job such as a driller. Chosen persons need more training. In such situations drill automation system augments the skill of such ‘yet semiskilled’ drillers greatly.

12.8.3

Advantages of GPS positioning system

The benefits of a GPS positioning system are many in number and vital in nature. GPS positioning is done automatically i.e. without any surveying and leveling activity. The coordinates of the desired locations of the blastholes have to be fed to the computer only once. Without GPS positioning, marking the blasthole collaring points on the ground at the coordinates has to be done through work of a surveyor and two helpers. GPS positioning not only eliminates the need of these employees but can do a better job. The points need not be visibly marked on the ground but the GPS system knows the position of each point very precisely and the blasthole drill can be navigated to the point either by the driller with visible help from the GPS positioning system or even automatically. The GPS positioning described earlier is to be carried out by the operator looking at the screen only i.e. without looking out for the marking on the ground, hence the navigation can be carried out even in the night or under very low visibility. A GPS positioning system attains very high accuracy, even of the order of 10 mm with the help of GLONASS. For this reason the actual blasthole collaring point is very accurately positioned near the precisely intended location. In one study the precision of the blasthole location with and without GPS was measured, and the observations were plotted. The results were as shown in Figure 12.13. If blastholes are not drilled at their precise positions and blasting is carried out, then depending upon how blastholes are positioned, some part of the bench does not fragment properly. This leaves very large rock fragments that have to be refragmented by use of an hydraulic hammer like the one shown in Figure 1.11. This is a wasteful expense.

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Figure 12.13 Precise positioning (Right) attained by GPS system.

A bench blast is normally designed with the assumption that the blastholes are at their precise locations. The blast can be designed in such a way that the throw of rocks from the bench forms a heap with such a profile that the shovels can lift the rock very efficiently. Large deviation of blastholes from their precise position does not result in the formation of such a heap. Thus higher shovel utilization is needed which also is a wasteful expense. Deviation in the position of blastholes from their precise position also results in flyrock hazard. Before starting blasthole drilling on any bench, a dozer has to level the top of the bench. This is done for two reasons. One is that on such leveled ground a relatively unstable blasthole drill can tram without any danger. The second is that from such leveled ground, if all the holes are drilled to same depth, their bottom is also in a plane surface in the same level. A GPS positioning system not only knows the blasthole position in terms of horizontal X and horizontal Y directions but also in the vertical Z direction. Therefore it is capable of knowing the reduced level of the collaring point of each of the blastholes. Using this information each of the blastholes can be given a separate depth so as to ensure that their bottoms are at the same plane surface in the same level. If blastholes of such different depth are properly loaded and blasted, the bench top, formed after removal of blasted material, is in level and the leveling efforts by use of a dozer are reduced. Another important advantage of a GPS positioning system is that data on each blasthole location as gathered by it can be used in the mine office to design the blast very effectively. In such blast design it is possible to ensure that the explosive

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concentration in each blasthole is calculated separately so more explosive is used in the layers where the formation is more difficult to blast.

12.8.4

Advantages of integrated mining system

As the name implies integrated mining attempts to take into consideration the working of all the major pieces of production equipment such as blasthole drills, loading shovels, dumpers, draglines, dozers, graders etc. With this the whole mining system can be made efficient by scheduling different activities in an appropriate manner. All this results in enormous cost saving and easily offsets the investment in the integrated mining system. A generalized idea about the reasons for cost savings in respect of different equipment can be gained through Table 12.5. The estimated cost saving in Aus$ is presented in Table 12.6. Both these tables are based on experience gained in the St. Ives Mine in Australia and are based on an article by Andrew P. Jarosz and Raleigh Finlayson. Table 12.5 Potential for benefits and cost saving through GPS. GPs guidance and positioning used for Reason for benefits and cost saving Reduce misidentified loads* Correct floor elevation* Increased Productivity Reduce idle time Improved mine planning Mine to design Mining subgrade blocks No dig lines Edge of blast information Reduced survey requirement 24/7 survey assistance* Working around voids* Reduce floor dilution & ore loss* Establish dropcut on grade Flatten heave Waste removal on ore zones Waste dump construction Backfilling pits Material tracking* Accurate overhaul cost Tonnage/truck determination Reduced idle time Improved fragmentation* Reduced labor cost

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Excavator / shovel Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes

Dozer

Grader

Yes Yes

Yes Yes Yes

Haul truck

Yes Yes Yes

Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes Yes

Yes Yes Yes

Yes

Yes

Yes

Yes

Yes

Drill and blast

Yes Yes Yes Yes

Yes Yes

Yes Yes Yes

Yes Yes Yes Yes Yes Yes

Yes Yes

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Table 12.6 Potential cost saving in Australian $. Machine using GPS Excavator or Shovel Dozer

Cost saving reason

Monthly savings A$

Annual savings A$

Cons. %

Adjusted annual savings A$

Mis identified loads

297123

3565476

50%

1782738

93150 19000 4138 19244 10491 90772 45221 17603 11250 6667 614658

1117800 228000 49650 230929 125892 1089264 542652 211236 135000 80000 7375899

50% 25% 50% 50% 25% 50% 75% 50% 25% 25% 43%

558900 57000 24825 115465 31473 544632 406989 105618 33750 20000 3681390

Mine to correct floor elev. Increased productivity Mining to design Doze to design floor level Doze/Grade within orebody Material tracking Haul Trucks Accurate overhaul costs Improve fragmentation Drill & Blast Minimal survey requirement Minimal dipping requirement Total estimated savings

12.8.5

Advantage of added safety

Apart from the above advantages the drill automation system inherently incorporates several safety features in the drill operation. Some of them have been described but are repeated below: 12.8.5.1

Blasthole depth limiting input

When a value of limiting depth of a blasthole is fed to the computer, the system automatically stops drilling when the depth is reached and withdraws the drill string completely. 12.8.5.2

Tram interlock for pipe in hole

The system allows actuation of tramming motors only when the drill pipe and drill bit has been completely removed from the blasthole. 12.8.5.3

Centralizer damage prevention

The centralizer, also called rod support or gripper arm, on the blasthole drill that is positioned in the middle of the mast height and supports long drill pipes, obstructs the drill head advance. It must be taken to its stowed position in order to continue further drilling. The system automatically actuates the hydraulic cylinders that take the centralizer into stowed position when the drill head is slightly above it.

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12.8.5.4

373

Leveling jack interlock

After the drill is leveled the system actuates valves on the leveling jacks in such a way that, even if the hose or metal pipes supplying or taking away the hydraulic oil to or from the jacks are punctured, the oil in the hydraulic jacks does not leak and the drill remains in level position 12.8.5.5

Carousel damage prevention

The drill head or carousel can get damaged if the drill head is moved down when the carousel is in the path of drill head during feed movement. The system does not allow the feed movement of the drill head when the carousel is not taken to its stowed position. 12.8.5.6

Breakout wrench damage prevention

Like the centralizer or carousel, the breakout wrench also operates in the feed path of the drill head and can get damaged if the drill head is allowed to move and collide with it. The system prevents the feed movement of the drill head if the breakout wrench is not in its stowed position. Apart from the above, a lot of information about working of the engine or electric motor and compressor is shown by the system to the driller on the screen of the computer. Such display is kept under view by the driller or his assistant and appropriate corrective action is taken whenever needed.

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Chapter 13

Concepts for rotary blasthole drill design

13.1

INTRODUCTION

Like any other product, a rotary blasthole drill is required to be designed meticulously. Maximum operational versatility and maximum production at minimum operational cost are the main objectives behind its design. A rotary blasthole drill is a self-contained machine mainly used for drilling shallow large diameter holes in the ground in a vertical or somewhat inclined downward direction. Quite a bit of general information about rotary blasthole drills has been elaborated in the earlier chapters, particularly chapter 7, of this book. However, some vital aspects related to rotary drill design must be known to the designers as a very first step. It is very important to understand these concepts even if the reader is not likely to design a rotary blasthole drill, because such understanding will give a deeper insight into the subject and will prove to be of great help even to drillers and drilling supervisors for the planning and execution of their jobs. This chapter is not devoted to design of blasthole drills but only to look into the special requirements of a rotary drill from the viewpoint of the drilling process. From this angle many design tips have been mentioned in the forthcoming sections of this chapter.

13.2

PRIMARY REQUIREMENTS OF THE DRILL

Ever since the inception of the rotary blasthole drill in the late 1940s, their appearance has remained so similar that all rotary blasthole drills look alike. In this manner, since the form of the rotary blasthole drill is already well defined, the first step i.e. visualization of the end product, has become unnecessary unless one ventures into designing a rotary blasthole drill on the basis of totally different concepts of rock fragmentation. The basic requirements of any drilling process have been stated in the second chapter of this book. Of those requirements, a rotary blasthole drill must be able to fulfill following: 1 2

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Supplying energy to the drilling bit so it is able to fracture the formation. Supplying an adequate volume of the circulating fluid so that the cuttings are removed from the hole.

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3 Rotating the drill bit to ensure formation fracture across the cross section of the hole. 4 Exerting sufficient feed force to give adequate bit load for effectively fragmenting the formation. 5 Imparting linear movement in the forward direction to the drill bit through the drill string so the hole progresses. 6 Exerting sufficient pullout force to ensure that the drill string used for drilling to maximum blasthole depth can be lifted with all the heavy components and has sufficient reserve to cope with the situation when it gets stuck. 7 Imparting linear movement in the backward direction to the drill bit through the drill string so the drill can move to the next hole location. 8 Smoothly moving to the location of the next blasthole with adequate stability. 9 Positioning the drill with additional stability to cope with the reaction of the feed force exerted on the drill string. 10 Collecting very fine particles of the formation formed in the drilling process so as to prevent them from mixing with the atmosphere. 11 Reducing the noise generated in the drill so as to reduce noise pollution. All the above factors must be constantly kept in mind while one proceeds with the design of a blasthole drill.

13.3

SIZE AND WEIGHT OF THE DRILL

A mine is a privately owned area. The weight and size restrictions applicable for a vehicle to be travelworthy on public roads are not applicable to the vehicles that are used exclusively in mines. A rotary blasthole drill works exclusively in mines. It can therefore be stated that there are no particular restrictions in respect of its length, height, width and weight. As far as size is concerned, a drill can have length as much as 30 m, width as much as 10 m and height as much as 30 m. Unless some overriding considerations exist, which can happen in extremely rare instances, the undercarriage for a rotary blasthole drill must consist of crawler tracks. For traveling from one bench to another bench the drill travels on the slope, and the mast of the rotary blasthole drill must be in the lowered position so as to keep the center of gravity at the lowest level and thereby attain higher stability. Since the roads in a mine are designed with a maximum of 10% gradient, with a factor of safety of 2, a rotary blasthole drill must be able to travel up as well as travel down a slope of say 20%. This concept is very important not only from the viewpoint of engine power but also the drill must be sufficiently stable under these conditions. Naturally the length of the crawler track, and a to good extent the width between crawler tracks, is determined on this basis. Remember that a drill can travel up a 20% grade with sufficient stability but will topple even on a 15% grade while traveling down. As a rotary blasthole drill while working on a bench has to travel only say 10 to 20m while moving from one bench to another, the travel speed should be limited to only

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2 to 3 kmph. Higher speeds are of no use and in fact can prove counter-productive because accurate positioning of the drill on the exact location of the blasthole can be attained only at low travel speed. Since a rotary blasthole drill works on the top of a bench where the surface is not paved, the pressure at the contact between ground and crawler should not exceed 100 kPa. Where the mine site experiences heavy rainfall the ground can become slippery and the crawler track may be susceptible to slight sinking in the ground. In such circumstances the contact pressure should not exceed 70 kPa. With both the above concepts the width of each crawler track can be determined. As far as weight of the drill is concerned it has an implicit relation with the maximum blasthole diameter. Let us presume that the maximum bit load to be exerted by the rotary blasthole drill is X. When the blasthole drill is engaged in drilling an inclined blasthole at an angle of inclination of 30°, the vertical component of X, i.e. Xv, is 0.866X and the horizontal component of X, i.e. Xh, is 0.5X. When no bit load is being exerted, the total weight of the rotary blasthole drill is counter-balanced by the reaction W1 on the rear side and W2 on the front side. At this juncture one must understand that bit load includes the weight of the rotary head as well as the axial force exerted by the feed mechanism provided on the rotary blasthole drill. All these have been shown in Figure 13.1. In most rotary blasthole drills the axis of vertical blasthole position is within the rectangle formed by four leveling jacks. In many cases it is so near to the vertical plane of the rear jacks that it is safe to say that the reactive force X resulting from the feed force X is totally borne by the weight reaction W1 on the rear side. Needless to say that W1 should be more than X. Although there is no such rule, in order to account for inaccuracies in fabrication weighing and other uncertain factors, W1 should be equal to 1.2X. In such cases when levelling, none of the jacks will get lifted and the drill will not tilt to one side. For all practical purposes, rotary blasthole drills are designed in such a way that W1 equals to 0.5 W to 0.6 W. Therefore, W should be 2 to 2.4 times the maximum bit load capacity X of the drill. Another consideration in this regard is that while drilling inclined blastholes the resistance to the horizontal component of the feed force should be sufficiently large. Here again, there is no specific rule, but it can be said that in the rainy season the coefficient of friction between the pads of the leveling jacks and the ground can be as little as 0.4. Thus to ensure that the blasthole drill does not slip to the front side, the load exerted by the lifting jacks on the ground must be sufficient to resist the horizontal component Xh of the bit load reaction. For a 30° inclined blasthole Xh works out to 0.5X. In view of the above points we have: Xh = 0.4(W − Xv) or 0.5X = 0.4(W − 0.866X) i.e. W = 0.5X/0.4 + 0.866X = 2.116X

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W X

W1

Figure 13.1

W2

Determining desirable weight of the drill.

From all that has been said in this section, the weight of the rotary blasthole drill should be more than 2.116 to 2.4 times the maximum feed force it is expected to exert while drilling. The maximum bit load required to be exerted on a tricone bit for drilling in the extremely hard formations, as stated in Table 5.6, is 1.575 kN/mm. This parameter is somewhat arbitrary and is based on the recommendations of the tricone bit manufacturers. In actual practice, in almost all the instances, it is sufficient to exert a bit load of about 1.350 kN/mm. Thus, if a rotary blasthole drill is to be designed to drill a blasthole of diameter 311 mm, it should be capable of exerting 1.350 * 311 = 419.850 kN bit load. In more familiar unit it works out to 42813 kg. and thus the weight of the rotary blasthole drill should lie somewhere between 90592 kg to 102751 kg.

13.4

ROTARY SPEED AND TORQUE

At any given bit load, when the rotational speed of the bit is increased, a greater number of teeth penetrate the ground and more formation is fractured. This results

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in a faster penetration rate with increasing rotary speed. However, with higher rotary speed the instances of impacts of the teeth with the formation increase. The quantity of heat generated also increases to unacceptable levels. This is added to further by the vibrations at the bit induced by the vibration of the drill string. All these factors put a limit on the maximum rotary speed, beyond which the performance diminishes. In softer formations higher rotary speeds can be tolerated by the tricone bits. In hard formations the rotary speed should be lower. Giving due consideration to these factors, bit manufacturers have given recommendations for rotary speeds while using their bit for drilling in formations of different hardness. There is almost a unanimity in these recommendations. They have been summarized in Table 5.6. From these recommendations a matrix of ideal rotary speeds for blastholes of different diameters and for drilling in formations of different hardness are given in Table 13.1 When a blasthole drill is designed to exert a certain bit load, it can become suitable for drilling blastholes of certain diameter in very hard formations, but at the same time it is also equally suitable for drilling larger diameter blastholes in softer formations by using drill pipes of larger diameter. Hence, by carefully scrutinizing such possibilities, the appropriate speed range from the above matrix should be chosen for the drill head. Since tricone bits used in rotary blasthole drilling are thrusted with very heavy bit loads they require very high torque for rotation. Field observations and research in this regard have evolved different formulae for determining maximum torque requirement at the rotary head of a rotary blasthole drill. One empirical formula, called Nelmark’s Formula, states: T = 5252 K D2.5 W1.5 where T = Torque required at drill head (ftlb.) D = Blasthole Diameter (inches) W = Weight on the bit (thousands of lb. per inch of bit diameter) K = A constant related to the properties of the rock Values for the constant K lie between 14 × 10–5 to 4 × 10–5 mainly depending upon the drillability of rock. For hardest rocks, lower values are to be used and for soft rocks higher values are to be used. Table 13.1 Recommended rotary speeds in rpm for blastholes of different diameters in various formations. IADC code for indicating formation hardness Hole dia. in mm

Soft–412

Medium hard–512

Hard–612

Very hard–712

Extremely hard–812

152 250 349 445

150 125 100 80

140 110 90 72

130 100 80 64

120 95 64 52

100 80 60 40

For intermediate diameters values can be interpolated

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As the blastholes are drilled mostly in medium hard to extra hard formations, the high value of K for medium hard formation, estimated at 11 × 10–5, should be used in calculations. Values of W to be used should be 8. The above formula does not take into consideration the additional torque required due to the stabilizer that is often used in blasthole drilling, or the to increased side weight arising out of inclined blastholes. Hence it seems prudent to multiply the torque value by 1.1. On the above basis, calculations for the torque requirement of blasthole drill heads meant to drill holes of size 152 mm, 225 mm, 311 mm and 406 mm, work out as in Table 13.2. Nelmark’s formula is rather crude. It does not correlate the torque requirement to many variable factors which obviously affect the torque requirement. A far more realistic-appearing formula, proposed by Fedorov, based on research in the USSR, is as follows: N = α γ Ld2n1.7 + 0.7854 Nod2 + 8 *10–7Fbn where N = Rotary power for drilling (kW) d = Diameter of the hole (cm.) n = Rotary speed (rpm) γ = Density of flushing medium (kg/m3) L = Depth of hole (m) No = Power spent in fragmenting the bottom cross sectional area of the hole (kW/cm2) Fb = Weight on the bit (Newton) Factor α depends upon the inclination of the hole. Its value can be found from a chart in Figure 13.2. Factor No has a value from 0.1 to 0.15 kW/cm2 depending upon the formation hardness. The lower value applies for soft formations and the higher value applies for hard formations. Using the above formula rotary power can be calculated, and then knowing the rotary speed the torque can also be determined. Table 13.3 gives torque requirements for α = 4.7 * 10–8, γ = 1.2252, L = 30 m, n = 100 rpm, No = 0.135 for bit weight worked out on the same basis as in Table 13.3. Table 13.2 Torque required at drill head for drilling blastholes of different diameters as per Nelmark’s formula. Hole dia.

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Inches

mm

K value

W value

Torque in lbft

Torque in nm

6 9 12 16

152.4 228.6 304.8 406.4

11 11 11 11

8 8 8 8

1268 3177 6521 13386

1719 4307 8841 18149

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381

28

Angle of Inclination in Degrees

24

20

16

12

8

4

0

0

Figure 13.2

1

2 3 Value of α*108

4

5

Determining value of α for different mast inclinations.

Table 13.3 Torque required at drill head for drilling blastholes of different diameters as per Fedorov’s formula. Hole dia.

Bit load

Inches

mm

lb

Newton

Torque in lbft

Torque in Nm

6 9 12 16

152.4 228.6 304.8 406.4

48000 72000 96000 128000

213514 320271 427029 569372

2432 5315 10731 16347

3297 7206 14549 22164

Comparing the torque required as calculated by using Nelmark’s Formula and presented in Table 13.2, and as calculated by using Fedorov’s Formula and presented in Table 13.3, it can be easily seen that Fedorov’s Formula gives a higher torque requirement. Table 13.4 presents torque required for drilling largest blastholes to 30 m depth by using different rotary blasthole drills made by some well known manufacturers. Calculations are made by using Nelmark’s as well as Fedorov’s Formula. For calculations using Nelmark’s formula the value of W is calculated by dividing the maximum bit load by the diameter of the blasthole. As rotary blasthole drills often have to be used in soft formations like coal, the value of K is taken as 13.00. In the case of calculations using Fedorov’s formula, values of a, c, L and No are presumed to be 4.7 * 10–8, 1.2252 and 0.135 respectively.

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Table 13.4 Torque calculations for some crawler mounted rotary blasthole drills.

Maximum recommended hole diameter

Torque required nelmark

Maximum feed force

Calculated Estimated K value Newtons W value lbft

Total rotary power required

Torque required

Hole depth

Bottom power Rotary per speed sq cm

Fedorov

Torque available at Nelmark drillhead

m

rpm

kW

kW

Nm

Nm

Nm

Make

Model name

Inches

cm

lbs

Atlas Copco Atlas Copco Atlas Copco Atlas Copco Bucyrus Bucyrus Bucyrus Bucyrus Bucyrus Bucyrus Hausherr Hausherr NKMZ

PV351

16

40.64

125000 556028

7.813

13.00

15267

41.2

87

0.135

221.59

24322

20699

25761

PV275

10.625 26.9875

75000 333617

7.059

13.00

4712

59.4

150

0.135

129.72

8259

6388

12202

DM50

10

25.4

50000 222411

5.000

13.00

2414

54.86

107

0.135

93.19

8317

3273

10575

17.145

30000 133447

4.444

13.00

757

45.7

100

0.135

43.79

4181

1027

7321

733957 627199 542683 489304 373695 276679 160136 120102 400073

9.429 8.813 8.873 7.333 6.858 5.854 4.800 4.154 8.465

13.00 13.00 13.00 13.00 13.00 13.00 13.00 13.00 13.00

25324 18290 12651 11815 6440 3559 1106 623 6188

79.28 85.37 79.28 85 64 45.7 40 42 32

120 125 160 125 125 220 134 134 150

0.135 0.135 0.135 0.135 0.135 0.135 0.135 0.135 0.135

310.84 267.64 229.89 228.92 153.12 144.31 59.10 44.50 131.95

24736 20446 13721 17488 11697 6264 4212 3171 8400

34335 24798 17152 16020 8732 4825 1500 844 8389

20793 14602 12430 20880 15185 17354 6901 6901 6013

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NKMZ P&H Mining P&H Mining Sandvik Sandvik Sandvik Sandvik Schramm

DM30

6.75

59R 17.5 49R 16 39R 13.75 SK-L 15 SKSS-13 12.25 SKF 10.625 HBM 160 7.5 HBM-120 6.5 SBShS10.625 250/270–32 SBShS-250 N 9.875 320 XPC 17.5

44.45 40.64 34.925 38.1 31.115 26.9875 19.05 16.51 26.9875

165000 141000 122000 110000 84010 62200 36000 27000 89940

25.0825 78680 349986 44.45 150000 667233

7.968 8.571

13.00 13.00

4705 21950

32 50.5

120 101

0.135 0.135

104.27 278.08

8298 26292

6380 29761

6508 33895

250XP-DL

13.75

34.925

105000 467063

7.636

13.00

10101

73.1

100

0.135

179.59

17150

13695

17219

1190E DR460 D75 KS D50 KS T450BH

13.75 12.25 11 9 6.5

34.925 31.115 27.94 22.86 16.51

118000 100000 92000 60000 30000

8.582 8.163 8.364 6.667 4.615

13.00 13.00 13.00 13.00 13.00

12034 8364 6627 2856 729

85 51 53 45 45

97 175 94 126 110

0.135 0.135 0.135 0.135 0.135

184.30 183.42 118.93 87.35 42.73

18144 10009 12082 6620 3710

16315 11340 8986 3872 989

16900 10461 14236 9934 8518

524890 444822 409236 266893 133447


383

The table also presents the actual torque output of each of the rotary blasthole drills. From the comparison of values of torque required as calculated by Nelmark’s as well as Fedorov’s formulae with the actually available torque for a drill as presented in the table, it can be concluded that values of torque as per Fedorov’s formula match more closely with the actual torque output. It is therefore suggested that values calculated by using Fedorov’s formula should be used for actual design. It must be borne in mind that electric motors used to drive the drill heads have very high overloading capacity – often of the order of 1.5 or even more. Hence they can easily generate sufficient torque for drilling holes of maximum-rated diameters even when the table shows that they do not generate sufficient torque.

13.5

FEED FORCE AND SPEED

In rotary drilling with tricone bits, formation fracture takes place along the three radial lines below the axes of the three cones. For this reason the pressure required to be exerted on the drill bit is given as a linear measure i.e. feed force per inch of bit diameter. When the tricone bit is rotated the cones and hence these lines also rotate and cover the complete cross section. Formations to be drilled are often heterogeneous, and cases where a very hard rock layer is encountered in a relatively soft formation are not rare. Thus, the drill should be capable of exerting maximum feed force on the bit so a hole of desired diameter can be drilled even in such a very hard but relatively thin formation layer. Small diameter holes are drilled with drill pipes of small diameter. Such a drill string is slender. To avoid buckling of the drill string and consequent deviation of the hole, lesser feed force and hence lower bit weights are used while designing small rotary blasthole drills. In this context it can be said that capabilities of drills in drilling holes of a particular maximum diameter are based on the feed force ratings given in Table 13.5. In the absence of any governing standard, the feed force ratings mentioned in Table 13.5 are not accepted and adhered to by all the manufacturers. Sometimes it is seen that two blasthole drills have the same bit loading capabilities but are rated to drill blastholes of different diameters. When a blasthole is in the vertical direction, the weight of the drill string and the rotary head is fully complementing the load exerted on the bit. However, when the

Table 13.5 Criteria for selecting the feed force for rotary blasthole drills.

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Hole diameter range of the rotary drill

Maximum desired bit load

Inches

mm

lb/in

kN/mm

6–7.785 8–10.625 10.75–13.75 13.75–17.5

152–200 203–270 273–349 350–445

4000 6000 8000 9000

0.700 1.050 1.400 1.575

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384


blasthole is off the vertical the complementing components reduce in proportion to the cosine of the angle of inclination. If a blasthole drill is capable of exerting the desired feed force but its feed speed is very low, then it will not be able to drill the formation at a penetration rate higher than the feed speed. In rotary drilling the maximum recorded penetration rate, even in soft formations, is about 2 m/min. In hard formations penetration rate is about 0.3 to 1.2 m/min. Hence a blasthole drill should be designed for feed speeds of at least 2.5 m/min, while the maximum feed force is being exerted through the feed mechanism. The mechanism for exerting the feed force should be such that it should be capable of making very fine adjustments in feed speed.

13.6

PULLOUT FORCE AND SPEED

After completion of drilling the drillstring is required to be withdrawn. This requires lifting the weight of the drill string and rotary head by the feed mechanism. The weight of the drill string for the deepest hole, based on the length of all the drill pipes in the pipe changer and the base drill pipe coupled to the drill bit through the stabilizer, can be easily calculated. The likely weight of the rotary head can also be judged with fair accuracy. The depths of blastholes are very shallow. Though the formation in which blastholes are drilled are stable, the instances of drill strings getting stuck in the blasthole due to a piece of rock at the wall do occur. It is very difficult to determine the force needed to loosen the drill string in such instances. A factor of safety of about 2.5 over the combined weight of the longest drill string and rotary head mentioned above appears to be adequate. Pullout speed is not a critical criteria and any value above 30 m/min is usually acceptable. This speed must be achieved while the pullout force, as mentioned above, is being exerted on the drill string. In many blasthole drills, hydraulic cylinders are used for exerting feed and pullout force. Particular attention in such cases must be given to the above values because one side of the hydraulic cylinder, where there is no piston rod, can create enough feed force but the other side may not create enough pullout force for lifting the drill string.

13.7

BREAKOUT TORQUE

When a drill pipe is added to the drill string it is screwed onto the lower drill pipe. This action is called making up of the tool joint. This is always done by the rotary motion passed on to the drill pipe from the drill head. A drill head may exert only limited torque on the drill pipe but due to vibrations in the drill string, the drill pipe gets tightened so much that even if the drill head is reversible it is not able to loosen the upper drill pipe. Further, it is certainly unwise to use a very costly assembly like the drill head to loosen and uncouple drill pipes.

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385

For this reason, when drill pipes are to be loosened, the lower drill pipe is prevented from rotating by use of a spanner that holds the drill rod at the wrench slots as shown in Figure 7.36. After this a special hydraulic wrench like the one shown in Figure 6.21 grips the curved surface of the drill pipe and exerts very heavy torque on the drill pipe to make it come loose. This is termed breaking out the tool joint. The torque required for this operation is called breakout torque. Breakout torque is usually higher than the torque available from the drill head but certainly lower than the maximum torque rating for the BECO tool joint as stated in Table 6.5. Thus, the breakout wrench should be designed for the maximum torque rating for the BECO tool joint. It is not harmful in any way to have a higher torque rating of the breakout wrench for uncoupling the drill pipes because in any case, as per the principles of hydraulics, torque will be exerted only to the extent that is necessary for uncoupling the drill pipe.

13.8

HYDRAULIC LEVELING JACKS

A rotary blasthole drill equipped with four leveling jacks is certainly superior in terms of stability than a drill equipped with only three jacks. Therefore, unless a very compelling reason exists, blasthole drills must be equipped with four hydraulic jacks. This is particularly important in the case of a rotary blasthole drill because it has a very heavy mast, pipe changer and heavy drill head. All these assemblies take the center of gravity of the rotary blasthole drill to a higher level. The lifting capacity of each of the four jacks should be more than 75% of the estimated weight of the drill. While drilling inclined blastholes, apart from vertical forces, horizontal forces are also exerted on the drill frame and are eventually passed on to the hydraulic jacks. Since the angle of inclination of drilling a blasthole is limited to 30°, the maximum horizontal force to be resisted by the four jacks together works out to 0.5 times the bit loading. When maximum feed force is exerted, the rear jacks lose weight transferred to the ground through them and become less efficient in withstanding any horizontal force. Thus, keeping a factor of safety of 2, each of the jacks should be designed to resist 100% of the total horizontal force to be resisted. In most cases the piston rod of the hydraulic jack is insufficient to resist such heavy force. However, an additional circular or preferably square cylinder surrounding the hydraulic jack and telescoping within another cylinder fixed to the frame of the drill as shown in Figure 13.3 can do the job well.

13.9

GROUND LOAD BEARING

A rotary blasthole drill stands and travels on ground that is leveled and prepared. Even if the ground is somewhat loose, rubber-tired vehicles can travel on such prepared ground without much difficulty. The pressure exerted on the ground by these

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386


Hydraulic Fluid

Cylinder Fixed to the Drill Frame

Drill Frame

Hydraulic Cylinder

Piston Rod

Inner Cylinder Fixed to the Piston Rod

Load Bearing Pad

Figure 13.3

Construction of hydraulic jack to withstand horizontal forces.

tires is about 5 kg/cm2. However, rubber-tired vehicles have a very low center of gravity and are not prone to unequal settlement of ground under the tires. Therefore, the diameter of the bearing pads below the leveling jacks should be such that their contact pressure does not exceed about 1 kg/cm2. When it comes to determining the diameter of the pads below the hydraulic jacks a bearing pressure value of 3 kg/cm2 can be used because of practical limitations on the size of the pads, and the fact that when these pads are loaded the drill is made horizontal, and somewhat higher settlements below pads can be tolerated.

13.10

COMPRESSOR DISCHARGE AND PRESSURE

Discharge and pressure of compressed air from the drill compressor is of utmost importance. The procedure to be used for calculating the desired discharge of compressed air from the drill compressor is elaborated in article 10.4 in chapter 10. Similarly, in the same chapter, in article 10.5 the procedure for calculating pressure required from the drill compressor is elaborated.

13.11

ENGINE POWER

In rotary blasthole drilling, activities to be carried out include propelling, leveling, pipe handling, drilling and drill string withdrawal. To carry out these activities, many operations are required. Each of the operations needs power that can be calculated from mechanical engineering principles.

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387

In diesel-operated drills power is supplied by one large engine. In the case of electric drills, electric power goes directly to certain drive motors and for certain operations it is converted into hydraulic flow under pressure. All the operations, except mast raising and lowering, are listed in Table 13.6. The table clearly indicates which operations are to be carried out during different activities. In almost all the rotary blasthole drills the drilling activity requires more power because of many different operations as indicated in Table 13.6. Thus the engine power rating should be determined on the basis of drilling operations. If careful analysis of the percentage of total power required by each of the operations is made, it will show that operation-wise power requirements will be as shown in Table 13.6 after the name of the operation. Table 13.6 Simultaneous operations in drilling activities.

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Operations required to be carried out for the activities. Figures in the bracket show the approximate percentage of total engine power required for the operation

Different drilling activities

Run Propel Motors (30%–50%) Extend Leveling Jacks (10%–20%) Running Rotary Head Motors (15%–20%) Running Feed Motors/Cylinders (2%–4%) Air Compressor Full Working (45%–55%) Air Compressor Idling (25%–30%) Running Cab Air Conditioner (0.5%–1%) Running Cab Pressurizer (0.5%–1%) Running Cab Heater (0.5%–1%) Running Cable Reel Motors (0.5%–1%) Running Mach. House Air Con. (1.5%–2%) Running Mach. House Pressurizer (1%–2%) Running Water Injection Pump (1%–2%) Running Auxiliary Winch (3%–4%) Carousel Operation (2%–3%) Running Dust Collector (3% to 5%) Lighting (2%–3%)

Yes

Drill string Propelling Leveling Pipe handling Drilling Withdrawal Yes Yes Yes Yes Yes Yes Yes Yes

Yes Yes Yes

Yes Yes Yes Yes

Yes

Yes Yes Yes Yes

Yes Yes Yes

Slow Yes Yes Yes Yes Yes

Yes

Yes

Yes Yes

Yes

11/22/2010 2:39:33 PM

Chapter 14

Cost analysis of rotary blasthole drilling

14.1

INTRODUCTION

Systematic mining activities have been carried out for the past four centuries. The scale of these activities was always on the increase. In the initial period ore reserves were near the surface of the earth and could be easily accessed, directly i.e. without a need for removal of a significant quantity of overburden. In those days the final materials such as coal, or minerals of gold, iron, copper etc. were produced at low cost. With the passage of time the demand for such final materials increased. ‘Easy to recover’ reserves dwindled and larger quantities of overburden were required to be removed to access the mineral of interest. It became essential to take advantage of mining machinery. As a result, the cost of minerals increased. Today mining companies are forced to excavate overburden on a very large scale and gain relatively low quantities of ore. Buyers of the final products are reluctant to pay higher prices and look for suppliers willing to accept lower prices. Due to the resulting competition, every mining house is always attempting to reduce the ore production cost, so they can survive in the fierce race in the sale of final products to the processing units. This is applicable not only to the mining industry but construction and almost all other industries. Needless to say that today, engineering cost analysis has become a very important subject. The principles of engineering economics are applicable to mining engineering. Cost analysis enables taking major decisions about a mining project right from the question of economic viability of the project. Once a mining project is found to be economically viable, it becomes possible to choose one of the many options for carrying out various operations. It also enables taking measures towards cost reduction in any of the chosen operations. In opencast mines, that work with drill and blast activities, the cost of drilling blastholes is always a major part of the total cost, often amounting to as much as 15%. With the exception of tunneling, most other civil engineering projects have very small drill and blast activities. This chapter is devoted to explaining the method for calculation of rotary blasthole drilling costs.

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14.2 WORKING PATTERN OF A BLASTHOLE DRILL Before going ahead with the calculations of cost one must clearly understand the working pattern of a blasthole drill. When a blasthole drill is to be used for say 350 days in a year, for 3 shifts of 8 hours each, it is actually to be used for 350 × 3 × 8 = 8400 hours. During this time period a blasthole drill can be available for cyclic drilling operations, or not available as it is under maintenance or undergoing repairs. The term ‘Availability’ is often used in this context. It is often defined as, A = TC/(TC + TM + TR) where A = Blasthole drill availability TC = Time in hrs for which drill was available for cyclic drilling operations TM = Time in hrs for which drill was under maintenance TR = Time in hrs for which drill was under repairs In this book the availability definition as above is taken for granted. Sometimes availability is also defined as: A = (TC + TM)/(TC + TM + TR) How much time is required to be spent on maintenance and repairs depends upon the type of the drill and the maintenance program as well as the alertness and judgment of the driller. Inclusion of an automatic lubrication system in a blasthole drill improves the drill availability by a multiple M, that varies from 1.03 to 1.06 depending upon the type of drill. Further, drill availability decreases as the drill becomes older. If the availability of drill in the initial period of operations is AI and in the last year of operation is AL, then availability reduction factor is AE = (AI − AL)/AI where AR = Availability reduction factor AI = Availability of blasthole drill in first year AL = Availability of blasthole drill in last year Table 14.1 presents typical values of useful drill life in hours, initial availability, availability multiplier and availability reduction factor for different types of blasthole drills. From these values the time for which a drill is available during its life can be calculated as H C = L D ∗ M ∗ A I ∗ AR

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Table 14.1 Life and availability related factors for different types of rotary blasthole drills. Availability multiplier M if the drill has automatic lubrication system

Availability reduction factor AR

Size

Factor

Type

Factor

Type

Factor

Useful life

Initial availability factor AI = A ∗ B∗C

Extra Large Extra Large Large Large Large Medium Medium Small Small

0.97

Electric

0.98

Electric

0.97

85000–100000

0.922082

1.06

0.92

0.97

Diesel

0.96

Hydraulic

0.95

80000–95000

0.884640

1.06

0.9

0.95 0.95 0.95 0.92 0.92 0.88 0.88

Electric Electric Diesel Electric Diesel Electric Diesel

0.98 0.98 0.96 0.98 0.96 0.98 0.96

Electric Hydraulic Hydraulic Hydraulic Hydraulic Hydraulic Hydraulic

0.97 0.95 0.95 0.95 0.95 0.95 0.95

75000–90000 70000–90000 60000–80000 55000–75000 50000–70000 40000–60000 30000–50000

0.903070 0.884450 0.866400 0.856520 0.839040 0.819280 0.802560

1.05 1.05 1.05 1.04 1.04 1.03 1.03

0.92 0.88 0.85 0.88 0.85 0.85 0.82

Power source related

Size related

Power distribution related

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392


where HC = Cyclic operation hours in drill life LD = Drill life in hours M = Availability multiplier AI = Initial availability AR = Availability reduction factor A rotary blasthole drill works in cyclic activities as mentioned in Table 14.2. Some cells in the table contain the word ‘calculation’. For these cells the time has to be calculated from the values of timings given in other cells. Formulae to be used for such calculations and the results thus obtained, are shown in the example of blasthole drilling costs presented at the end of this chapter. One important factor, viz. net penetration rate, is required to be accurately predicted for correct calculations of the blasthole drilling costs. The methods and equations used for prediction of penetration rate have been described in chapter 3. As the values of the economic life of a blasthole drill and related factors mentioned in Table 14.1, as also cyclic activity timings in Table 14.2, depend upon a blasthole drill, more authentic data should be obtained from the blasthole drill manufacturers.

Table 14.2 Activities in a blasthole drilling cycle.

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Sr. no.

Cyclic activity

Typical time in s

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22

Moving from one Hole Location to the Other Lifting the Blasthole Drill by Hydraulic Jacks and Level It Lowering the Bit and Starting Drilling at Low Speed Drilling to Complete Depth of the Blasthole Uncoupling the Drill Head from the Drill Pipe Moving Up the Drill Head to the Top of the Mast Positioning the Pipe Changer in Drill String Alignment Coupling the Drill Head to the New Drill Pipe Removing the Pipe Changer from Drill String Alignment Coupling the New Drill Pipe to the Lower Drill Pipe Attaching the Desired Number of drill pipes Moving up the Drill Head + Drill Pipes by one Drill Pipe Length Moving up the Complete Drill String Uncoupling the Drill Pipe from Lower Drill Pipe Positioning the Pipe Changer Uncoupling the Drill Head from Drill Pipe Removing the Pipe Changer Lowering the Drill Head to the Bottom of the Mast Coupling the Drill Head to the Lower Drill Pipe Detaching the Desired Number of Drill Pipes Lowering the Drill on to the Crawler Base Complete Cyclic Operation

45 30 20 Calculation 25 30 25 20 15 25 Calculation 40 Calculation 30 30 25 18 35 25 Calculation 25 Calculation

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14.3

393

FACTORS RELATED TO BLASTHOLE DRILLING COST

In every business activity cost is incurred for output. In blasthole drilling the output is in terms of diameter and length of the blastholes drilled. As explained earlier the diameter and depth of the blastholes to be drilled are determined on the basis of production requirements and the other associated equipment such as dragline, shovel, dumper etc. used for mining operations. As the diameter of blastholes usually remains the same throughout the life of the drill, the cost of blasthole drilling is determined in terms of currency/m e.g. US$/m or Euro/m or Indian Rs/m etc. Cost is always dependent upon local factors, and as such it changes from nation to nation, from place to place, or in very strictest terms, even from one bench to another in the same mine. For this reason it will be very unwise to convert blasthole drilling cost on the basis of conversion rates for currencies. In other words, if the exchange rate is US$ 1 = Indian Rs 45 and if in a US copper mine operation the drilling cost is US$ 15/m, for drilling operations in a copper mine in India, blasthole drilling costs should not be taken as Indian Rs 675/m, even if the blasthole drill and rock properties are the same. Blasthole drilling cost can be divided into three major costing heads viz. owning, operating and overhead cost.

14.4

OWNING COST

Owning cost is the expenditure for owning a blasthole drill and keeping the ownership of the drill till the end of its life. For this the drill has to be purchased and yearly insurance and taxes have to be paid on it. In cost calculations the owning cost is depreciated over the life of the blasthole drill. It is calculated by dividing the total expenditure made for owning a blasthole drill by the expected drill life in cyclic operating hours. Owning cost is also called fixed cost. This cost has to be incurred irrespective whether the drill is used for production or not. In rotary blasthole drilling, the owning cost is very high as compared to the other drilling methods. This is due to the high price of rotary blasthole drills and accessories. Factors that are involved in owning cost are: 1 2 3

Purchase Expenditure Yearly Taxes, Duties and Levies Salvage Value Following are the individual elaborations.

14.4.1

Purchase expenditure

Purchase expenditure involves all the cost items till the blasthole drill is purchased, transported to the site, erected and commissioned at the site to start the drilling operations. Components that together form purchase expenditures are given in Table 14.3.

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394


Most of the expenditures mentioned in Table 14.3 occur once in the lifetime of a rotary blasthole drill. The table also gives the information as to where the values of expenditure are to be obtained from and to whom the amounts are payable as per common commercial practice. The following points in respect of purchase expenditure are worth noting. Table 14.3 Items of purchase expenditure for a blasthole drill. Sr. no. 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19

20

Details of the items in purchase expenditure Ex Works Price of the Drill Expenses for Arranging Pre-dispatch Inspection of the Drill at Plant Payment to Inspection Agency for the Pre-dispatch Inspection Packing or Export Packing Loading the Drill or Its Components on Trucks or Trailers Freight and Insurance up to Worksite or Port of Export En Route Charges, Such as Toll, Tax etc., to Worksite or Port of Export Documentation for Transport to Worksite or Export Port Charges at Port of Export Unloading the Drill from Trucks or Trailers Loading the Drill on the Ship Ocean Freight and Insurance up to Port of Import Unloading the Drill from Ship Import Duties and Custom Clearance Charges All Other Duties Payable at Port of Import Port Charges at Port of Import Loading the Drill on Trucks or Trailers Freight and Insurance up to Worksite En Route Charges, Such as Toll, Tax etc., to Worksite Purchaser’s Transporter Unloading the Drill at Worksite

Firm values provided by

Amount payable to

Manufacturer Manufacturer


Inspection Agency

Inspection Agency



Manufacturer/Transporter

Manufacturer/ Transporter Manufacturer/ Transporter Manufacturer/ Transporter Manufacturer/ Transporter Manufacturer/ Transporter Manufacturer/ Transporter Manufacturer/ Transporter Manufacturer/ Transporter Clearing Agency/ Govt. Clearing Agency/ Local Bodies Clearing Agency/ Port Authority Purchaser’s Transporter Purchaser’s Transporter

Manufacturer/Transporter Manufacturer/Transporter Manufacturer/Transporter Manufacturer/Transporter Manufacturer/Transporter Manufacturer/Transporter Manufacturer/Transporter Import Clearing Agency Import Clearing Agency Import Clearing Agency Purchaser’s Transporter Purchaser’s Transporter Purchaser’s Transporter

Purchaser’s Contractor

Purchaser’s Contractor (Continued )

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395

Table 14.3 (Continued) Sr. no. 21

22 23 24 25 26

1

2

3

4

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Details of the items in purchase expenditure Bank Charges for Arranging Payments through Letter of Credit and Its Confirmation Erection Charges Payable to Manufacturer Charges for Tools, Tackles, Power etc. Used in Erection Charges for Labor Used for Erection Charges for Providing Lodging and Boarding to the Erection Engineering Charges for Commissioning and Inauguration

Firm values provided by

Amount payable to

Purchaser’s Bank/Lending Agency

Purchaser’s Bank/ Lending Agency

Manufacturer

Manufacturer

Purchaser’s Contractor Purchaser’s Contractor Usually Arranged by Purchaser Usually Purchaser

Purchaser’s Contractor Purchaser’s Contractor Absorbed by Purchaser Usually Absorbed by Purchaser

Large blasthole drills are always dispatched in component form. They are assembled at the work site. For such erection of the drill at the work site, the help of experienced erection engineers from the manufacturer is always required. For medium and small blasthole drills the mast is often sent separately from the assembly that contains the upper and undercarriage, so as to reduce the volume and transport cost of the drill. Assembling the mast components on site is fairly easy and can often be done by the service engineers of the purchaser or some local agencies specialized in rendering such services. The manufacturer’s shop manual usually contains all such information. In marine transport, charges for transportation are calculated on the basis of volume of the item or weight of the item, whichever yields more payment to the shipping agency. For this purpose the component or drill data about dimensions and weight stated by the manufacturer is taken as standard. The volume is calculated on the basis of L × W × H of the component or drill. In most of the cases 1 m3 volume is taken as 1 tonne i.e. 1000 kg. Therefore if an item has a volume equal to 10 m3 but the weight is only 3500 kg, the item attracts freight on the basis of 10 m3 volume. However, if an item weighing 2000 kg weight has volume of 1 m3 the freight charges are based on 2000 kg weight and not on 1 m3 volume. Frequently, a purchaser makes payment to the manufacturer by taking a loan from banks or a financial institution. In such cases the loan has to be repaid within a certain period of time by way of monthly or yearly installments. Value for installments are worked out by the banks or financial institutions and the purchaser is duly informed about such installments to be paid to the bank or financial institution. The total interest paid by the customer can easily be calculated from standard financial formulae presented later. In most purchases an irrevocable letter of credit is required to be opened by the purchaser in favor of manufacturer. Further, the letter of credit may have to be

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confirmed to the manufacturer by the manufacturer’s banker. All the charges for opening and confirming the letter of credit are normally payable by the purchaser directly to their bankers. For calculating owning cost these charges must be treated as a part and parcel of the owning cost.

14.4.2 Yearly taxes, duties and levies Even if the blasthole drill works within the private premises of a mine, some taxes, duties and levies are payable to the government or local bodies on a yearly basis as per the laws, rules and regulations laid down by the government. The amount payable towards these can be easily found from the concerned government departments or the local bodies. In cost calculation practice, it is usually taken as a percentage of the blasthole drill price.

14.4.3

Salvage value

When a blasthole drill reaches the end of its useful life it can be sold. The amount received by such sale is called the salvage value. The salvage value of a rotary blasthole drill is to be judged carefully after giving due consideration to the following factors. 1 2 3

As per present scenario of technological advances there is hardly any likelihood that the rotary blasthole drilling method will become obsolete in near future. There are cases where rotary blasthole drills, particularly the all-electric type, have been used for as long as 30 years. The refurbishing of a rotary blasthole drill is relatively simple and the cost involved is relatively low in comparison to the price of a new blasthole drill of the same capabilities.

Salvage value cannot be judged easily. If some reliable data are available they can be used for the purpose of determining the owning cost. In the absence of any data, the salvage value of a crawler-mounted rotary blasthole drill can be determined on the basis of Table 14.4. The value of S can be determined from following equation. S = f * 0.95 * R * W where S = Salvage value in a currency f = Salvage value factor (Col. 4 of Table) R = Scrap steel selling rate in currency/kg W = Weight of the drill indicated in the manufacturer’s literature The factor 0.95 has been used in consideration of the fact that the drill has undergone wear and tear and its weight has reduced.

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Table 14.4 Salvage values of rotary blasthole drills.

Rotary

Blasthole Drill

Salvage value factor f

Extra Large

Electric

Electric

1.4

Extra Large

Diesel

Hydraulic 1.3

These drills are also rarely purchased for repair and resale. They have only one electric motor and therefore do not have large proportion of copper in their weight.

Large

Electric

Electric

These drills are also rarely purchased for repair and resale. They have fewer copper components.

Large

Electric

Hydraulic 1.1

These drills are also rarely purchased for repair and resale. They have only one electric motor and therefore do not have a large proportion of copper in their weight.

Large

Diesel

Hydraulic 1

These drills are also rarely purchased for repair and resale. They have do not have any copper components.

Medium

Electric

Hydraulic 1.4

Due to their smaller size these drills are purchased for repairs and resale. It is easier to get most of the components required for their repairs. The cost and risk involved in refurbishing such blasthole drills is also low.

Medium

Diesel

Hydraulic 1.5

Due to their smaller size these drills are purchased for repairs and resale. It is easier to get most of the components required for their repairs. Diesel engine drive makes these drills more easily saleable. The cost and risk involved in refurbishing such drills is also low.

Small

Electric

Hydraulic 1.7

As the size of these drills is very small, these drills are purchased for repair and resale to a much greater extent. It is easier to get most of the components required for their repairs. The cost and risk involved in refurbishing such blasthole drills is also low.

Small

Diesel

Hydraulic 2

As the size of these drills is very small, these drills are purchased for repair and resale to a much greater extent. It is easier to get most of the components required for their repairs. Diesel engine drive makes these drills more easily saleable. The cost and risk involved in refurbishing such drills is also low.

14.4.4

1.4

Reasoning behind suggested salvage value Large electric rotary blasthole drills are rarely purchased for repair and resale because it is very difficult to find a purchaser. Such drills are normally scrapped at the end of their life. Electric blasthole drills, however, have substantial quantity of copper and hence can fetch a good salvage value in proportion to their scrap value.

Calculation of owning cost

The sale proposal from a manufacturer gives firm prices on FOB (Free On Board) works basis and includes estimates or firm offers for inland freight to work site or port of export and ocean freight up to port of import. In the case of large drills the proposal also gives charges for providing services of their erection engineer.

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A purchaser should ask the prices for accessories and tires (in case of rubber tired blasthole drills) included in the price of the blasthole drill because they are considered as consumables and are taken into consideration in the operating cost. Therefore, the prices of drilling accessories and tires has to be deducted from the drill price. More often than not, a purchaser takes a loan from a bank or lending agency for making payment to the manufacturer. In such transactions the purchaser agrees with the bank or lending agency to bear a certain part of the payment immediately and requests a loan on the balance part. The amount immediately paid by the purchaser is called the down payment. The bank in turn makes full payment directly to the manufacturer and debits the down payment to the purchaser’s account. The loan, at a certain prefixed interest rate, given for the balance part is recovered by the bank or lending agency from the purchaser by way of a single installment or monthly or yearly installments of equal amounts payable in future. In loan transactions there is always an agreement between the purchaser and the bank or lending agency on down payment amount, loan amount, interest rate on yearly basis, number of installments and their period. On this basis the interest paid for the purchase of the drill can be calculated as follows. 14.4.4.1

Single installment

In this case the equation which relates various factors is: F = L * (1 + i/100)n where F = Amount of the single installment L = Loan amount n = Period of loan repayment in years I = Yearly interest rate in% Thus, the interest amount Ai, paid for the purchase works out to Ai = F − L and can be calculated easily. 14.4.4.2

Multiple installments

When the loan repayment is to be made through multiple installments the following equations can be used for calculation of the interest amount: Ai = A * n − L where Ai = Interest amount A = Amount of each installment n = Number of installments L = Loan amount

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The value of A is invariably specified by the bank or lending agency in the loan agreement. If it is to be calculated prior to the agreement the following equation can be used. A = L * {[(i/(k * 100)) * (1 + i/(k * 100))n]/[(1 + i/(k * 100))n − 1]} where A = Amount of each installment n = Number of installments L = Loan amount I = Yearly interest rate in% k = 12 or 1, for monthly or yearly installments When a loan has been taken, the total interest amount payable by the purchaser is to be added to the price of the blasthole drill in owning cost calculations. If no loan is taken, the down payment value should be made equal to the purchase expenditure for drill and accessories so the loan amount will become zero.

14.5

OPERATING COST

Operating cost arises when a blasthole drill is operated i.e. its engine or electric motors are being operated. A blasthole drill does not work when it is being repaired. If the drill is not provided with an automatic lubricating system it has to be stopped for carrying out manual maintenance. Thus, it can be said that the blasthole drill is operative during all the cyclic operation hours. Several cost factors contribute to the operating cost of a blasthole drill. Some of these are linked to the cyclic operation hours, whereas the others are based on drilled meterage that has to be calculated on the basis of net penetration rate observed during activity 4, 6 etc. of the cyclic operations. Groups in which the operating costs can be divided are: Cost of Maintenance and Repairs Cost of Consumables Cost of Operating Labor Cost of Accessories and Bits The following elaboration gives all the necessary details.

14.5.1

Cost of maintenance and repairs

For almost every blasthole drill, its manufacturer lays down a procedure for scheduled maintenance based on the data collected during field operation of the drill. As per this procedure certain parts used in the assemblies of the drill are replaced after a specific operating period. It is in the best interest of the owner to follow such procedure as this way the need for unscheduled maintenance or repairs reduces greatly.

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Such preventive maintenance procedure changes from machine to machine. The cost of the preventive maintenance is by way of the price of the replaced parts and labor used for the purpose. Proper records of preventive maintenance kept for one machine can greatly help in calculating maintenance costs for similar blasthole drills in a very reliable way. If such records are not available for estimation of hourly maintenance and repair cost, the multiplier presented in Table 14.5 can be used. The formula for calculations is as under. Cm = fm * P where Cm = Maintenance and repair cost for each shift hour fm = Multiplication factor (Table 14.5) P = FOB Works price of the blasthole drill without accessories and tires The following points are to be noted when the maintenance and repair cost is found by using the above formula: 1 2

Maintenance and repair cost does not take into account the consumption of lubricants. The values presented in the table have been judged on a logical basis and are arbitrary. For the sake of uniformity this cost must be converted in terms of cyclic operation hours.

14.5.2

Cost of consumables

Consumables are those which are either fully or partly burnt or become useless for their purpose or are lost irrecoverably. The following elaboration will provide adequate guidelines for calculating quantities of consumables used during drill operation. The consumables cost can be found by appropriate multiplication of the price of the consumable. Consumable costs in a blasthole drill are as under. Table 14.5 Multiplication factor to be used for estimation of hourly maintenance and repair cost.

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Rotary blasthole drill

Maintenance cost multiplier fm

Extra Large Extra Large Large Large Large Medium Medium Small

Electric Diesel Electric Electric Diesel Electric Diesel Electric

Electric Hydraulic Electric Hydraulic Hydraulic Hydraulic Hydraulic Hydraulic

4 * 10–5 5 * 10–5 4.5 * 10–5 5.5 * 10–5 5.75 * 10–5 6 * 10–5 6.5 * 10–5 6.5 * 10–5

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14.5.2.1

401

Cost of power

Most of the rotary blasthole drills are powered by diesel engines. Maximum rated power output of such engines varies over a wide range from 175 to 1120 kW. For most of the diesel engines the ASFC (Average Specific Fuel Consumption) is about 215 g/kWh. As diesel has a density of 0.839 kg/L, the ASFC equals 0.256 L/kWh. If a more accurate estimate is to be made, the manufacturers have the necessary data on the basis of the tests conducted by them under well-controlled conditions. As explained in one of the earlier chapters, the operation of a rotary drill seldom requires the maximum power from the diesel engine, and the average power requirement can be taken as 75% of the maximum rated power output. Thus, if a blasthole drill has a diesel engine rated at 250 kW power output, it is likely to consume 250 * 0.256 * 0.75 = 48 L of diesel oil per cyclic operation hour. With this value of fuel consumption the cost can be found by using the diesel price. In some blasthole drills only one electric motor is used for driving the compressor from one end of its shaft and the hydraulic pump drive unit from the other end of its shaft. Since electric motors have good overloading capacity, the motor rating is about 90% of ‘in lieu’ diesel engine rating. Therefore, the power consumption of the electric motor can be taken as about 75% of the rated power of ‘in lieu’ diesel engine or 0.75/.90 i.e. 0.83 or 83% of the rated power of the electric motor. By their nature, electric motors are quick acting. If a sudden additional power requirement comes up, a diesel engine may take say 5 s to adjust to the power requirement but an electric motor can adjust to the requirement within 1 s. Due to such characteristics, the need of power from an electric motor comes further down to say 0.9 × 0.95 i.e. 85.5% compared to a diesel engine. Thus, a blasthole drill equipped with a single motor rated at 225 kW will consume electric power equal to 225 * 0.83 * 0.95 = 177.4 kW during each hour of its cyclic operation. If the blasthole drill is all-electric type, the power consumption is less because the power drawn by a non-working component is simply switched off and there is no idle running involved. In all-electric drills the motors are separate and the sum of their rated power is about 25% higher than the rated power of a single motor or about 0.9 * 1.25 i.e. 1.125 or 112.5% of the equivalent diesel engine. Thus the power requirement of an all-electric blasthole drill can be taken as 0.75/1.125 i.e. 66% of the total power installed on the drill by way of electric motors. In view of the above quick overloading capabilities of electric motors, an allelectric blasthole drill having a total rating of 280 kW, will consume electric power equal to 280 * 0.66 * 0.95 = 175.5 kWh. The above values are already based on cyclic operation hour. The known rate of electric power, in currency/kWh, enables the calculation of hourly power cost. 14.5.2.2

Drill lubricants

Drill lubricants used in rotary blasthole drills are of two types viz. grease or high viscosity oil.

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An automatic lubrication system in a blasthole drill uses high viscosity oil. In some blasthole drills grease is manually fed to the component through grease nipples provided for the purpose. Consumption of lubricant depends the number of points at which lubricant is required to be fed. It can be approximately related to the size of the blasthole drill. Hourly consumption of lubricant required for a rotary blasthole drill equipped with an automatic lubricating system can be judged from Table 14.6. For drills where lubricant is fed manually, the values of lubricant consumption stated in the table may be treated as kg/h instead of L/h. Hourly cost of the lubricant can be found from the known price of the lubricant in currency/L or currency/kg as the case may be. It has been noted that the life of tricone bits increases significantly if a small quantity of lubricant is mixed with the flushing air. But such practice has to be adopted only after careful experiments because if a large quantity of lubricant is fed, it may choke up the air passages of the tricone bit resulting in reduced bit life. It may be worthwhile to use the lubricating oil injection mechanism intermittently. If this practice is adopted the extra quantity of lubricating oil required for the bit lubrication can be found from the data gathered during experiments. For the purpose of estimation the consumption can be correlated to the bit diameter as under. Q = 6 * 10–3 * D where Q = Bit lubricating oil required in L/h D = Bit diameter in mm It must be borne in mind that the cost worked out by the above formula is for each drilling hour and not for cyclic operation hour. Hence, the cost must be spread over cyclic operation hour by using the appropriate factor. Here again, hourly cost of the bit lubricating oil can be found from the known price of the bit lubricating oil in currency/L.

Table 14.6 Lubricating oil consumption for drills with automatic lubricating system.

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Rotary blasthole drill

Lube oil consumption L/h

Extra Large Extra Large Large Large Large Medium Medium Small Small

Electric Diesel Electric Electric Diesel Electric Diesel Electric Diesel

Electric Hydraulic Electric Hydraulic Hydraulic Hydraulic Hydraulic Hydraulic Hydraulic

1.2 1.1 1.06 1.06 1.06 0.88 0.88 0.77 0.77

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14.5.2.3

403

Oils

Depending upon the type of drill the following oils are used in various blasthole drill components. Compressor Oil Engine Oil Hydraulic Oil Transformer Oil These oils burn due to heat generated by friction between moving parts or heat generated in power loss. Some users simply throw away the used oil and replace it with new, whereas others reuse the oil after appropriate filtering and replenishing the extra quantity needed. In any case a cost is involved in replenishment (full or partial) as well as filtering. The time interval for such filtering and/or replenishment, as stipulated by the manufacturer, should be followed. Hourly cost for each of the above four types of oils can be found by using the following equation. Co = (Qn * Pn+Qf * Pf)/H where Co = Hourly oil cost in currency/h Qn = Volume of new oil used in L Pn = Price of new oil in currency/L Qf = Volume of filtered oil used in L Pf = Price of filtered oil in currency/L H = Oil change interval in hours The following points are worth noting. 1 2 3 4

A blasthole drill driven by a single electric motor usually does not have any transformer as the motor is run directly on the main supply voltage. An all-electric drill has many motors with vastly different power ratings. Such drills often need a transformer. Needless to say that if there is no transformer, there is no transformer oil. The interval for changing transformer oil is very long.

If data needed for calculating individual oil cost are not available then the cost of all the oils used can be taken as 12% to 20% of fuel cost depending upon the type of blasthole drill. 14.5.2.4

Water

Rotary blasthole drills require water for dust suppression, for the radiator that cools the diesel engine and for the radiator that cools the compressor. How much water is required for these purposes can be estimated easily from experience of using the drills or similar equipment on site. The need for water depends upon weather conditions.

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At the work site water may be sufficiently available and the cost of the water itself may be negligible but the cost for its transportation to the drill and pumping the water into appropriate components may need consideration. Despite the fact that water cost is usually neglected, wherever necessary it can be estimated and used in calculation of blasthole drilling costs.

14.5.2.5

Tires

Truck-mounted rotary blasthole drills are used very rarely. As the price of their tires is relatively high and working life is low, their replacement cost is quite significant. Therefore, tires are treated as consumable. Unlike many types of excavation/hauling equipment, e.g. wheel loaders, trucks, motor graders etc., the wear and tear of tires on blasthole drills is much less because they are used only while the blasthole drill is moved from location of one blasthole to another. Unless tire condition warrants otherwise, all the tires are replaced simultaneously. Hourly tire cost can be found by dividing the price of the tires replaced by time interval in hours over which they are replaced. In rotary blasthole drills the tire replacement interval can be taken as 8000 working hours and hourly cost is worked out on the basis of tire price.

14.5.2.6

Power cable

In electrically operated blasthole drills a long trailing power cable is required for supplying electricity to the blasthole drill. This cable undergoes wear and tear and is required to be changed after a certain time interval. As the cable is quite an expensive item, and is required to be used only in the case of electrically operated drills, its cost has to be taken into consideration separately. The cost of the cable can be found by knowing the length of cable used and its price per meter length.

14.5.3

Cost of accessories and bits

The process of drilling requires drilling bits and many accessories. As these items undergo heavy wear and tear, they are replaced frequently. The useful life is of these items is usually given in terms of drilled meters rather than hours. It is extremely difficult to predict the lifetime of accessories and particularly drill bits as it depends upon numerous factors related to drilling parameters as well as the properties of the rock mass being drilled. Rather than relying on the data presented in a forthcoming part of this chapter, attempts must be made to obtain more realistic data. Manufacturers of the drilling bits and accessories often have performance records of their drilling bits and accessories working under vastly different conditions. While making inquiries with them, many details about the environments, rock mass characteristics, rock properties, types of machines to be used as well as diameter and depth of the intended blasthole are required to be given so the data sent by them will be more realistic. For the purpose of cost estimation, drill bits and accessories are treated separately as the differences in their lives is very large.

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14.5.3.1

405

Drilling accessories

In rotary blasthole drilling the main accessories used are shock absorber, crossover sub, drill pipe and a stabilizer. All these accessories are subjected to very heavy feed force, torque and vibrations. A shock absorber always remains outside the blasthole. Its life is depends upon its construction, the environmental factors, the vibration levels to which it is subjected etc. The estimated life of shock absorbers is given in Table 14.7. Other accessories go inside the blasthole and are continuously subjected to impacts from the drill cuttings traveling up at very high speeds. Such continuous sandblasting causes very heavy wear on the outer surface of the accessories. If the formations are very hard and abrasive the life expected is much lower than that in soft and non-abrasive formations. For rough estimation, the life of some rotary drilling accessories is given in Table 14.8. These estimates and the prices of the accessories enable evaluation of the cost of accessories in terms of currency/m. 14.5.3.2

Tricone bits

Evaluating the life of the tricone bit in terms of m is a very difficult task. Fortunately, most of the bit users and manufacturers keep meticulous records of the drill bits used/ supplied by them. With such data, the bit manufacturers are the best advisers to a user in his efforts in evaluating bit life in this computer age. Table 14.7 Life of shock absorbers. Life in M Feed force rating of shock absorber

Hard fractured

Medium monolithic

550 to 600 kN 300 to 350 kN 200 to 250 kN 139 to 150 kN

8000 6000 5000 3000

12000 10000 8000 6500

Table 14.8 Life of rotary blasthole accessories. Useful life in M

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Name of the accessory

Hard, abrasive rocks

Soft, non-abrasive rocks

Drill Pipes Diameter 381 mm, 32 mm Wall Drill Pipes Diameter 273 mm, 25 mm Wall Drill Pipes Diameter 140 mm, 19 mm Wall Welded Blade Stabilizers Diameter 381 mm Welded Blade Stabilizers Diameter 251 mm Roller Stabilizers Diameter 381 mm Roller Stabilizers Diameter 251 mm Crossover Sub Diameter 381 mm, 32 mm Wall Crossover Sub Diameter 273 mm, 25 mm Wall Crossover Sub Diameter 140 mm, 19 mm Wall

24,000 20,000 15,000 6,500 4,000 7,500 6,000 5,000 3,500 3,000

35,000 30,000 25,000 8,000 6,000 11,000 9,500 7,000 6,500 5,000

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In the absence of such data an empirical equation as under is recommended for the purpose. L = 149.82D1.55 * P/(N * K1.67) where L = Bit life in hrs D = Diameter of the bit in mm P = Penetration rate in m/h N = Rotary speed in RPM K = Bit load in terms of thousand kg

14.5.4

Cost of operating labor

Blasthole drilling operation has mostly been made fully automatic in most of the modern blasthole drills. The drill itself is operated by one driller but in most cases a helper/ trainee driller accompanies the operator for assistance. The operating cost to be included on account of such labor is calculated by dividing the total annual emoluments (including all the benefits and bonuses) of the driller and helper/trainee driller, as the case may be, by cyclic operation hours. As the emoluments of driller and helper or trainee driller are dependent on local conditions, the variation is too great to give any estimations. 14.6

OVERHEAD COST

In every organization many expenditures are incurred by way of salary, rent, purchases etc. that cannot be attributed to a particular production machine like blasthole drill, shovel or truck. The total of such emoluments and expenditures is called the overhead cost. Depending upon the accounting practice adopted, sometimes these overhead costs are not considered in the internal communications. The management of the organization apportions these overhead costs in some logical manner to various production activities like blasthole drilling, blasting, loading, hauling etc. The part of overhead costs debited to blasthole drilling activity has to be further divided between the number of operating drills depending upon their ability, price etc. No specific formula can be given for evaluating the overhead cost but it can be easily done by applying the accounting practice used by the organization and common sense. 14.7 TOTAL BLASTHOLE DRILLING COST When all the above components of cost are added the sum gives the total blasthole drilling cost. Thus, Total Cost = Owning Cost + Operating Cost It is possible to expand the above formula to take into account various factors but the presentation is not of much use in the present world of computers because the formula

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407

can be written more purposefully in the form of a worksheet in a spreadsheet program as has been done later in this chapter.

14.8

GENERAL CONSIDERATIONS IN COST ESTIMATION

Elaborations in the earlier part of this chapter do enable calculation of blasthole drilling cost, but for improving accuracy the following factors need to be given due consideration. 1

2

3 4 5

6

7

Manufacturers of drills often give typical timings of different cyclic activities. They are usually based on the speed of electric motors or volume of hydraulic oil supplied to various components that are used in the activities. In actual practice a little more time is needed for completion of the activity. Sometimes pauses in cyclic operation are necessary for inspection of a bit or other drill string components. During these pauses the engine keeps on working and the compressors may also have to supply air. Hence consumables are used in the operation and the operating cost is incurred. Some pauses in cyclic activities may also be needed when the work site experiences severe weather. Stormy winds may require lowering the mast. Apart from the pauses there is always a time lag between starting the engine and start of the cyclic activity. Environmental conditions affect blasthole drilling costs to a very great extent. For example when a blasthole drill is to be operated in extremely cold weather not only must it be provided with several devices built-in but the driller and his assistants also have to be given several additional facilities like warm clothing, gloves etc. Every such item costs money and becomes a part of additional expenditure. Penetration rates estimated on the basis of rock sample tests may have good accuracy for drilling vertical holes but need to be adjusted because the actual blastholes may be at an angle. When blastholes are to be drilled at an angle with the vertical the drilling cost increases considerably because of the decrease in the life of drilling bits, accessories and even the blasthole drill.

14.9

EXAMPLE OF COST ESTIMATION

Table 14.9, spread over the next four pages, presents an example of calculations of blasthole drilling cost for an all-electric drill. Overhead costs have not been included as they depend upon the accounting method adopted in an organization. An attempt has been made to keep the values of inputs as realistic as possible. However, some input values may be somewhat unrealistic. In any case, when realistic values are used for calculations, the cost estimate will also be realistic. The table is in spreadsheet form. All the formulae required are given in the second column. From these, readers can construct their own spreadsheet on their computers.

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Book.indb 408

Table 14.9 Cost analysis of rotary blasthole drilling. Cost head

Source of input or formula

C1

Input/results

Type of Drill

Drill Specifications

C3

All Electric

Is the Drill Provided With Automatic Lubrication System?


C4

Yes

Blasthole Drill and Its Working Pattern

Weight of the Blasthole Drill in kg


C5

135000.00000

Scrap Value Factor

From Table 14.4

C6

1.40000

Estimated Economic Life of the Blasthole Drill in Shift Hours

Table 14.1

C7

95000.00000

Initial Availability of the Blasthole Drill

Table 14.1

C8

0.90307

Availability Multiplier

Table 14.1 (= 1.00 for Manual Maintenance)

C9

1.06000

Availability Reduction Factor

Table 14.1

C10

0.92000

Number of Drill Shift Hours Per Day

Rules of Mine

C11

24.00000

Number of Drill Operation Days in a Year

Rules of Mine

C12

355.00000

Number of Shift Hours in a Year

C13 = C11 * C12 C14 = C13 * C8 * C9 * C10

C13

8520.00000

C14

7503.34132

Number of Cyclic Operation Hours in the Life of Drill

C15 = C7 * C8 * C9 * C10

C15

83664.01708

Maintenance and Repair Cost Multiplier

Table 14.5

C16

0.00005

Maximum Power Rating of the Diesel Engine in kW (0 If Not Applicable)


C17

0.00000

Power Rating of the Single Electric Motor in kW (0 If Not Applicable)


C18

0.00000

Number of Cyclic Operation Hours in One Year

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Power Rating of All the Electric Motors in kW (0 If Not Applicable)


C19

530.00000

Type of Drill Lubricating System


C20

Automatic

Lube Consumption Multiplier

Table 14.6

C21

1.06000

Recommended Compressor Oil Change Interval in Cyclic Operation Hours

Compressor Specifications

C22

3000.00000

Volume of Compressor Oil Changed


C23

200.00000

Is the Drill Equipped With Diesel Engine?


C24

No

Recommended Engine Oil Change Interval in Cyclic Operation Hours

Engine Specifications

C25

1500.00000

Volume of Engine Oil Changed


C26

100.00000

Recommended Hydraulic Oil Change Interval in Cyclic Operation Hours


C27

2000.00000

Volume of Hydraulic Oil Changed


C28

400.00000

Is the Drill Equipped With Transformer?


C29

Yes

Recommended Transformer Oil Change Interval in Cyclic Operation Hours

Transformer Specifications

C30

8000.00000

Volume of Transformer Oil Changed

Transformer Specifications

C31

150.00000

Capacity of Radiator Tank for Engine Cooling


C32

150.00000

Interval for Changing Water in Engine Radiator


C33

24.00000

Capacity of Radiator Tank for Compressor Cooling


C34

100.00000

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(Continued )

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Table 14.9 (Continued) Cost head


C1

Input/results

Interval for Changing Water in Compressor Radiator


C35

48.00000

Is the Drill Equipped With Tires?


C36

No

Number of Tires on the Drill


C37

0.00000

Drill Bit Diameter

Blasthole Parameter

C38

349.00000

Average Net Penetration Rate in m/h

User Input Based on Tests

C39

30.00000

Single Pass Length for the Drill in m


C40

19.80000

11/22/2010 2:39:35 PM

Depth of the Drill Hole

Blasthole Parameter

C41

53.00000

Number of Drill Pipes Required for Reaching the Hole Depth

C42 = ROUND(C41/C40,2)

C42

2.68000

Number of Pipe Attachments and Detachments

C43 = INT(C42–0.001)

C43

2.00000

Time in s for Moving from one Hole Location to the Other

Drill Working Parameter

C44

45.00000

Time in s for Lifting the Blasthole Drill by Hydraulic Jacks and Level It


C45

30.00000

Time in s for Lowering the Bit and Starting Drilling at Low Speed.


C46

20.00000

Time in s for Drilling to Complete Depth of the Blasthole

C47 = 3600 * C41/C39

C47

6360.00000

Time in s for Uncoupling the Drill Head from the Drill Pipe


C48

25.00000

Time in s for Moving Up the Drill Head to the Top of the Mast


C49

30.00000

Time in s for Positioning the Pipe Changer in Drill String Alignment


C50

25.00000

Book.indb 411

Time in s for Coupling the Drill Head to the New Drill Pipe


C51

20.00000

Time in s for Removing the Pipe Changer from Drill String Alignment


C52

15.00000

Time in s for Coupling the New Drill Pipe to the Lower Drill Pipe


C53

25.00000

Time in s for Attaching the Desired Number of drill pipes

C54 = C43 * SUM(C48:C53)

C54

280.00000

Time in s for Moving up the Drill Head with Drill Pipes by one Drill Pipe Length


C55

40.00000

Time in s for Moving up the Complete Drill String

C56 = C55 * (INT(C42) + (1-(C42-INT(C42))))

C56

92.80000

Time in s for Uncoupling the Drill Pipe from Lower Drill Pipe


C57

30.00000

Time in s for Positioning the Pipe Changer


C58

30.00000

Time in s for Uncoupling the Drill Head from Drill Pipe


C59

25.00000

Time in s for Removing the Pipe Changer


C60

18.00000

Time in s for Lowering the Drill Head to the Bottom of the Mast


C61

35.00000

Time in s for Coupling the Drill Head to the Lower Drill Pipe


C62

25.00000

Time in s for Detaching the Desired Number of Drill Pipes

C63 = C43 * SUM(C57:C62)

C63

326.00000

Time in s for Lowering the Drill on to the Crawler Base


C64

25.00000

11/22/2010 2:39:36 PM

(Continued )

Book.indb 412



C1

Input/results

Time in s for Complete Cyclic Operation

C65 = SUM(C44:C47) + C54 + C56 + C63 + C64

C65

7178.80000

Drilled meters in each Cyclic Operation hour

C66 = ROUND(3600 * C41/C65,2)

C66

26.58000

Calculation of Owning Cost Depreciation

C68

FOB Works Price of the Blasthole Drill

Manufacturer’s Quotation

C69

1783500.00000

Payment to Inspection Agency for the Pre-dispatch Inspection

Inspection Agency Quotation

C70

12000.00000

Inland Freight to Destination or Port of Export (As Applicable)


C71

6800.00000

Ocean Freight and Marine Insurance (If Applicable)


C72

72000.00000

Clearing Charges at the Port (If Applicable)

Import Clearing Agency

C73

26000.00000

Inland Freight and Insurance Up to Work Site (If Applicable)

Purchaser’s Transporter

C74

27800.00000

Charges for Unloading the Drill at Work Site

Purchaser’s Heavy Lift Contractor

C75

3000.00000

Erection and Commission Charges

Manufacturer + Purchaser

C76

7000.00000

Expenditure for Arranging Letter of Credit and Its Confirmation

Purchaser’s Bank

C77

1500.00000

Price of the Drill and Accessories After Commissioning

C78 = SUM(C69:C77)

C78

1939600.00000

Down Payment

As Per Agreement with Bank

C79

600000.00000

11/22/2010 2:39:36 PM

Book.indb 413

Loan Amount

C80 = C78-C79

C80

1339600.00000

Interest Rate


C81

12.00000

If Installments are Monthly or Yearly


C82

Monthly

Value of k

C83 = IF(C82 = “Monthly”,12,1)

C83

12.00000

Loan Period in Years


C84

5.00000

Number of Installment for Repayment of the Loan

C85 = C83 * C84

C85

60.00000

Amount of Each Installment

C86 = ROUND(C80 * ((C81/(C83 * 100)) * (1 + (C81/(C83 * 100)))^C85/((1 + (C81/(C83 * 100)))^C85–1)),0)

C86

29799.00000

Total Interest Paid on the Loan

C87 = C85 * C86-C80

C87

448340.00000

Total Purchase Expenditure

C88 = C87 + C78

C88

2387940.00000

Price of Accessories and Tires Purchased with the Drill


C89

57000.00000

Estimated Scrap Rate, Per kg, at the End of Drill Life

Purchaser’s Estimate

C90

0.50000

Salvage Value of the Drill at the End of Economic Life of Drill

C91 = C5 * C6 * C90

C91

94500.00000

Total Cost to be Depreciated

C92 = C88-C89-C91

C92

2236440.00000

Depreciation Per Cyclic Operation Hour

C93 = ROUND(C92/C15,2)

C93

26.73000

Insurance and Taxes

C94

Insurance and Taxes Payable in a Year

Government/Local Bodies

C95

12000.00000

Insurance and Taxes Payable Per Cyclic Operation Hour

C96 = ROUND(C95/C14,2)

C96

1.60000

Total Owning Cost for the Drill Per Cyclic Operation Hour

C97 = C93 + C96

C97

28.33000

Calculation of Operating Cost 11/22/2010 2:39:36 PM

(Continued )

Book.indb 414



Maintenance and Repair Cost

C1

Input/results

C99

Amount to Which the Multiplier is to be Applied

C100 = C78-C89

C100

1882600.00000

Maintenance and Repair Cost Per Cyclic Operation Hour

C101 = ROUND(C16 * C100/(C8 * C9 * C10),2)

C101

96.20000

Price of Diesel Per L (In Case of Diesel Drill)


C103

NA

Price of Electricity Per kWh (In Case of Electric Drill)


C104

0.07000

Power Cost Per Cyclic Operation Hour

C105 = ROUND(IF(C17 > 0,C17 * 0.75 * 0.256 * C103, IF(C18 > 0,C18 * 0.83 * 0.95 * C104, IF(C19 > 0,C19 * 0.66 * 0.95 * C104,0))),2)

C105

Power Cost

C102

Drill Lubrication Cost

23.26000

C106

Price of Lubricating Oil Per L


C107

4.00000

Price of Lubricating Grease Per kg


C108

3.50000

Drill Lubricant Cost Per Cyclic Operation Hour

C109 = IF(C20 = “Automatic”,C21 * C107,C21 * C108)

C109

4.24000

Volume of New Compressor Oil Used in Oil Change


C111

100.00000

Price of New Compressor Oil Used in Oil Change


C112

11.00000

Compressor Oil Cost

C110

11/22/2010 2:39:36 PM

Book.indb 415

Price of Filtered Compressor Oil Used in Oil Change


C113

9.00000

Compressor Oil Cost Per Cyclic Operation Hour

C114 = ROUND((C111 * C112 + (C23-C111) * C113)/ C22,2)

C114

0.67000

Volume of New Engine Oil Used in Oil Change


C116

NA

Price of New Engine Oil Used in Oil Change


C117

NA

Price of Filtered Engine Oil Used in Oil Change


C118

NA

Engine Oil Cost Per Cyclic Operation Hour

C119 = IF(C24 = “Yes”,ROUND((C116 * C117 + (C26C116) * C118)/C25,2),0)

C119

0.00000

Engine Oil Cost

C115

Hydraulic Oil Cost

C120

Volume of New Hydraulic Oil Used in Oil Change


C121

300.00000

Price of New Hydraulic Oil Used in Oil Change


C122

12.00000

Price of Filtered Hydraulic Oil Used in Oil Change


C123

9.00000

Hydraulic Oil Cost Per Cyclic Operation Hour

C124 = ROUND((C121 * C122 + (C28-C121) * C123)/ C27,2)

C124

2.25000

Volume of New Transformer Oil Used in Oil Change


C126

80.00000

Price of New Transformer Oil Used in Oil Change


C127

12.00000

Transformer Oil Cost

C125

(Continued ) 11/22/2010 2:39:36 PM

Book.indb 416



C1

Input/results

Price of Filtered Transformer Oil Used in Oil Change


C128

9.00000

Transformer Oil Cost Per Cyclic Operation Hour

C129 = IF(C29 = “Yes”,ROUND((C126 * C127 + (C31C126) * C128)/C30,2),0)

C129

0.20000

Water Cost

C130

Price of Water Used in the Radiators


C131

0.05000

Water Cost Per Cyclic Operation Hour

C132 = IF(C24 = “Yes”,ROUND((C32 * C131/C33 + C34 * C131/C35),2),ROUND((C34 * C131/C35),2))

C132

0.10000

Tire Cost Life of Each of the Tires in Shift Hours

C133 Purchaser’s Estimate

C134

NA

Price of Each of the Tires


C135

NA

Tire Cost Per Cyclic Operation Hour

C136 = IF(C36 = “Yes”,ROUND(C135 * C37/C134,2),0)

C136

0.00000

Power Cable Cost Length of Power Cable Used


C138

300.00000

Price of Power Cable Used Per meter


C139

30.00000

Life of Power Cable in Cyclic Operation Hours


C140

12000.00000

Power Cable Cost Per Cyclic Operation Hour

C141 = ROUND(C140/(C139 * C138),2)

C141

1.33000

Cost of Drilling Accessories

C142 C143 = INT(C42–0.001) + 1

C143

3.00000

Estimated Life of a Drill Pipe in meters

Table 14.8

C144

25000.00000

Price of a Drill Pipe


C145

5700.00000

Number of Drill Pipes Used in the Drill String

11/22/2010 2:39:36 PM

Book.indb 417

Estimated Life of a Shock Absorber in meters

Table 14.7

C146

10000.00000

Price of a Shock Absorber


C147

22000.00000

Number of Crossover Subs Used in the Drill String


C148

3.00000

Estimated Life of a Crossover Sub in meters

Table 14.8

C149

6000.00000

Price of a Crossover Sub


C150

1200.00000

Is Stabilizer Used for Drilling?


C151

Yes

Estimated Life of a Stabilizer in meters

Table 14.8

C152

12000.00000

Price of a Stabilizer


C153

7500.00000

Total Cost of Drilling Accessories Per meter

C154 = ROUND(C145 * C143/C144 + C147/C146 + C148 * C150/C149 + IF(C151 = “Yes”,C153/C152,0),2)

C154

4.11000

Cost of Tricone Bits Estimated Life of a Tricone Bit in meters


C156

1960.00000

Price of a Tricone Bit


C157

7900.00000

Cost of Tricone Bit Per Meter

C158 = ROUND(C157/C156, 2)

C158

4.03000

Bit Lubricant Cost

C159

Price of Bit Lubricant Per L


C160

12.00000

Cost of Bit Lubricant Per Cyclic Operation Hour

C161 = ROUND(2 * 10^-3 * C38 * C160/(C8 * C9 * C10),2)

C161

9.51000

Cost of Operating Labor Number of Drillers Required to Run the Drill


C163

3.00000

(Continued ) 11/22/2010 2:39:36 PM

Book.indb 418



C1

Input/results

Number of Helpers Required to Assist the Drillers


C164

3.00000

Apportioned Number of Supervisor for the Drill


C165

0.50000

Yearly Emoluments of the Driller


C166

44000.00000

Yearly Emoluments of the Helper


C167

29000.00000

Yearly Emoluments of the Supervisor


C168

70000.00000

Cost of Operating Labor Per Cyclic Operation Hour

C169 = ROUND((C163 * C166 + C164 * C167 + c165 * C168)/C14,2)

C169

33.85000

Total Owning and Operating Cost Calculated on Cyclic Operation Hour Basis

C170 = C97 + C101 + C105 + C109 + C114 + C119 + C124 + C129 + C132 + C136 + C141 + C161 + C169

C171

199.94000

Total Operating Cost Calculated on Per Meter Basis

C171 = C154 + C158

C172

8.14000

Total Owning and Operating Cost Calculated on Per Meter Basis

C173 = ROUND(C171/C66 + C172,2)

C173

15.66000

Total Costs

11/22/2010 2:39:36 PM

Chapter 15

Procurement of rotary blasthole drills

15.1

INTRODUCTION

Many of the previous chapters in this book have been devoted to elaborating various technical aspects of rotary blasthole drills and associated drilling equipment viz. drill bits and drilling accessories. In some other chapters, topics such as interaction between rock formations and the drill bits, flushing of the blasthole etc. that are inseparably connected with rotary blasthole drilling have been discussed at length. All such background knowledge on the subject will certainly be useful to the drillers and drilling engineers to be better at their job. However, it is desirable to cover some more topics associated with rotary blasthole drilling so as to leave no gaps in the background knowledge. Procuring the right drill for the job is of utmost importance. Topics related to this are: 1 2

Pre Procurement Considerations Procurement Methodology. This chapter covers these topics.

15.2

PRE PROCUREMENT CONSIDERATIONS

When a very large excavation work or a medium to large size mine is planned, unless some overriding considerations exist, such as the proximity of the project to a particular locality etc., in almost all cases the drill and blast method is the best choice. To actually carry out mining operations, many types of equipment are to be procured. Blasthole drills are required right from the start. Before procuring blasthole drills, the following matters must be thoroughly investigated. 1 2 3 4

Book.indb 419

What are the blasthole parameters? Which drilling method should be adopted? What type of drills – diesel or electric? How many blasthole drills are needed?

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420


The parameters of the blasthole, viz. diameter, depth and inclination are related to blasting and are covered in one of the forthcoming chapters on blasting. Guidelines on other topics are as under.

15.2.1 Which drilling method? In medium to very large mining projects or very large scale construction excavations, the diameter of blastholes to be drilled usually works out to 170 mm or larger. For such blasthole drilling work the choice of method is restricted only to Rotary or DTH drilling. Guidelines for the choice can be availed from Figure 15.1. It is to be borne in mind that such choice of drilling method essentially needs further verification about likely penetration rate and consequent production rate, because a wrong choice can lead to disastrous consequences. Verification can be done by sending samples of rocks to one or more well-equipped laboratories and getting their test results and estimation of penetration rates. Even better verification can be done by hiring a truck-mounted rotary-cum-DTH drill and using it in a very calculated manner to actually drill a few trial blastholes with appropriately chosen tricone bits as well as DTH hammers and bits. Very often the diameter of blastholes actually to be drilled is much larger than the holes drilled by the hired rotary-cum-DTH drills. In such an event the prediction of penetration rate, bit life etc., is done by extrapolating the results in the trial drilling of 150 to 165 mm blastholes. Such a prediction is usually more reliable than that made from laboratory test data.

15.2.2 What type of drill? As stated in chapter 7 of this book electric-electric, electric-hydraulic and diesel-hydraulic are the three main types of rotary blasthole drills. Which of these types is better, depends entirely on the availability of an uninterrupted power supply without voltage fluctuations at the mine site. In most of the large mines, where shovels or draglines are to be used, a reliable power supply is certain. Therefore, in such situations technologically the best choice is electric-electric. Then comes electric-hydraulic and lastly diesel-hydraulic.

DTH Drilling

Rotary with Tricone Rotary with Bit or DTH Tricone Bit

Extremely Hard Formations Very Hard Formations Hard Formations Medium Hard Formations 150

200 250 300 350 400 Blasthole Diameter in mm

450

Figure 15.1 Suitability of blasthole drilling method in medium to very hard formations.

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421

Whether this ranking stands on costing considerations must be verified carefully on the basis of several factors mentioned under article 7.7 in chapter 7.

15.2.3

How many blasthole drills?

In the mine planning stage the desired production rate is conclusively fixed on the basis of economic considerations. How much explosive in terms of kg, is necessary for proper fragmentation of 1 m3 of rock is called the powder factor. For a specific explosive the powder factor is dependent upon the rock properties, particularly compressive strength and toughness, joint structure in the rock mass, diameter of the blasthole etc. When the powder factor for blasting of the formation is ascertained, the length of blasthole needed to accommodate the specific quantity of explosive can also be calculated from the density of explosive attainable inside the blasthole. The penetration rate that can be achieved in the formation can be estimated as per the procedure described in chapter 9 in this book. In this manner, when the average penetration rate is determined the number of blasthole drills required to be procured can easily be calculated. In all such calculations the stemming length and subdrilling required must also be taken into account.

15.3

PROCUREMENT METHODOLOGY

When purchasing a rotary blasthole drill, apart from the technical aspects of the blasthole drill, many other factors must also be considered. These factors revolve around: 1 2 3 4 5

How the specifications should be drafted Erection and commissioning of the drill Manuals pertaining to the drill Operator and maintenance training Miscellaneous items needed. Elaborations of these are as under.

15.3.1

Draft of specification for the drill

Once a purchaser decides to purchase one or more blasthole drills, manufacturers of blasthole drills must be informed accordingly so they can submit a proposal. The information is often referred to as a tender or invitation to bid and the manufacturer’s proposal is called a quotation or bid. In the tender the purchaser must state: 1 The working location of the drills. Relevant details such as longitude and latitude and elevation of the place, annual rainfall, range of atmospheric temperatures, likelihood of storms, nearest town, nearest airport, nearest sea port to which the drills can be shipped.

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422


2 The type of mine i.e. whether metal mine or coal mine or lignite mine etc. 3 Properties of rock, and rockmass in which blastholes are to be drilled. Commonly a large mine has many different rocks. Details of all types are to be given. 4 Expected yearly meterage of drilling from each of the blasthole drills. 5 The mining method to be adopted, i.e. whether loading and mucking by dragline or by shovel and dumper combination, or by bucket wheel excavators. In either case, more details are needed about planned excavation depth i.e. maximum depth of a blasthole and inclination of blastholes. 6 Voltage and frequency of electric supply and its continuity at the worksite. 7 The number of rotary drills to be purchased. 8 Specification of the drill intended to be purchased. 9 Desired time frame for the supply of drills and their erection and commissioning at site. 10 Commercial aspects related to the purchase. Some points, as stated below, must be emphasized in the specifications. 1

2

Most of the assemblies and subassemblies intended to be incorporated in the drill should be mentioned in the specifications, but the manufacturers should also be advised to include a proposal for any additional or alternative items that may prove beneficial in the operation to be carried out by the drill. The specifications of the blasthole drill intended to be purchased should be treated as general in nature and purchasers should give a proposal for any blasthole drill that may prove better than the specified.

These points induce the manufacturers to give many ideas to the purchaser that can prove greatly beneficial.

15.3.2

Erection and commissioning of the drill

Small rotary blasthole drills, in their transport position, often have length, width and height within the limits of 15 m, 3.0 m and 5.0 m respectively. As their weight is usually less than about 30000 kg, they can be easily transported in fully assembled condition on trailers, so long as the destination is within about 1500 km and does not involve any transhipment. However, in many instances even such small blasthole drills may have to be transported by sea. For such modes of shipment, depending upon the size and weight, the blasthole drill is sent in the form of two or more sub-assemblies to save freight. When the blasthole drill reaches the worksite, the sub-assemblies must be assembled, ballast must be added and some welding must be carried out before the drill can be put to actual use at the mine. In such circumstances due consideration must be given to the following factors. Such erection and commissioning takes about 7 to 15 days and requires many facilities such as welding, grinding etc. Most usually the manufacturer sends an experienced engineer to supervise the erection and commissioning task but the work has to be carried out by the technical and non-technical labor force provided by the purchaser. Cranes of capacity up to as

Book.indb 422

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423

much as 100 tonnes and several other items may have to be used in the process of erection and commissioning. All such help has to be given by the purchaser to the drill manufacturer without any payment from the drill manufacturer to the purchaser. It is also necessary on the part of the purchaser to arrange reasonably good lodging and boarding near the worksite and arrangement of transport from residence to worksite for the engineer and other personnel involved in erection of the drill. The need for an interpreter to ease the communication between the service engineer deputed by the manufacturer and the labor force involved in erection and commissioning of the blasthole drill must also be thought of in some cases. The purchaser must consider this erection and commissioning as an opportunity and deploy diligent maintenance personnel for this job so that these persons are very well acquainted with the machine.

15.3.3

Drill manuals

Manufacturers of drills usually compile three different types of manuals for each of their rotary drill models. First one is called the Operators Manual. It gives all the necessary procedures to be adopted and precautions to be taken while carrying out actual drilling and other associated operations like leveling, tramming etc. Every driller, drilling supervisor and assistant driller must thoroughly read this manual. The second manual is called the Spare Parts Manual and contains lists of spare parts with their part numbers and the drawings of the assemblies in which the spare parts are fitted. This manual is essential for ordering spare parts correctly. Usually a third manual is also available from the drill manufacturer. It is called the Maintenance Manual. As the name indicates the manual contains detailed instructions on maintenance procedures to be carried out periodically. This manual must also be thoroughly studied by all the persons associated with maintenance of the drill. It is to be noted that if a computerized drill system is incorporated on the blasthole drill, the system has a different set of manuals altogether. Similarly separate manuals are available for diesel engines, compressors and some other major components incorporated in the drill by the manufacturer. It is the responsibility of the drill manufacturer to make sure that such manuals are made available to the purchaser of the blasthole drill. The purchaser must specifically ask for such manuals from the manufacturer. In this era of computerization the manuals are in the form of an electronic pdf file format document on a compact disc. If such is the case the purchaser should ensure that he gets three sets of such CD unless they are copyable. It is worthwhile to make at least one paper copy of the manual from such CD. Some manufacturers keep this information on their website and allow the user to connect to the website. In such cases the user should download the relevant information from the website and preserve it in the form of a CD.

15.3.4

Operation and maintenance training

By any standard, a rotary blasthole drill is an expensive item. However sturdy it may be, it can get severely abused by incorrect operation. For this reason, the drillers who operate the drill on a daily basis must be thoroughly trained.

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424


Manufacturers of blasthole drills are ideally suited for imparting training to the drillers and the maintenance personnel who are often referred to as service engineers. Part of a driller’s training is given on a drill simulator. This computerized piece of equipment is nothing but a cabin to be placed in a special room. When the driller enters the cabin, the simulator creates all the audio visual environments as if the driller is actually sitting in the operator’s cabin of the blasthole drill. The driller operates consoles in the simulator and the impression of actual drilling is given to the driller. The computer monitors whether the driller is taking appropriate operational actions and gives him instructions or warnings. Such simulator training goes a long way in training the driller ideally. Usually the training courses for maintenance personnel are different than those for drillers. Maintenance training courses are oriented towards the procedures and precautions to be adapted in effectively maintaining the drill. Sometimes these training courses also cover minor repairs of the drill. It is strongly felt that while procuring every new blasthole drill, at least one driller and one service engineer should be sent to the manufacturer for their training.

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Chapter 16

Tips for operating rotary blasthole drills

16.1

INTRODUCTION

The knowledge of rotary blasthole drilling presented so far in this book is no doubt very essential for every driller engaged in rotary blasthole drilling practice. But such need is essential because the knowledge is to be used in achieving perfection in the job. To achieve such perfection some direct tips about rotary blasthole drilling given to a driller will certainly prove beneficial. The purpose of this chapter is to give such tips under various activities of the operation.

16.2

KNOWING THE ENVIRONMENT

This is a very important step in achieving a highly satisfactory drilling operation. The driller must be well acquainted with the surroundings. If he knows what circumstances are going to follow he remains well equipped to deal with the situation. The following are the tips in this regard. 1 2 3 4

5 6 7 8

9

Book.indb 425

Get to know where and when the blasting is scheduled for the day. Find out the other equipment, power lines, cables or structures that have to be moved or avoided during movement of the drill. Know the bearing capacity of the ground surface and judge if the blasthole drill can tread over it without any difficulty. Get to know about the weather forecast in respect of ambient temperatures, rain, storms etc., before moving to the worksite to take control of the blasthole drilling operations. Know in advance of any likely but unavoidable interruptions in the drilling operations due to power cut, visit of VIP etc. Find out the grade of the slope which the drill will be working on. Usually in large surface mines the bench tops are well leveled. Think of the steps necessary to keep visitors or other onlookers and equipment at a safe distance from the drill area. Before starting any operation of the drill get to know about important characteristics of the rotary blasthole drill, particularly its weight, height, width and the restrictions in its operation. Make sure that the blasthole drill is equipped with fire extinguisher and first aid kit.

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426

16.3


MOVING THE ROTARY BLASTHOLE DRILL

Rotary blasthole drills have to be moved several times per day. Usually such travel is very short in distance – merely 10 to 20 m every time because it is from one completed blasthole to the location of a new blasthole. It takes anywhere between 0.5 to 2 minutes. For such operation the following tips are of great importance. 1 Before starting the movement of the drill ensure that the drill bit is withdrawn well above the ground level. Also ensure that all the three or four jacks, as the case may be, are lifted well above the ground. 2 A rotary blasthole drill invariably has its center of gravity at a level much higher than that of a dozer, grader or other construction equipment. Hence it is far more sensitive to the unevenness of the ground. Therefore, ensure that the tramming surface is very well leveled and the blasthole drill will have no difficulty in tramming on the ground. It is to be started and moved slowly. 3 Ensure that the path over which drill has to travel must be free from ditches, boulders etc. 4 Ensure that all the on-board equipment on the drill is securely fastened to upper or mast as the case may be. Be particularly attentive to the locking arrangements of drill pipes. 5 Before starting movement of the drill, make sure that no one is in the machinery house or on any other part of the drill. 6 Some blasthole drills, in order to lower their center of gravity before starting to travel, may require uncoupling of all the drill pipes and lower the drill head to a level sufficient only to hold the drill bit and stabilizer. If your blasthole drill is of this type make sure that these operations are completed before tramming. 7 Make use of the audible alarm before starting the movement of the drill to alert people in the area. 8 Whenever needed let the assistant give guiding signals so the movement of the drill is safer. 9 The trailing cable of a blasthole drill always poses problems to a driller. Make sure that the winding mechanism of the cable reel is working appropriately. Maintain cable slack while moving. Let the assistant give signals while moving so the damage to the cable or cable connection is prevented. 10 If your blasthole drill has a slewing upper ensure that the upper is slewed back to its recommended position and securely locked before starting ground movement. 11 Look and listen for falling rocks when tramming near banks or high faces. 12 If the blasthole drill is equipped with a remote tramming device, tram the drill in remote tram mode rather than using the tramming controls in the drill. Less frequently but periodically a blasthole drill has to be moved over longer distances to allow charging and blasting. The distance to be traveled at a stretch may be long – of the order of 500 m or so. For such travel operations remember the following.

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1 2 3 4

427

Make sure that the manufacturer of the blasthole drill does recommend travel of the drill over such distance at a stretch i.e. without any intermediate pause. Give a warning, well in advance, to all the operators of other equipment in the area about such travel. Make sure that the path to be traveled by the blasthole drill is level and without any road obstacles like boulders, ditches etc. Make sure if the blasthole drill has to pass under power transmission lines. In such conditions, for the sake of safety, any part of the drill, including the mast, must not be allowed to pass through a certain minimum distance from the transmission line as indicated in Table 16.1.

On a few occasions a blasthole drill has to be moved from one bench to the other bench. In such operation it has to travel over a gradient that is as high as 10%. In some drills the specified gradeability is far more than 10% but this figure is based on the power rather than stability of the drill. Due to these factors a driller must remember the following. 1

2

3 4

5

6

Check the ground you intend to travel for stability. When traveling off-road, first walk the route of travel, inspecting for depressions, stumps, gullies, ruts or other obstacles. This is particularly important when working on steep slopes or near a bench face. No attempt should be made for propelling the drill on such slope without lowering the mast and bringing the drill head in a position recommended by the manufacturer. It must be remembered that stability of the blasthole drill may be different while traveling in a forward or backward direction. While moving from one bench to another, many turns may have to be negotiated. While tramming in such conditions, the driller must allow for mast overhang when approaching corners, other vehicles or structures. Driller must give careful look to the canopies of service stations and rest rooms near the path to be traveled. These structures are most often too low for clearance of the mast. Always remember that a blasthole drill is not a rough terrain vehicle. Table 16.1 Distance from the power line within which no part of the blasthole drill must come while traveling.

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Voltage in power line in V

Minimum clearance of line from ground in m

Minimum allowable distance of any part of the blasthole drill (OSHA regulations) in m

6600 to 11000 11000 to 33000 33000 to 50000 50000 to 132000 132000 to 275000 275000 to 500000 More than 500000

5.2 5.2 5.2 6.7 6.7 7.0 7.0

3 3 3 3.35 3.35 4.0 5.5

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16.4

SETTING UP THE ROTARY BLASTHOLE DRILL

Positioning the rotary blasthole drill at a precise point on the ground, and with precise alignment of the mast, is of utmost importance. Imprecise setup of the drill can give extremely unfavorable fragmentation of the rock mass. The following tips will help in avoiding such possibilities. 1 2 3

4

5 6 7

Inspect the ground for lifting support and wherever necessary add cribbing or support mats before lowering jacks. The drill must be leveled in such a way that its crawler tracks are only about 200 to 300 mm above the ground after machine leveling. When lowering the machine to the ground, always lower the machine slowly and in stages, maintaining near level condition until the crawler tracks contact the ground and support the machine fully. Always take the help of the assistant in giving appropriate signals because he can observe the position of the drill, hydraulic jacks and crawlers far more vividly by standing on the ground. While drilling blastholes lying in the first row from the edge of the bench, ensure that the center of the blasthole drill is as far away from the edge of the bench as possible. Watch for any cracks in such vulnerable areas. These cracks may be indicative of backbreak or unstable ground conditions. Start drilling blastholes in such a sequence that the drill will not have to tram through an area where blastholes have already been drilled.

16.5

OPERATING A ROTARY BLASTHOLE DRILL

It is once again reiterated that the driller and his assistant must carefully read and understand the methods of operation and maintenance manuals provided by drill manufacturers. Study of the manuals of important components in the blasthole drill e.g. engine, electric motors, compressors, hydraulic system etc., is also essential. The following tips, addressed to the driller, will go a long way in getting most beneficial results. 1 Reach the blasthole drill at least ten minutes before the start of your duty. When you reach the drilling site carefully observe the heaps of cuttings around the blastholes already drilled. They will give you a good idea about the rock mass, seepage of water etc. 2 A few minutes should also be spent in knowing how drilling and other activities progressed during the previous shift. Have attentive dialogue with the departing driller and assistant. 3 Before taking charge of the drill, check all the gages and indicators to make sure that the drill has an adequate quantity of engine oil, compressor oil, fuel and other essential commodities. 4 While starting a new blasthole operate the drill at low bit pressure and low rotary speed.

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5 If there is likelihood of seepage of water in the blasthole, operate the compressor at full air pressure. 6 Keep the drill string rotating at low speed while withdrawing it after completion of the blasthole or lowering it after changing the drill bit. 7 Ensure that the drilling parameters, such as weight on the bit, rotary speed, air flow etc., are within the limits recommended by the drill or drill bit manufacturer. 8 Do not lower the drill string into the blasthole at fast speed because its impact on bottom is the most common cause of bit damage. 9 If rotation of the drill string has been stopped for some reason, before resuming the rotation ensure that air flushing through the bit is restarted. 10 Always remember that rotary speed is inversely proportional to the bit weight. 11 If the drill is equipped with an automatic lubricating or leveling system, keep an eye on the working or calibration of the system from time to time. 12 Ensure that air pressure is kept at an appropriate level while drilling. 13 Avoid the use of an excessive quantity of water for dust control as it shortens bit life. 14 Long, slender items like drill pipes, Kellys etc., are very susceptible to bending or thread damage. Therefore, before taking them into the drilling operation check that their threads are not damaged or they are not bent beyond acceptable tolerance limits. 15 If a new bit is used for completing a half finished blasthole, the gage diameter of such new bit will be more than the diameter of the blasthole. For this reason the cones, bearings and gage teeth of the new bit will get damaged. To avoid such damage use a reconditioned bit to complete the hole. 16 Ensure to follow the recommendations made by the manufacturers in respect of oils, lubricants and other consumables to be used. 17 Keep a careful watch to ensure that making and breaking tool joint operations are being performed smoothly. 18 If the drill has been idle for a long time for maintenance or repairs and drilling is to be started once again, clean the bit by passing air or water through it while rotating the cones of the bit by hand. 19 Let the assistant driller inspect the bit periodically for damage or impending failure. If there is considerable difference between the temperatures of two cones, it is a sign of some obstruction in the bearing and likely bit failure in the near future. Appropriate corrective action to avoid such failure must be taken. 20 Do not use the lifting cable run on the mast and wound by the auxiliary winch for pulling distantly placed accessories. The mast of a blasthole drill is not designed for such activities. 21 If stormy weather or other adverse conditions are impending, continue the operations cautiously and stop at an appropriate moment. 22 During drill operation, keep all people off the drilling platform and drill mast, and away from drill stems. Moving components or rotating drill stems can entangle clothing and can crush, pinch or strangle personnel. 23 Do not allow anyone to tie any object to the mast or hang any object on the mast by means of small hooks. It can prove disastrous while the mast is lowered or lifted.

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24 Before lowering the mast, notify personnel to evacuate the roof and drill deck and inspect the mast storage area for obstructions. After storing the mast, check for secure mast attachment on the front jack caps. 25 While drilling is in progress do not allow anyone to stand on the drilling platform. 26 If you have a feeling of uneasiness or drowsiness etc., do not attempt to operate the drill. Medications that induce drowsiness or headache must not be consumed. Do not consume alcohol or drugs prior to or while on the job. 27 You must not allow unauthorized persons to come near the blasthole drill. Even if an authorized person has to be taken on the blasthole drill he should be kept away from the control console for their safety, your safety and the safety of the drill. 28 If you are a fully trained, authorized and experienced driller, only then with the permission of management may you undertake to impart primary training to the helper assistants. This training must only be verbal and not hands-on. The helper assistant must spend at least 500 hrs in primary training on the blasthole drill. 29 You must not leave the operator’s cabin when the drill is in operation. You must not handle the controls with oily or greasy hands so as to avoid accidental slip in the grip and consequent disaster. 30 If you have to leave the operator’s cab, before doing so you must ensure that the feed and rotation controls are brought into neutral position and then the engine has been disengaged from the power train. 31 You must not carry out drilling in electrical storms, high wind conditions or conditions of poor visibility, unless such operation is acceptable owning to automatic safety devices incorporated in the drill. 32 If for some reason you have to drill a blasthole near another blasthole that is either partially of fully loaded with explosive, make sure that alignment of such new blasthole is not near to the loaded blasthole by less than 6 m in any of its part. 33 Before climbing the blasthole drill ensure that grease, oil etc., are wiped off from your shoes so the staircase, cabin floor or catwalks do not become slippery. 34 Hold each of the two handrails by each of the two hands while climbing up the staircase of the blasthole drill. 35 When a tool is to be carried on the blasthole drill, first keep the tool on the catwalks, then climb the staircase and then lift and carry the tool to the appropriate point. 36 Keep an eye on the exhaust of the dust collector to ensure that its pollution level is not exceeding the visible limits. If the dust collector is not working within such limits and drilling has to be continued, make use of the water injection system. 37 At the end of the shift do not stop work; continue the drilling operation until you handover the charge to the driller in the next shift. When drilling has to be discontinued in the next shift ensure that the blasthole is completed to its final depth, and drill pipes as well as drill bit have been removed from the blasthole. 38 Ensure that the tool-handling wire rope has a hook with safety latch at its end. 39 Do not exceed the safe working load capacity of the auxiliary winch meant for tool handling. 40 In cases when the drill string gets stuck in the blasthole, try to pull it up only with the feed mechanism. Do not think of giving additional pullout force by simultaneously lifting the hydraulic leveling jacks.

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431

41 Do not play around or play with the blasthole drill or any of its components. 42 While drilling blastholes lying in the row near the toe of the bench floor on which the drill is standing, ensure that center of the blasthole drill is as far away from the cliff as possible. 43 Remove the starting keys of the drill engine when the drill has to be left unattended during a no-work period or when the engine is to be stopped for drill maintenance.

16.6

CONTROLLING DUST EMISSIONS

As stated earlier, the dust and debris formed at the bottom of the blasthole while drilling and ejected from the blasthole are always hazardous if allowed to freely mix with atmospheric air. Breathing in such polluted air can cause one or more respiratory disorders depending upon the contents of the dust. The following steps and tips are necessary to avoid such happenings. 1 2 3 4 5

Inspect the dust hood curtains from time to time and ensure that they are not torn and are enclosing the surroundings of the blasthole adequately. Before collaring make sure that the curtains are lowered and effectively prevent the drilling dust from mixing with the atmosphere. While the drilling is in progress do not lift the curtains to remove the cuttings. Avoid contamination of work clothes, drinking water etc., by the drilling dust. Rigorously follow mine procedures for air monitoring, exposure limitations, and means of protection from crystalline silica exposure.

16.7 TRANSPORTING THE DRILL When a blasthole drill has to be moved from one mine to another, it may have to be dismantled to a varying degree depending upon the size of the drill. This becomes necessary to comply with the regulation of the road authorities. For such transportation of the blasthole drill, the following factors must be given due consideration. 1

2

3

4 5

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If the blasthole drill has to be loaded on a trailer for long-distance travel, choose a low bed trailer so the drill on the trailer will easily pass under overhead obstacles. Some drills must be dismantled before transportation. Ensure that the center of gravity of the blasthole drill is kept on the center line of the trailer. If the trailer is a semi-trailer ensure that the blasthole drill is positioned in such a way that adequate weight is transferred to the tractor through the fifth wheel. A ramp that has adequate capacity to bear the weight of the drill must be used for loading the blasthole drill onto the trailer. During such loading operation make sure that movement of the trailer is adequately blocked so it will not move when the drill is trammed onto it from the ramp. Always take guidance from an assistant on the ground when you are loading the drill on a trailer from a ramp. Firmly secure all the loose items on the drill to appropriate points on the drill and the drill to the trailer before moving the trailer.

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16.8

MAINTENANCE ACTIVITIES

The precautions to be taken by service personnel are expressed by the following advice or commands: 1 Carry out a pre-shift safety and maintenance inspection as per the specific form provided for the purpose. 2 Do not smoke during any maintenance or fuel fill procedure. 3 Use gloves when handling wire rope. 4 When new rope is used for the first time, lift a light load with it. 5 While replacing a wire rope ensure that new wire rope is exactly as per the specifications of the wire rope that is being replaced. 6 Use of hands, even with gloves, for guiding the wire rope when it is being wound onto the winch drum is very dangerous. 7 Ensure that drilling accessories are not lifted and kept hanging in the midst of the air unattended while they are being handled by the tool handling winch. 8 Carry out all the maintenance procedures as per the instructions given by the component manufacturers e.g. drill, compressor, diesel engine, electric motors, hydraulic motors, hydraulic pumps, cables, hydraulic hoses etc. 9 Always wear safety glasses or face shield while servicing the batteries. 10 Maintenance of batteries is to be carried out in a well ventilated place. 12 If battery acid droplets splash on your body wash the body area with plenty of water. 13 When charging a battery, turn off the power source to the battery before connecting the charger. Cell caps should be loosened to allow the gases to escape. 14 All rotating or moving parts, such as drive lines or fan belts, should be covered. Be sure to replace guards after maintenance. 15 Flammable substances should be transported in cans that are specifically accepted by the Fire Prevention Agencies.

16.9

PERIODICAL CHECK UPS

A blasthole drill manufacturer designs and writes the operation and maintenance manuals through the experience gained about the operation of their rotary blasthole drills on worksites. From time to time such manuals are revised and provided with additional information. Invariably such manuals have several instructions about periodical check ups. Some of the check ups are to be performed on a daily basis whereas others are to be performed on a weekly, monthly or half yearly basis. These check ups and remedial actions suggested in the manuals must be followed strictly so as to maximize the life of the blasthole drill.

16.10

DRILLING RECORDS

Drilling records are very important for planning further drilling and blasting activities. No doubt the best records are kept by a computerized drilling system or to a lesser extent by a logging system.

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Table 16.2 Drilling record. Indian Copper Ltd.

Head Office: 1732, Uttung Towers, 9th Road, Andheri, Mumbai-400 053, INDIA Mines - Anypore, Bihar; Malanjkhand, Madhya Pradesh. Daily Drilling Report

Date :- April 14, 1997

General Information and Parameters Common to All the Blastholes Mine Location

Malanjkhand Pattern Name

10C

Pattern Altitude (m)

980

No. of Holes in Pattern

330

Ore Name

Chalcopyrite Av. UCS of Ore (MPa)

200

Av.DRI of Ore

75

First Shift Start Time (hh)

07.00

Waste Type

Granite

Av. UCS of Waste (MPa)

220

Av.DRI of Waste

68

Second Shift Start Time (hh)

15.00

Pattern Type

Triangular

Blasthole Diameter (mm)

349

Burden (m)

Spacing (m) 85.5 Comp. Air Pressure (kPa)

Blasthole Drill Make and Model

Bucyrus Intl. 39R

Maximum Bit Weight (kN)

540

Bench Height (m)

16.5

Inclination 0

15

Maximum Rotary Speed (RPM)

160

First Drill Pipe Length (m)

13.7

Next Drill Pipe Length (m)

13.7 Drill Pipe Dia (mm)

Max. Compressor Capacity (m3/min)

Blasthole Length (m)

448 17.09 273

Serial Blasthole Time In Time Out Bit Bit Sr. Pulldown Rotary Speed Vibration Level General Comment or No. No. (Hrs) (Hrs) Type No. Range (kN) Range (RPM) (High/Med/Low) Comment About Delays If Any. 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 Specific Comments

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But if such a system is not incorporated in the rotary blasthole drill, as is often the case in small rotary blasthole drills, the driller must keep records of the drilling activities. For this purpose a sample form is shown in Table 16.2.

16.11

SAFETY PRECAUTIONS

Nowadays, blasthole drills have become very safe due to numerous improvements made in the last six decades. However, careless operation with no concern for safety, can lead to disasters and occasionally even fatal accidents. To avoid such instances, precautions must be taken by the driller as well as the assistant while carrying out blasthole drilling, maintaining or repairing the drill, and traveling or tramming the drill to the next blasthole position. Manufacturers of rotary blasthole drills have vast experience with which they write and update their operation and maintenance manuals. These manuals contain a lot of operating as well as maintenance instructions and precautions to be taken while drilling as well as maintaining the drill.

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Chapter 17

Mechanics of blast fragmentation

17.1

INTRODUCTION

Previous parts of this book dealt exclusively with rotary blasthole drilling, as is practiced in large surface mines. In mining operations the drill and blast operations are so closely associated with each other that a book on either subject is likely to be considered incomplete without adequate coverage of other. Therefore, chapters hereinafter cover theoretical and practical aspects of blasting in large surface mines. Blasting is a very devastating form of rock destruction. As these operations are extremely dangerous, they are subjected to very stringent regulations imposed by controlling authorities – usually Governments of the countries. These regulations not only apply to blasting operations but also for transporting, storing and handling of explosives. All the persons carrying out blasting operations are required to be specially trained. Without a certificate from an appropriate training institute, nobody is allowed to undertake blasting operations. Blasting is not only dangerous in practice but even as a subject it is far more vast and relatively complicated than drilling. In fact, this is the reason for which almost every other book on drilling and blasting published so far has laid far more emphasis on blasting. Adequate knowledge of blasting is very important for everybody involved in blasthole drilling. This chapter is devoted to elaborating different aspects related to the process of explosion.

17.2

EXPLOSION PROCESS

Any inflammable material filled in a hole in the ground, will burn continuously so long as the supply of oxygen necessary for the burning process continues. The process of burning propagates at a certain speed from the point of initiation to the end point after which no more inflammable material is available for burning. When an inflammable material burns, the most common byproducts are oxides. Therefore, oxygen is very essential for the process of burning. If a material burns by taking oxygen from the atmosphere, it does not burn quickly. But if it generates an adequate quantity of oxygen by self-decomposition, it can burn much faster.

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Actual burning always takes place within a small layer confined by two surfaces. This layer always moves to burn new material. Behind this layer the material is in burnt form and ahead of the layer the material is in unburnt form. The thickness of the burning layer depends upon the burning material itself and also on the physical extent of the burning material in continuum. Deflagration and detonation are the two distinct types of burning. When the speed of propagation of the burning layer is less than the speed of sound, the burning is called deflagration. In deflagration the burning layer advances by transferring heat energy to the adjacent layer in the unburnt portion. When the speed of propagation of the burning layer is in excess of 2000 m/s the burning is termed detonation. In detonation the burning layer advances by transferring energy through a shock wave. The speed of propagation of the burning front depends upon the intensity of heat with which the burning process has been started and how quickly, and also how much oxygen the burning process gets. Therefore, some explosives merely deflagrate when brought in contact with a common match-stick fire but detonate and explode if brought in contact with a spark that has very high intensity of heat. Usually all materials that detonate are called explosives. Detonation of all the explosives results in a very large volume of gases and a very large quantum of heat. Most of the explosives are chemicals formed by carbon, hydrogen, oxygen and nitrogen. When they detonate, they mainly generate common and harmless chemical compounds such as H2O (Water), CO2 (Carbon Dioxide), N2 (Nitrogen) with a very small quantity of poisonous gases such as NO (Nitrous Oxide), CO (Carbon Monoxide), NH3 (Ammonia), CH4 (Methane). Figure 17.1 shows the propagation of a burning layer in a blasthole that has uniform diameter through out its length. In the case of explosives the velocity of propagation of the burning layer is called the velocity of detonation. The frontal surface of the burning layer, more commonly called the detonation zone, is termed the shock front and the rear surface is called the Chapman-Jouguet surface. The length of the detonation zone depends upon the components of the explosive, particle size, density and confinement of the explosive. The width of the detonation zone is proportional to the length of detonation zone. For every explosive there is a certain blasthole diameter called the ‘critical diameter’. If the particular explosive, filled in a blasthole having a diameter smaller than the critical diameter is detonated, there is a likelihood that detonation of the

Products of Detonation

Width of Detonation Zone Length of Detonation Zone

Detonated Zone

Shock Front Yet to detonate Zone

Point of Initiation C-J Surface

Figure 17.1 Propagation of detonation.

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437

explosive will be incomplete and blasting will be ineffective. Thus for effective use of the explosive, the actual blasthole diameter must be more than the critical diameter. Dangers of blasting operations are many and arise From: 1 2 3 4 5

Generation of poisonous gases Throw of rock pieces in the air Pollution of atmospheric air by dust Vibrations caused in the ground mass Propagation of air shock waves.

Improper transportation, storing and handling can cause even more dangerous explosions than the explosion engineered through proper blasting operations. Accidental explosions are always unexpected and without any control, whereas the proper blasting operations are expected and controlled.

17.3

FORMATION OF A CRATER BY A BLAST

The most fundamental action behind disintegration of the rock mass by a blast is the formation of a crater. A crater is formed when a projectile of any heavy object impinges upon a rock mass. Such a crater is observed very commonly. A crater is also formed when an anchor embedded in the rock mass is subjected to a pull-out force that is sufficient to break the rock mass. Such a crater is shown in Figure 17.2. The pullout force actually exerts pressure on the rock mass through the plate as shown by small arrows in the figure. These forces in turn generate shear stress in the rock mass along the surface AB. Eventually the failure occurs at this surface when the stress exceeds the strength of the rock mass. When an explosive, confined in a spherical cavity inside a rock mass, is detonated, the pressure created inside the cavity exerts forces in the radial direction. Further, the extreme rapidity with which the pressure inside the cavity is generated causes a compressive stress wave moving radially in all directions. As the compression stress wave travels, the magnitude of stress decreases. When this compressive stress wave F R

B

d

A

Figure 17.2 A crater formed by pullout force exerted on an anchor embedded in rock mass.

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reaches the free surface it rebounds backward in the form of a tensile stress wave. The magnitude of tensile stress in this wave depends upon the compressive stress wave. As a result of the tension wave, the failure of the rock mass occurs in the volume enclosed by the free surface and the curved surface in the form of a crater shown in Figure 17.3. The depth and volume of the crater is at a maximum for a certain depth of the explosive cavity below the free surface. This depth is called the critical depth dc. When the depth of the explosive cavity is much higher than the critical depth, no fragmentation at the free surface takes place and there is no crater. As the depth of the cavity of explosive below ground level d approaches dc, the crater forms and the volume of the crater keeps on increasing till d = dc. All this is shown in Figure 17.3. When d = dc the corresponding crater is called a full crater. The radius R1 on the free surface, in the case of a full crater formed by an explosion, is much larger than the radius R observed in the crater formed by anchor pullout. This is shown in Figure 17.4. The critical depth depends upon the properties of the rock mass as well as properties and weight of the explosive. The depth dc for full crater increases with increasing mass of the explosive, W. A relation between the two is as under. n

W ∝ dc where

W = Mass of the explosive dc = Depth of the full crater from the free surface.

d=dc d >>dc

d>dc

Figure 17.3 Progressively increasing depth of crater formed in the rock mass by explosive in a cavity below free surface.

F

R1 BB

d=dc

AA

Figure 17.4 Geometry of a full crater.

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439

100 80 60 40

Log W (Weight of Charge in lb)

20

10 8 6 4

2

1.0 0.8 0.6 0.4

0.2

0.1 0.2

0.4

0.6 0.8 1 2 4 6 8 10 Log d ( Depth of Charge in ft)

20

Figure 17.5 The plot of relationship between weight of explosive w and critical value of depth of explosive.

If the explosive is contained in a long horizontal blasthole instead of a spherical cavity, then the depression made in the ground by the explosion is in a long canal-like shape rather than a conical crater. Experiments were carried out on a gneiss rock mass at Idaho Springs to ascertain various parameters of the geometry of a full crater. The plot for various values of mass of explosive and corresponding depth of the crater was as shown in Figure 17.5. It was found that the value of n is 1.85 for the particular rock mass. It was also found that R1 = 2.8dc. Experiments have indicated that the value of n is between 1.5 for elastic and 3.4 for plastic rocks.

17.4

BLAST FRAGMENTATION MECHANISM

A very rudimentary idea of rock fragmentation by blasting is contained in the first chapter of this book. It needs to be elaborated.

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The many disintegration modes that contribute to the process of rock fragmentation are as under. 1 2 3 4 5 6 7 8

Crushing of Rock Radial Cracking Circumferential Cracking Spalling Radial Push by Gases Cracking by Flexure Collision of Fragments Separation of Beds Each of the above points is elaborated below.

17.4.1

Crushing of rock

Explosive filled in a blasthole is called the charge. The process of systematically filling charge in a blasthole is called charging. Initially when the explosive charge is detonated by a spark at the bottom of the blasthole the detonation zone starts moving up the blasthole at a speed of 2500 to 7000 m/s. The energy released and the gases generated in the detonation zone exert an extremely high pressure against the walls of the hole. Simultaneously the heat generated in the detonation gives rise to very high temperatures. The formulation of velocity of detonation, detonation pressure and detonation heat is dealt with in one of the forthcoming sections. In most cases the pressure ranges between 7 to 10 GPa and temperature ranges between 2500 to 4500°C. Such high temperature and pressure act together to pulverize and crumble a thick cylindrical layer around the blasthole wall. This is called the crushed zone. The thickness of the crushed zone depends upon the magnitude of the detonation pressure, heat of detonation, and the strength and porosity of the rock. All this culminates in an increase in the volume of the blasthole. If VO is the original volume of the blasthole and VE is its volume after explosion, the increase in ratio VE/VO with time is as shown in Figure 17.6. This ratio is of the order of 2 to 4 for strong rocks and of the order of 10 for weak and porous rocks. As much as 30% of the energy generated in the detonation zone is spent in the crushing of rock.

17.4.2

Radial cracking

A compression wave is also created at the detonation zone. It starts moving away from the blasthole in all directions at the velocity of a sound wave in that medium, as shown in Figure 17.7A. All rocks are formed by minerals of different composition. Therefore, planes of weakness are left during the process of formation of a rock mass. The cracks, which are in a direction tangential to the wave propagation, remain nearly unaffected by the compression wave. The compression wave generates shear stress along the radial planes of weakness because the response of different crystals to compressive stress

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Blasthole Volume Ratio VE /VO


441

10 8 6 4 2 1

1

2

3

4 5 7 10 20 Time in milliseconds

30 40 50 70 100

Figure 17.6 Increase in blasthole volume after detonation. A Detonation wave moving along the blasthole at 2500-7000 m/s Compression wave moving away from the blasthole at 1000-5000 m/s

B Crushed Zone

Original Diameter of the Blasthole

Region of Slightly Enlarged Weakness Diameter of the Blasthole

Figure 17.7 Initiation of rock fracture by blasting.

differs. The number of cracks in the region adjacent to the walls of a blasthole is higher, but at longer distances from the blasthole the formation of cracks is lesser as shown in Figure 17.7B. This happens because there is a significant reduction in the energy contained in the compression stress wave. As the rock mass on the rear side of the blasthole is in horizontal confinement, there is no significant radial cracking on the rear side of the blasthole.

17.4.3

Circumferential cracking

Within a very short period the compression wave reaches the free surface at the bench face and gets reflected as a tension wave. The speed at which the tension wave travels backward is about half of the speed of the compression wave, i.e 500 to 2500 m/s. The lower speed of the tension wave is due to the fact that a significant part of the energy contained in the compression wave is lost as it moves to the bench face. The quantum of energy lost depends upon many factors such as composition of the rock mass, the distance between the blasthole and the free surface nearest to the

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blasthole at the time of detonation (i.e. the burden), the distribution of pre-existing cracks in the rock mass, the porosity of the rock mass etc. Tension waves are very important from the viewpoint of blasting because they generate circumferential cracks in the rock mass as shown in Figure 17.8A. Without such circumferential cracks there cannot be any significant fragmentation in the rock mass. The portion between the blasthole and the bench face is heavily cracked by the tension wave as shown in Figure 17.8B. The reason for large scale circumferential cracking in the rock mass by the tension wave is that all the rocks have very low tensile strength as can be seen from Table A14.5 in Appendix 14. Since there is no free surface on the rear side of the blasthole there is no circumferential cracking on the rear side.

17.4.4

Spalling

When a compression wave reaches the bench face and reflects as a tension wave, in the very initial stage circumferential radial cracks develop near the bench face. Since there is no confinement of the rock fragments on one side of the bench, they separate from the bench face and fall down. This fragmentation is called spalling.

17.4.5

Radial push by gases

After the radial and circumferential cracking the rock mass on the front side of the blasthole is already loosened, but remains without significant movement. Gases formed in detonation and held at exceedingly high pressure start flowing to the circumferential cracks through the radial cracks and exert outward pressure. This causes throw of the rock fragments in an outward direction i.e. away from the bench face as shown in Figure 17.9.

17.4.6

Cracking by flexure

In a blasthole the detonation process is almost invariably started at the bottom of the blasthole. In the initial stages gases form at the bottom and exert upward pressure. A Circumferential Cracks Crush Zone

B Large Scale Cracking of the Rock Mass in Front of the Blasthole

Rock Fragments Formed at Bench Face by Spalling Enlarged Diameter of the Blasthole

Figure 17.8 Rock mass cracking by tension wave.

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Broken Rock moving away at about 50 m/s or even higher speed

Figure 17.9 Throw of rock fragments.

Figure 17.10 A blast in a large mine.

As the detonation layer moves upward, the gases formed in the upper region also add to this pressure. Thus, the pressure in the middle of the blasthole length is higher than the pressure at the bottom. This differential pressure creates a bending moment on the burden wall as shown in Figure 17.9 and as is evident from the photograph in Figure 17.10. With such flexure given to the burden wall, horizontal cracks form in the wall on the outer side i.e. near the bench face.

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A photograph of a blast carried out in a mine bench is shown in Figure 17.10. Minute observation of the photograph reveals that the rock mass portion towards right side has bulged more whereas the rock mass at the middle portion of the photograph has not bulged to that extent. This is because the middle portion actually represents the boundary between the non-blasted area on the left side of the photograph and the blasted area on the right side. Further, the right side portion appears to have vents through which explosion gases have escaped along with fine material. The fine material has formed a clearly visible cloud of dust.

17.4.7

Collision of fragments

When fragments formed by blasting are thrown away from the rock mass in burden, they collide with each other. Such collisions result in further fragmentation of the pieces of rock.

17.4.8

Separation of beds

If the rock mass consists of different layers, as usually happens in sedimentary rock formations, the response of different layers to the action of the compression wave is different. For this reason, in many cases the beds on two sides of the bedding plane get separated by cracks formed at the bedding plane. Such cracks invariably add to the process of fragmentation by blasting. The speed at which the fragments are thrown away depends upon the pressure and rock mass characteristics. All the modes of rock fragmentation described in earlier sections take place in such a short time interval that it is very difficult to give any sequence.

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Chapter 18

Properties of explosives

18.1

INTRODUCTION

The sole purpose of drilling blastholes is to fill them with explosives and blast them so that the shock wave formed, the gases generated and the heat evolved by the explosion after blasting the holes causes fragmentation of the rock mass. Thus, for blasting, the most important material is explosive. This chapter is devoted to giving broad information about different properties of explosives. Explosives used in large surface mines are the center of attention in the treatment of the topics. Tests carried out for evaluating the properties are dealt with separately in chapter 22.

18.2

EXPLOSIVES

Many materials burn. Whenever they burn, there is a layer in which the burning action is in progress. On one side of the layer the material is in burnt form. The burning layer propagates towards the other side where the material is yet to be burned. The speed of propagation of the front surface of this layer differs from material to material. Based on the speed of propagation, five classes of burnable materials viz. combustible, inflammable, highly inflammable, low explosives and high explosives are usually recognized. Table 18.1 clarifies the differences in these classes. For a chemical to be considered as an explosive, it must exhibit the following characteristics. 1 2 3 4

Rapid expansion Quick liberation of heat Fast reaction Need to initiate the chain of reactions

Rapid expansion can occur when a material burns to evolve gases. Thus, an explosive must be largely composed of gas-forming elements viz. carbon, hydrogen, oxygen and nitrogen. Further, for the rate of burning to be high, either the particles must have sufficient oxygen content so the chemical action readily gives oxygen, or the size of particles of the material must be very fine so that the atmospheric oxygen is available in close vicinity of a very large surface area.

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Table 18.1 Classes of burnable materials. Class

Classification norm (m/s)

Typical material

Combustible Inflammable Highly Inflammable Low Explosive High Explosive

Less than 0.01 Between 0.01 to 1 Between 1 to 700 Between 700 to 2000 More than 2000

Wood, Coal Kerosene, Paraffin Petrol, Sulfur, Magnesium Black Powder Nitroglycerin, TNT

Coal burns slowly in atmospheric air. If it is finely powdered it can burn in air very rapidly. If it is mixed with liquid oxygen it can burn violently and cause an explosion. Quick liberation of heat in the chemical reaction can occur depending upon the composition of the material. The gases formed in the chemical reaction expand and have devastating effect only when quickly-released intense heat is available. A kilogram of coal has five times the calorific value of one kilogram of nitroglycerin. However, detonation of nitroglycerin causes an explosion and gives much higher shattering effect than the slow burning of coal. Rapid expansion of the material and quick liberation of gases can take place only when the chemical reaction takes place very rapidly. For a chemical to be considered as an explosive, the explosion must occur only when an instance designed by a human occurs. Without such a characteristic the explosive cannot be very useful. The process of very rapid burning of the low explosives is called deflagration, whereas the process of lightning-fast burning of high explosives is termed detonation. Low explosives do cause a heavy push or powerful lift of the material that surrounds them but do not cause a shatter. Often, low explosives are mixtures of a combustible substance and an oxidant that decomposes rapidly to give a high quantity of oxygen for the quick burning. High explosive materials decompose very rapidly through detonation under certain circumstances to evolve a huge volume of gases, extremely high quantity of heat, rapidly traveling shock waves in atmospheric gases as well as the ground rock mass, light, deafening noise and hazardous flyrock. The build-up of pressure on the walls of a blasthole upon blasting low and high explosives is shown in Figure 18.1. In this book we will be concerned with high explosives as they are used for rock blasting. High explosives are further classified as shown in Figure 18.2. Primary explosives are very easily influenced by heat, friction or shock to cause detonation. Commonly used primary explosives are mercury fulminate, lead styphnate and lead azide. These are also termed as initiating explosives. PETN is a benchmark explosive. All explosives more sensitive than PETN are primary explosives. Secondary explosives are relatively insensitive to heat, shock or friction. They are also called base explosives. Since they are not easily influenced by

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Pressure of Detonation


447

Detonation of Powerful High Explosives like Nitroglycerin, PETN, RDX etc.

Detonation of Intermediate High Explosives like ANFO, Pyroxylin etc.

Detonation of Low Grade High Explosives like Phenobel, Tetracene etc. Deflagration

Time

Figure 18.1 Build-up of pressure on the walls of blasthole. High Explosives (Detonating)

Primary Explosives

Initiating Explosives

Secondary Explosives

Military High Explosives

Commercial Explosives

Tertiary Explosives

Commercial Blasting Agents

Figure 18.2 Classification of high explosives.

heat, friction or shock, they can be cautiously used in large quantities during warfare as well as rock blasting. Some secondary explosives may burn when exposed to heat or flame in small, unconfined quantities, but detonation can occur. Secondary explosives usually need a small device containing a very small quantity of primary explosive for their detonation. Most of the secondary explosives are chemical compounds rather than mixtures. Explosives, such as dynamite, TNT, PETN, RDX, HDX etc., are mainly used for warfare activities. Some of these, particularly dynamite, were used for rock blasting in the early days but now they are being replaced by tertiary explosives.

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Tertiary explosives are those which are very insensitive to shocks. They cannot be reliably detonated by a limited quantity of primary explosive. Therefore, the detonation device used for them contains a small quantity of secondary explosives. Explosives in a confined state that cannot be detonated by means of test blasting cap no. 8 are classified as tertiary explosives. Tertiary explosives are also called blasting agents. Tertiary explosives contain primarily an inorganic nitrate and a carbonaceous fuel. For certain purposes they also contain non-explosive substances like powdered aluminum. The most commonly used tertiary explosive is ANFO. Almost 80% of mining blasts are made by using it. ANFO is actually a mixture of ammonium nitrate and fuel oil. Slurry or “wet bag” explosives are also tertiary explosives. Apart from the above classification, explosives are also classified on other criteria such as their consistency or form. Permissible explosive is a term used to refer to low explosives which are somewhat freely permitted for use. They are often used for making fire crackers and entertainment fireworks. Manufacturers add several other ingredients to the explosives for some specific purposes. These include combustibles, oxygen carriers, antacids, absorbent and antifreeze.

Table 18.2 Chemicals used in explosives and fireworks.

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Ingredient

Formula

Purpose

Ammonium Nitrate Chalk/Calcium Carbonate Charcoal or Lampblack Ethylene Glycol Dinitrate Kieselguhr Lead Azide Lead Styphnate Liquid Oxygen Mercury Fulminate Nitrocellulose (Guncotton) Nitroglycerin Paraffin Pentaerythritol Tetranitrate (PETN) Potassium Nitrate Sodium Nitrate Sulfur Tetranitro diglycerine Trinitrotoluene (TNT) Zinc Oxide

NH4 NO3 CaCO3 C C2H4(NO3)2 SiO2 Pb(N3)2 PbC6H(NO2)3O2 O2 Hg(ONC)2 (C6H7(NO3)3O2)n C3H5(NO3)3 CnH2n+2 C5H8 N4O12

Explosive Base, Oxygen Carrier Antacid Combustible Explosive Base and Antifreeze Absorbent Primary Explosive Primary Explosive Oxygen Carrier Primary Explosive Explosive Base, Gelatinizing Agent Explosive base Combustible Explosive Base for Caps and Detonation Fuse

KNO3 NaNO3 S C6H10 N4O13 C7H5 N3O6 ZnO

Oxygen Carrier Oxygen Carrier Combustible Explosive Base and Antifreeze Explosive Base Antacid

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A combustible is meant to burn and consume excess oxygen formed in the explosion so as to reduce the formation of poisonous gases such as nitrous oxide, nitrogen dioxide etc. due to their burning; they also lower the heat of explosion. Oxygen carriers are meant to ensure complete oxidation of the carbon in the explosive. In this manner they inhibit the formation of carbon monoxide. Sometimes their purpose is to lower the heat of explosion and create a less efficient blast. Antacids are added to an explosive compound to increase its long term storage life, and to reduce the acidic value of the explosive base, particularly nitroglycerin. Absorbents, like saw dust, nut shells etc., are mixed with dynamite and some other explosives so as to hold the explosive base from exudation, seepage and settlement to the bottom of the cartridge or container. Antifreeze is meant to lower the freezing point of the explosive. Table 18.2 lists some chemicals and compounds used in making explosives and fireworks.

18.3

PROPERTIES OF EXPLOSIVES

Explosives have many distinct properties. They are summarized in Table 18.3.

18.3.1 Velocity of detonation The velocity of detonation is the most important property of an explosive. It is a measure of the speed at which the detonation front moves. The velocity of detonation depends upon composition of the explosive, density achieved in charging the blasthole, diameter of the blasthole, degree of confinement, presence of voids in the rock mass, rock mass temperature, and temperature generated at the initiation element of the detonators that are used for firing the explosive. Different explosives have different velocities of detonation. Some military explosives have velocities of detonation reaching to 10300 m/s but explosives used in rock blasting have velocities of detonation ranging between 2000 to 7000 m/s. Velocities of detonation of some explosives are given in Table 18.4. If the explosive is filled in a blasthole by proper tamping or compressive force as recommended, the density of explosive in the blasthole increases. Similarly when cartridges of explosives are filled in the factory with higher compression and achieve higher density, the velocity of detonation increases with density of explosive. For example, the detonation velocity of ANFO at a density of 0.8 g/cc is about 3048 m/s and at a density of 1.2 g/cc it is about 4572 m/s. For the same explosive, detonated in the same manner, the velocity of detonation increases with increasing hole diameter. Variation of detonation velocity with diameter of holes for some common explosives is shown in Figure 18.3. It is very natural that detonation velocity depends upon the degree of confinement. If an explosive is detonated in a blasthole drilled in soft rock, it gives lower detonation velocity than that in a blasthole drilled in hard rock. This is due to the fact that the increase in the diameter of the blasthole is larger in the soft rock than in hard rock. If the rock mass has voids or joints in it, the gases formed in the explosion first fill the voids and joints. Therefore, the pressure remaining for expansion of the blasthole is less. This reduces the velocity of detonation.

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Table 18.3 Properties of explosives. Explosive property

Meaning

Velocity of Detonation Detonation Pressure

Velocity in m/s at which the shock front of the detonation layer travels within the column of explosive. Pressure developed by detonation of the explosive in the detonation zone. It is usually measured in GPa. Pressure exerted on the wall of the blasthole immediately after the detonation. Total amount of energy released by the explosion in MJ for each kg of explosive. This includes the energy released in the form of heat as well as the pressure exerted by the gases generated in detonation. Total amount of heat released by the detonation in kcal for each kg of explosive. Amount of gas generated by detonation of one kg of explosive under normal conditions. Possibility of causing detonation by such means as friction, pressure, heat etc. How easily the explosive can be handled and transported through different modes of transport. Brisance value indicates the shattering effect of the explosive. Weight of the explosive in kg, contained in each liter volume of the blasthole. Volume of poisonous gases generated in terms of liters per kg of explosive detonated. Whether the properties of explosive remain unchanged by mixing the explosive with water. Hygroscopicity is a measure of water-absorbing capacity of an explosive. What is the smallest diameter of blasthole in which the explosive can be charged and detonated to get the desired explosion effect. How long the explosive can be stored in originally packed and unpacked condition without a change in it’s properties. How much is the volatility of the explosive. This is the ability of the explosive to coexist with other materials.

Blasthole Pressure Strength

Heat of Explosion Specific Gas Volume Sensitivity Transport and Handling Safety Brisance Value Charging Density Toxic Fumes Water Resistance Hygroscopicity Minimum Hole Diameter Storage Life Volatility Material Coexistence

Table 18.4 Velocities of detonation of some explosives.

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Explosive

VOD (m/s)

Explosive

VOD (m/s)

Lead Azide Mercury Fulminate Picric Acid Trinitrotoluene (TNT) PETN RDX HMX

4630 4250 7350 6900 8400 8750 9100

Nitroglycerin Dynamite (65% Gelatine) Ammonium Picrate Black Powder Lead Styphnate Nitrocellulose Nitroglycol

7700 6500 7150 400 5200 4492 8250

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7500 Velocity of Detonation in m/s

Cast Pentolite (High Explosive) Straight Gelatin (60% High Explosive) 6000

Semigelatin (45% Bulk Strength Explosive) Slurry Blasting Agents

4500 ANFO 3000

1500

0

25

50

75

100 125 150 175 200 225 250 Blasthole Diameter

Figure 18.3 Increase in detonation velocity with blasthole diameter.

If the rock mass temperature is very low, the heat energy evolved in the detonation is absorbed by the surrounding rock to a much greater degree. This also reduces the velocity of detonation. To initiate the detonation of the main explosive filled in a blasthole, some small volume of very high explosive material is used in the blasting cap i.e. detonator. If the temperature generated in such initiation element is lower, then the velocity of detonation is also less. It must also be borne in mind that if the explosive is being used a long time after its manufacture, it loses some of its strength. Therefore, the velocity of detonation of such explosive is low.

18.3.2

Detonation pressure

Detonation pressure is the pressure developed immediately after the detonation of the explosive just behind the detonation front. It cannot easily be measured by any test but is usually estimated from the known properties of the explosive by using the following equation. Pd = (1/2) * ρe * Cd * 10–9 where Pd = Detonation pressure in GPa ρe = Density of explosive in kg/m3 Cd = Velocity of detonation in m/s Detonation pressure can also be calculated with reasonably good accuracy by analysis of chemical equations in the detonation process.

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18.3.3

Blasthole pressure

Pressure exerted on the wall of the blasthole by the expanding gases formed in the detonation of the explosive is called the blasthole pressure. It is obviously dependent upon the detonation pressure generated by the explosive and the confinement of the explosive in the blasthole. If the rock mass has cracks, joints, cavities etc., the blasthole pressure tends to be very low, but otherwise it is often of the order of 50% to 60% of the detonation pressure.

18.3.4

Strength

Strength of an explosive means the energy output by unit weight or unit volume of the explosive. Such strength is related to density and detonating velocity as well as heat and gas volume liberated in the detonation of the explosive. The total energy Et released by an explosion can be divided into two components, viz. shock energy Es and bubble energy Eb. The shock energy is actually caused by the shock wave. It travels away from the place of its generation, i.e. the detonation zone, in the form of a strain wave. The bubble energy is caused by the heat evolved by the chemical reactions involved in the detonation process. Table 18.5 gives shock wave and bubble energy of some explosives. Energy released in the detonation is either usefully consumed in fragmentation or wasted through heat, ground vibrations, light, sound waves and flyrocks. The efficiency of an explosive depends upon the explosive formulation, blasthole parameters, rock mass conditions, environmental factors and the manner of blasthole charging. Terms used in context of strength of explosive are: 1 2

Weight and bulk strength. Absolute and relative strength.

18.3.4.1

Weight and bulk strength

The energy released by one unit weight of explosive, expressed in terms of the concurrent energy units is called weight strength measured as kcal/kg. Energy released by one unit volume of explosive, expressed in terms of the concurrent energy units is called bulk strength, for example kcal/liter. Table 18.5 Bubble, shock and total energy of some explosives.

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Explosive name

Eb in Cal/g

Es in Cal/g

Et in Cal/g

Pentolite (50% TNT+ 50% PETN) TNT OCG Emulsion ANFO Heavy ANFO

490 505 495 460 500 470

234 215 408 370 340 360

724 720 903 830 840 830

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18.3.4.2

453

Absolute and relative strength

When the strength of an explosive is given in terms of kcal/kg or kcal/liter, it is called absolute strength because it is expressed in absolute units. It is difficult to measure strength of explosives in terms of absolute units. However, several tests enable the effect of the strength of an explosive to be assessed in an easily measurable way. Such tests give indicators of the strength of an explosive with respect to the strength of a common explosive, which is treated as 100. Strength expressed in terms of such an indicator is called relative strength.

18.3.5

Heat of explosion

The quantity of heat generated by an explosive after its detonation is a very important aspect of an explosion. It is treated in detail in chapter 19 on the thermochemistry of explosives.

18.3.6

Specific gas volume

As stated earlier, specific gas volume means the volume of gas evolved by explosion of one kg of explosive measured in liters. It obviously depends upon the chemical composition of the explosive. For an explosive to be very effective, it must generate an optimum volume of gas and an optimum quantum of heat upon detonation. This is because heat is required to expand the gases. Only after such expansion are the walls of the blasthole subjected to high pressures, which is one of the reasons for rock fragmentation. More on the specific volume is contained in chapter 19 on the thermochemistry of explosives.

18.3.7

Sensitivity

Sensitivity of an explosive is a measure of the ease with which it can be detonated. An explosive is called very sensitive if it can be detonated very easily by such factors as heat, pressure or relatively low flame temperature. Till they are consumed, explosives have to be handled very frequently. They have to be loaded in vehicles, godowns and blastholes. They require transportation from factory to godown to worksite etc. Naturally, an ideal explosive for use in rock blasting should be very insensitive so it does not detonate in all these storing or transportation processes. Explosives used in olden days, e.g. nitroglycerin, were very sensitive and exploded even by gentle scratching. Thousands of miners must have died while handling explosives. Today explosives used for rock blasting are far less sensitive and thus have become far safer. Sensitivity of an explosive can be reduced by adding certain components, like glass bubbles, to the main explosives material. Sensitivities of explosives can be in relation with different type of disturbances as under. 1 2

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Sensitivity to shocks Sensitivity to friction

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3 4 5


Sensitivity to heat Sensitivity to detonator strength Air gap sensitivity Details are given below.

18.3.7.1

Sensitivity to shocks

When an explosive is subjected to shock, the molecules in the explosive move vigorously. In the process they impinge upon each other and the bonds that hold the atoms in the molecule together are disturbed. If the bonds are loose they break and release energy. This release of energy causes many more molecules to release energy. In a very short time the material detonates if the bonds in it are loose. Materials which contain nitrogen along with carbon, hydrogen and oxygen are usually very prone to detonate. Sometimes the magnitude of the shock to which an explosive is subjected is too low to cause detonation but deflagration of the explosive occurs. Shock sensitivity indicates how easily an explosive can detonate when it is subjected to a shock. 18.3.7.2

Sensitivity to friction

Friction always generates heat. If an explosive material rubs against another hard material, heat is generated and is imparted to the molecules of the explosive. The bonds between the atoms in the explosive molecules can withstand only a certain quantum of heat. If the heat imparted to explosive molecules exceeds this quantum, the bonds which have been holding the atoms together break and release energy. The consequent chain reaction leads to detonation or deflagration of the explosive material. 18.3.7.3

Sensitivity to heat

If an explosive is subjected to slowly increasing temperature by applying controlled heat to it, at a particular temperature it suddenly decomposes and starts deflagrating. This temperature is called the Ignition Temperature. It is to be noticed that some explosives, like black powder, have an ignition temperature between 300 to 350°C but are detonated by even a tiny spark. As against this some industrial explosives have an ignition temperature between 180 to 230°C but need a detonator for causing detonation. 18.3.7.4

Sensitivity to detonator strength

This is more commonly known as cap sensitivity. It is essential that an explosive should detonate only by use of a blasting cap of a certain strength, and not less than that strength. Further, such detonation of the explosive with the particular blasting cap must occur with 100% certainty. Thus, explosives are classified by the strength of the blasting cap.

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18.3.7.5

455

Gap sensitivity

In many blastholes the explosives are in the form of cartridges. Similarly bulk explosives in polythene bags are also used when the blastholes have water seeping into them. Naturally there is a gap in the two explosive containers. In most situations it is desirable that once one cartridge or bag of explosive is detonated, the neighboring cartridge must also be detonated automatically by the shock wave generated by the first. This type of sensitivity naturally depends upon the gap between the explosives and hence it is termed the gap sensitivity. From the viewpoint of storage of explosives, the gap sensitivity is measured in unconfined conditions. From the viewpoint of blast in a blasthole, the gap sensitivity is measured in confined conditions. Gap sensitivity of an initiating explosive is the highest, often of the order of 8D or even more. For blasting agents it can be as low as 2D.

18.3.8

Handling and transport safety

While an explosive is being handled and transported it is subjected to shocks and elevated temperatures. There are no specific indicators for handling and transportation safety of an explosive, because it can be correlated to different types of sensitivities like shock, friction, heat, gap etc. Handling and transport safety is usually expressed in relative terms.

18.3.9

Brisance value

The term brisance originates from the French word briser which means to break or shatter. It is a measure of the rapidity with which an explosive builds the maximum pressure in its confined state. Usually the explosives that have high detonation velocities also have high brisance values because the shock wave generated by them is highly energetic. If two explosives having equal energy output and confined in identical enclosures are exploded, the one with higher brisance value will break the casing into smaller fragments. The brisance values of different explosives are estimated by the sand crush test. They are always relative to TNT. One of the most brisant of the explosives is cyclotrimethylene trinitramine, i.e. RDX. Explosives with high brisance value are very desirable in military explosives but in mine blasting this property is somewhat undesirable because such explosives create a large volume of very fine rock fragments and a lot of dust. This is one of the reasons why explosives like RDX, PETN, HMX etc. are not used for rock fragmentation in mines.

18.3.10

Charging density

Charging density means the density of the explosive as it exists in the blasthole. It is measured as mass per unit length. It is computed by the following equation.

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ρc = 7.854 * 10–4 * D2 * ρe where ρc = Charge density in kg/m ρe = Density of explosive in g/cc D = Diameter of blasthole in mm In many charging methods the density achieved in charging the blasthole may be quite different than ρe. Hence in such cases it is more practical to calculate the charging density as ρc = W/L where ρc = Charge density in kg/m W = Weight of explosive in kg L = Blasthole length in m over which the explosive is filled If the charging density of an explosive in a blasthole is very low, it is very sensitive to the detonating cord. In such circumstances the detonation of the explosive starts from the top of the blasthole before the primer cartridge detonates and the blast becomes very ineffective. When charging density is higher the molecules of the explosive are nearer to each other. Naturally the detonation of the explosive having higher charging density takes place in a shorter time, and the devastation effect of the explosion is greater. Charging density depends upon many factors such as the density of explosive itself, the manner in which the hole is filled with explosive, the diameter of the blasthole and the inclination of the blasthole. Several methods of charging a blasthole e.g. tamping, pellet loading, press loading, cast loading etc. are practiced. When the diameter of the blasthole is larger it is possible to achieve a higher loading density by using the same explosive and the same loading method. Depending upon the loading method used, a charge density of 80 to 99% of theoretical maximum density can be achieved. When the molecules of explosive are densely packed they do not have much room to move with respect to each other. With this, sensitivity of the explosive reduces. However, beyond a certain value of density the crystals of the explosive get crushed and the sensitivity increases. Some explosives can be pressed very heavily. Such pressing is called “dead-pressing”. Dead-pressed explosives do not detonate easily. Some do not detonate at all.

18.3.11 Toxic fumes If an explosive produces only water vapor, carbon dioxide and nitrogen after its detonation, it can be looked upon as an ideal explosive because all these products are harmless. However, the end products of most explosives contain the following ingredients in small quantities.

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1 2

457

Non-toxic but irritating gases like methane (CH4), ammonia (NH4), acetylene (C2H2). Poisonous gases, such as carbon monoxide (CO), nitrogen dioxide (NO2), nitrous oxide (N2O), nitrogen oxide (NO), sulfur dioxide (SO2), hydrogen sulfide (H2S).

Since all these products are in gaseous form they are called fumes. Exposure of a human to these gases is dangerous. How some of these gases are hazardous is explained in Table 18.6. Three limiting values of tolerable concentrations of these gases in atmospheric air have been recognized. The values are in ppm, meaning parts per million. 1

2 3

Threshold Limit Value (TLV) – the amount of exposure to a gas for an 8 hour day for 5 days a week without any harmful effects. This is also called the permissible exposure limit. Ceiling Limit (CLV) – the amount of gas a person cannot be exposed to at any time. Immediately dangerous to life or health (IDLH) – the maximum concentration of a gas at which one could escape without any irreversible health effects.

TLV, CLV and IDLH values for different gases likely to be evolved in a mine explosion are given in Table 18.7. Table 18.6 Hazardous effects of poisonous gases on humans. Poisonous gas

Hazardous effect

Carbon Monoxide

Since carbon monoxide has about 300 times higher affinity to hemoglobin than oxygen, when carbon monoxide enters the lungs, the hemoglobin molecules start supplying carbon monoxide to cells in the body rather than oxygen. Cells of the human brain need very high quantity of oxygen. They start dying quickly if they are not given oxygen. Therefore, carbon monoxide quickly results in brain death of a human. Nitrogen oxide, in very low concentrations, dilates blood vessels and increases blood supply in the human body. However, when it is inhaled it acts the same way as carbon monoxide. In addition it also generates more toxic free radicals. When nitrogen dioxide is inhaled it goes into the lungs and immediately forms Nitric acid by combining with water molecules in the blood plasma. This causes pulmonary edema and heavy damage to the cells. Final outcome is death. Hydrogen sulfide causes paralysis of the respiratory system.

Nitrogen Oxide Nitrogen Dioxide Hydrogen sulfide Sulfur Dioxide

Hazardous effects of sulfur dioxide are similar to those of nitrogen dioxide but it forms sulfuric acid rather than nitric acid. Table 18.7 Exposure limits of poisonous gases.

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Poisonous gas

Formula

TLV (ppm)

CLV (ppm)

IDLH (ppm)

Carbon Monoxide Nitrogen Oxide Nitrogen Dioxide Hydrogen Sulfide Sulfur Dioxide

CO NO NO2 H2S SO2

50 20 1 10 5

200 ? 3 15 10

1500 100 20 300 100

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Depending upon the quantity of undesirable toxic gases formed during detonation, the explosives are given fume class ratings. If an explosive produces a small quantity of gases, it has a better rating. For blasting in the open atmosphere, such as surface blasts, the fume classes are not as important as they are in a confined space. However, they should not be ignored in surface blasting operations. The US Bureau of Mines recognizes two fume classes viz. A and B as specified in Table 18.8. This classification is applicable for permissible explosives and is based on the measurements of CO, NO, NO2 and H2S. Another classification is stipulated by the Institute of Makers of Explosives. It has three classes viz. 1, 2 and 3 as shown in Table 18.9. This classification is applicable for non-permissible explosives. It is based only on measurements of CO and H2S. The factors which influence the volume of toxic fumes formed in an explosion are: Insufficient charge diameter Improper delay timings Deterioration of explosive by water Insufficient priming Use of plastic liners or paper wrappers in blasthole

18.3.12 Water resistance Water reduces the effectiveness of an explosive to a very great extent. This happens because one or more ingredients of the explosive dissolves in water and becomes ineffective. Another way in which water can affect the process of explosion in a blasthole is the formation of hot spots. If the pressure exerted on the explosive by a water column in the blasthole is high, the size of air bubbles reduces by compression. Such Table 18.8 Fume classes for permissible explosives (USBM). Noxious gases produced Fume class

ft3/lb

l/kg

A B

<1.25 1.25–2.5

<78 78–156

Table 18.9 Fume classes for non permissible explosives (IMA). Noxious gases produced

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Fume class

ft3/lb

l/kg

1 2 3

<0.36 0.36–0.75 0.75–1.52

<22.5 22.5–46.8 46.8–94.9

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compressed air is not easily usable in the process of explosion. Therefore, some portion of the explosive can become desensitized. This portion is called the hot spot. In very low temperature regions, the water may cool a water-resistant explosive to such a low temperature that a much higher detonation energy is required to ensure its detonation. Resistance of an explosive to water is measured in terms of the number of hours over which the explosive can remain submerged below water and still retain its ability to reliably detonate. On the basis of results of tests carried out on samples, the manufacturers express water resistance of the explosive as Excellent, Very Good, Good, Limited or Poor.

18.3.13

Hygroscopicity

Hygroscopicity is the tendency of a substance to absorb moisture. Water is the strongest enemy of an explosive. It affects explosives in many ways as under. 1 2

3

4

Water forms a film around an explosive molecule and inhibits the effective transfer of heat, shock etc. Thus, velocity of detonation is reduced greatly. The heat of explosion is quickly absorbed by water. It vaporizes but in the process the heat is lost and the effectiveness of the explosive in fragmentation reduces. In other words energy output is reduced. Water can cause some undesirable chemical reactions with one or more components of the explosive, and in the process the explosive detonates very weakly or sometimes it does not detonate at all. This way the sensitivity of the explosive is reduced. Some chemical reactions of water with an ingredient of the explosive give rise to some highly corrosive chemicals which severely affects the containers of explosive. This finally results in reduction of shelf life.

For all these reasons explosives must be made with only such ingredients which have very low hygroscopicity.

18.3.14

Storage life

Storage life, also called shelf life, of an explosive is measured in terms of the number of years over which it can be stored and used later without any loss or degradation of its properties. Storage life depends upon the characteristics of the explosive as well as the storage place. Explosives which contain radicals such as nitrites (−NO2), nitrates (−NO3) and azides (−N3), are intrinsically under a state of stress. They are easily affected by external factors such as presence of heat, moisture, radiation and electromagnetic fields etc. It is very difficult to ascertain the shelf life of any explosive by taking into consideration all these factors because they change with every storage place. For most explosives products, a shelf life of one year can be taken for granted so long as their original packing is not broken. In most cases satisfactory performance can be expected from most products two, three, and even four years later. This is

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because the explosive manufacturers are over-cautious and usually do not stipulate a longer shelf life.

18.3.15 Volatility Volatility means the ease with which the substance vaporizes. This obviously depends upon the temperature. Higher volatility is an undesirable property of an explosive because if one or more components of the explosive reduces in its proportion after vaporization, the explosive may become dead or dangerous. All mining explosives have insignificant volatility at the atmospheric temperatures at which they are stored or used. Volatility has very significant importance in the case of military explosives, which have to be stored over a long period of time. The maximum allowed volatility for military explosives is 2 mL volume of gas per kg weight of the explosive in 48 hrs.

18.3.16

Material coexistence

An explosive may chemically affect certain materials or may be affected itself by certain materials if they are brought in contact with each other. Under such circumstances the explosive may become dead or dangerous. Thus, the two materials have very poor coexistence. Since explosives are supposed to be stored in the containers in which they are supplied, the data about their coexistence with the packing materials is not published but the manufacturer always verifies the safe coexistence between the explosive and packing material. Coexistence gains more importance in the case of bulk explosives because when they are charged, they are in direct contact with the minerals in the rock mass. There have been instances when ANFO loaded in the blastholes drilled in copper mines exploded prematurely. Later investigations revealed that ammonium nitrate, being a very strong oxidizer, chemically reacted with pyrite and oxidized it. In the process substantial heat was generated. This heat caused the rise in temperature to a level that detonation occurred.

18.3.17

Minimum hole diameter

When an explosive is charged in a blasthole and the detonation is initiated, the detonation layer moves away from the point of initiation. In a mine blast it is very essential that the movement of the detonation layer is neither discontinued nor diminished in any manner. For this objective to be achieved, the diameter of a blasthole cannot be less than a certain value called the critical diameter. The critical diameter depends upon the chemical composition of the explosive, and to a much lesser extent the temperature as well as the thermal conductivity of the rock mass. In general it can be said that the critical diameter is inversely proportional to the velocity of detonation. The mathematical relation between the two, however, includes constants that have to be determined by tests.

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Table 18.10 Critical blasthole diameters. Explosive

Critical diameter in (mm)

Cast Pentolite (High Explosive) Straight Gelatin (60% High Explosive) Semigelatin (45% Bulk Strength Explosive) Slurry Blasting Agents ANFO

About 10 About 28 About 42 About 75 About 100

Critical diameters of the blastholes for certain explosives as can be found from measurements of the velocity of detonation and their plots as in Figure 18.3. In the plot it can be seen that for an explosive the velocity of detonation remains unchanged beyond a certain blasthole diameter. In other words this is the maximum velocity of detonation. Since there is no specific definition of critical diameter, it can be said the critical blasthole diameter of an explosive is the one where the velocity of detonation in the blasthole exceeds 90% of its maximum velocity of detonation. Critical diameters of certain explosives arrived at on the above basis are given in Table 18.10.

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Chapter 19

Thermochemistry of explosives

19.1

INTRODUCTION

Thermochemistry is a subject in which the changes in internal heat of the reacting products is studied. Every explosion is a series of several reactions that take place in a particular sequence at extremely rapid speed. These reactions are invariably exothermic, i.e. heat is evolved in the reaction. In an explosion the ingredients of the chemicals, or mixtures of them, first decompose into elemental form and again recombine by consuming some of the energy evolved. Energy changes in explosive reactions are calculated either from known chemical laws or by analysis of the products. This chapter deals with all such aspects of the chemical reaction of an explosive.

19.2

CHEMICAL NATURE OF EXPLOSIVES

As mentioned earlier, explosives are either pure chemicals with one type of molecules, or mixtures of many chemicals with different molecules. Exceptions do exist but as said earlier almost all explosive chemicals comprise of carbon, hydrogen, nitrogen and oxygen. Nitrogen is probably the most essential constituent of a chemical to attain the explosive nature. This is because nitrogen by itself is an inert gas. This means that it has least ability of sharing electrons with other atoms to form a molecule. In other words when a molecule is comprised of nitrogen, excessively high energy must be used to keep the bond intact. Such bonds are broken easily and are not reformed easily. Thus, a very high quantum of energy is released at extremely rapid speed when a chemical containing nitrogen decomposes. It has also been noticed that generally for the organic explosive elements, the strength increases with increasing molar weight.

19.3

REACTIONS OF EXPLOSIVE CHEMICALS

Thermochemical studies of reactions have much greater importance in the case of explosive chemicals than other chemicals because of the extreme speed of the reactions.

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This is because the end products of a chemical reaction of an explosive are extremely short-lived. Both pressure and heat of explosion fall within a few miniseconds. It is almost impossible to correctly measure their magnitudes experimentally. Therefore indirect or theoretical methods are required to be used to determine the maximum temperature and pressure values. Whenever a chemical reaction takes place, it progresses in a particular sequence as shown in Table 19.1 In the very initial stage all the molecules contained in an explosive chemical or mixture thereof, decompose to form atoms. Once this happens the atoms start combining with each other in the priority steps shown in the table. When a chemical is formed in this manner it does not undergo decomposition again. This means that the sequence proceeds only in increasing priority number in the table. If both the atoms required by a particular priority are not present the step is skipped and the reaction proceeds as per the next priority. Let us take an example of picric acid. In the initial step molecules of the elements are formed. Hence, C6H2(OH)(NO2)3 ⇒ 6C + 3H + 3N + 7O As the mixture does not contain any metal or chlorine there are no reactions as per priority step 1, 2 and 3. As per step 4 the mixture gets converted into 6CO + 3H + 3N + O As per step 5 the mixture gets converted into 6CO + H2O + H + 3N

Table 19.1 Priorities of formation of end products of explosive reactions.

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Priority

Composition of the product mix

End products

End product phase

1 2 3 4 5 6 7 8 9 10

A metal and chlorine Hydrogen and Chlorine A Metal and Oxygen Carbon and Oxygen Hydrogen and Oxygen Carbon Monoxide and Oxygen Nitrogen Excess Oxygen Excess Hydrogen Carbon

Metallic chloride HCl Metallic oxide CO H2O CO2 N2 O2 H2 C

Solid Gas Solid Gas Gas Gas Gas Gas Gas Gas

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As there is no unreacted oxygen left in the mixture there is no reaction as per priority step 6 and the mixture remains 6CO + H2O + H + 3N As per step 7 the mixture gets converted into 6CO + H2O + H + 1.5N2 As there is no excess oxygen left in the mixture there is no reaction as per priority step 8 and the mixture remains 6CO + H2O + H + 1.5N2 As per step 9 the mixture gets converted into 6CO + H2O + 0.5H2 + 1.5N2 Since there is no unreacted carbon left, the mixture remains 6CO + H2O + 0.5H2 + 1.5N2 In the above equations the symbols such as 0.5H2 or 1.5N2 are correct because they do not represent molecules but represent moles each of which contains 6.02257 * 1027 molecules. It must be specifically noted that the water formed after the explosive is in vapor i.e. gaseous form. End products formed after the detonation of different explosives are given in Table 19.2. The important aspects of the chemical reactions of explosives are: 1 2 3 4

Oxygen balance Volume of products of explosion Heat of explosion Strength of explosive Each of these aspects is detailed below.

19.3.1

Oxygen balance

An explosive does not use oxygen available from the atmosphere in its chemical reaction. If it is required to do so the reaction will be very slow and the material will not qualify to be called an explosive. If the atoms contained in an explosive molecule are in such numbers that the oxygen molecules evolved can completely convert all metal atoms to metal oxide

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Table 19.2 Characteristics of some chemical reactions of explosives and the end products.

Explosive name

Chemical formula

End products

Gunpowder Nitrocellulose

2KNO3 + 3C + S C24H29O9(NO3)11

Nitroglycerin Ammonium Nitrate Trinitrotoluene Picric Acid Ammonium Picrate Tetryl Mercury Fulminate Lead Azide

C3H5(NO3)3 NH4NO3 C7H5(NO2)3 C6H2(OH)(NO2)3 C6H2(NO2)3ONH4 C7H5N5O8 Hg(ONC)2 PbN6

N2 + 3CO2 + K2S 20.5CO + 3.5CO2 + 14.5H2O + 5.5N2 3CO2 + 2.5H2O + 1.5N2 + 0.25O2 2H2O + N2 + 0.5O2 6CO + C + 2.5H2 + 1.5N2 6CO + H2O + 0.5H2 + 1.5N2 6CO + H2O + 2H2 + 2N2 7CO + H2O + 2H2 + 2N2 Hg + 2CO + N2 Pb + 3N2

Qv in J/kg

Te in °C

f kg/cm2

V in m/s

Trauzl expansion in cc/10 g

Energy potential in kg · m

2098 5234

2090 2800

291 981

– 6100

30 420

2.1 5.3

6389 1608 2714 3546 2604 3802 1759 2864

3360 1100 2200 2717 1979 2781 4105 3180

964 500 822 977 837 1062 511 791

8500 4100 6800 7000 6500 7229 3920 5000

590 300 260 300 230 320 213 250

6.5 1.6 2.8 3.6 2.6 3.9 1.8 2.9

Qv = Heat of explosion at constant volume, Te = Temperature of explosion, f = Pressure exerted by 1 kg explosive in volume of explosive at explosion temperature, V = Velocity of detonation in m/s. From Science of Explosives by Myers.

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molecules, convert all carbon atoms to carbon dioxide molecules, convert all hydrogen atoms to water molecules and leave no oxygen atoms to form oxygen molecules, it is considered to have zero oxygen balance. It has been noticed that the strength, brisance and sensitivity generally tend to be maximum when the explosive molecule has perfect oxygen balance. The explosive molecule is said to have a positive oxygen balance if it holds more oxygen than is needed and a negative oxygen balance if it holds less oxygen than is needed for such conversion. The oxygen balance can be calculated by using the following formula OB = −(100%) * (MWo/MWe) * [2C + H/2 + M − O] where MWo = Molecular weight of oxygen MWe = Molecular weight of explosive C = Number of carbon atoms in the explosive H = Number of hydrogen atoms in the explosive M = Number of metal atoms in the explosive O = Number of oxygen atoms in the explosive As an example, consider picric acid, the end products of which are 6C + 3H + 3N + 7O It will need 12O to convert 6C into 6CO2 and 1.5O to convert 3H to 1.5H2O i.e. a total of 13.5O. Since it already has 7O the additional need is for 6.5O. By using the formula the oxygen balance of picric acid can be found as OB = −100 * (15.999/229.1037) * (2 * 6 + 3/2 − 7) = 45.39%

19.3.2 Volume of products of explosion A mole is a scientific term given to a numerical quantity of molecules. The number of molecules contained in a mole is 6.02257 * 1027 and is often termed as Avogadro’s Number. It can be shown that at NTP conditions i.e. temperature of 0°C and pressure of 101.325 kPa the volume of one mole of any gas equals to 22.3933 m3. Let us now consider the chemical decomposition of nitroglycerin. The chemical equation is as follows. C3H5(NO3)3 ⇒ 3CO2 + 2.5H2O + 1.5N2 + 0.25O2 In a way this is an ideal explosive because it produces totally non-toxic end products and it has a positive balance of oxygen. As indicated by the above equation one mole of nitroglycerin molecules produces a total of 7.25 g-moles of gas at 0°C. Hence, its volume at 0°C will be 7.25 × 22.3933 = 162.3514 L.

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By using Charles’ Law the volume for 20°C works out to 162.3514 * ((273.15 + 20)/273.15) = 174.2387 L.

19.3.3

Heat of explosion

Some energy is required for formation of each molecule. When a molecule of explosive breaks down, the energy required for its formation is released. Immediately a part of the energy is spent in formation of the end products of the reaction. Thus, the heat evolved in an explosion can be calculated as ΔE = ΔEfe − ΔEfp where ΔE = Energy evolved in kJ/mol ΔEfe = Energy required in kJ/mol for formation of the explosive ΔEfp = Energy required in kJ/mol for formation of the end products of the explosive reaction For actual calculations, the heat of formation of some explosives and the end products of the reaction are given in Table 19.3. As an example, let us calculate the heat of explosion of Trinitrotoluene i.e. TNT. From Table 19.3 we have ΔHfe = −54.39 kJ/mol From Table 19.1 the end products of explosion of TNT are 6CO + C + 2.5H2 + 1.5N2 Table 19.3 Heat of formation of some explosive and the end products of explosive reactions.

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Name of end product or explosive

Chemical formula

Molecular weight

Formation energy in kJ/mol

Oxygen molecule Hydrogen molecule Nitrogen molecule Carbon monoxide Carbon dioxide Water Nitroglycerin RDX PETN HMX Trinitrotoluene Tetryl

O2 H2 N2 CO CO2 H2O C3H5N3O9 C3H6N6O6 C5H8N4O12 C4H8N8O8 C7H5N3O6 C7H5N5O8

31.998 2.0158 28.014 28.01 44.009 18.0148 297.1313 222.1164 316.1342 296.1552 227.1315 287.1435

0 0 0 −111.8 −393.5 −240.6 −333.66 +83.82 −514.63 +104.77 −54.39 +38.91

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Therefore, ΔEfp = 6(−111.8) + 1(0) + 2.5(0) + 1.5(0) = −670.8 With this we get ΔE = −54.39 − (−670.8) = 616.41 kJ/mol Since ΔE is positive the reaction is exothermic and heat is evolved and not absorbed. Since one mole of TNT weighs 227.1435 grams the heat evolved by the explosion of one kg of TNT will work out to 616.41/0.2271435 = 2713.74 kJ.

19.3.4

Strength of explosive

As soon as an explosive is detonated, gases and other end products, and also a huge quantity of heat, evolves from the chemical reaction. The gases and heat do lot of work on the surroundings such as expansion of the blasthole, crushing of the rock in the peripheral wall of the blasthole, fragmenting of the rock mass, throwing of rock pieces etc. The capacity to do all this work is called the strength of the explosive. From the basic principle, the strength of the explosive will be equal to the volume of gases evolved and the change in internal energy. In this regard the concept put forth by Berthelot is of great importance. Even Berthelot also knew that it is not a perfect concept hence it is called the Berthelot Approximation. It indirectly states that the strength of the explosive is proportional to the product of change in internal energy in the explosion and the volume of gases evolved in the explosion for each mole of the explosive. From this an equation for the strength of any explosive can be written as St = 840 * Δn * ΔE/(Mw)2 where St = Strength of Explosive as a percentage in relation to the strength of trinitrotoluene. Δn = Volume of gases produced per mole by the reaction of the explosive. ΔE = Change in internal energy due to chemical reaction. Mw = Molecular weight of the explosive in g/mol. The strength is relative to the strength of trinitrotoluene because the constant 840 accounts for the units and values of Δn and ΔE for trinitrotoluene. Let us consider the example of RDX having chemical formula C3H6N6O6. The steps of the chemical reaction of RDX explosion are: Step 1: C3H6N6O6 ⇒ 3C + 6H + 6N + 6O Step 4: 3C + 6H + 6N + 6O ⇒ 3CO + 6H + 6N + 3O Step 5: 3CO + 6H + 6N + 3O ⇒ 3CO + 3H2O + 6N Step 7: 3CO + 3H2O + 6N ⇒ 3CO + 3H2O + 3N2

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Thus, the end products of the explosion reaction of RDX are 3CO + 3H2O + 3N2. All these products are in gaseous form hence 1 mole of RDX evolves 9 moles of gas. In other words, ⇒Δn = 9. For RDX we have ΔEfe = 83.82 and for the end products DEfp = 3(−111.8) +3 (−240.6) = −1057.2. Hence ΔE = 83.82 − (−1057.2) = 1141.02. This gives St = 840 * 9 * 1141.02/(222.1164)2 = 174.845% Thus, the strength of RDX is about 175% that of TNT.

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Chapter 20

Explosives – their history and composition

20.1

INTRODUCTION

After having dealt with the properties of explosives, desirable and undesirable from the viewpoint of their safe and effective use in mines, it is a place to take detailed account of the actual explosives, their ingredients etc. Even before taking such account, it is very important to look into the historical developments in the realm of explosives from the viewpoint of their use in the mining and civil engineering domains. This chapter deals with both the above topics.

20.2

BRIEF HISTORY OF EXPLOSIVES

The following paragraphs give a glimpse of the history of explosives. Some chemicals, or mixtures of them, that can be considered as explosives by current definitions of the word, may have been known to mankind long before the Christian era, but there are no specific written references to them. Armaments with explosive-type devastating power are said to have been used in the Mahabharata War supposed to have been fought as early as 5000 BC. Black powder was known to the Chinese from the eighth century but was used only for entertaining the public. The first written reference to an explosive substance is by Arabian writer named Abd Allah. In the year 1200, he mentioned that saltpeter is the main ingredient of black powder. In the year 1242, English friar Roger Bacon published the formula for making gunpowder through his book ‘De Mirabili Potestate Artis et Nature’. It was in coded language. He also gave details of the destructive power of the powder. In the year 1380, a German Franciscan monk, by the name of Berthold Schwarts, developed gun powder i.e. black powder. He also used it in guns. In the year 1617, the use of black powder for the purpose of rock blasting was first proposed by Martin Weigel, a mining superintendent at Frieberg, Germany. Casper Windt is said to have actually used black powder for rock blasting at a German town Schemitz in the year 1623. In the year 1670, German miners revealed the secret of using gun powder for the purpose of rock blasting in Cornwall England.

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In the year 1675, a factory for making black powder was established in what is now Milton, Massachusetts, USA. The first instance of the use of black powder for blasting rock in a road-widening project occurred in Switzerland in the year 1696. In the year 1745, Doctor Watson invented the technique of exploding black powder by electric spark. In the year 1750, Benjamin Franklin invented a cartridge by compressing black powder in a cylindrical shell with a electric spark generating device in it to cause detonation. In the year 1773, black powder was used for mine blasting for the first time in a copper mine in Connecticut in what is now the USA. In 1822 the first hot wire detonator was produced by Dr. Robert Hare. Using one strand separated out of a multi strand wire as the hot bridge wire, this blasting cap ignited a pyrotechnic mixture (likely to be potassium chlorate/arsenic/sulfur) and then a charge of tamped black powder. Moses Shaw of New York patented a device for electric firing of black powder in the year 1830. William Bickford of Cornwall, England invented a safety fuse in the year 1831. In the year 1832, Dr. Robert Hare developed the electric blasting cap. In the year 1846, an Italian chemist, Ascanio Sobrero, discovered nitroglycerin. While doing so he was nearly killed by the explosion. He therefore never revealed the formulation. Wilbrand invented TNT i.e. Trinitrotoluene in the year 1863. In the same year Alfred Noble developed a safer variety of nitroglycerin for use in rock blasting. It could not be exploded by a normal fuse. In the year 1864, Alfred Noble developed the first detonating cap. This became essential for detonating the safer variety of nitroglycerin. Alfred Noble invented dynamite by mixing kieselguhr with nitroglycerin in the year 1866. In the year 1870, a plant was set up by the Giant Powder Company for making dynamite at San Francisco, California, USA. In the year 1875, Alfred Noble patented blasting gelatine. In the same year Perry, Gardiner and Smith independently developed and marketed caps which combined the hot wire detonator with mercury fulminate explosive. These were the first generally modern type blasting caps. An electric blasting machine containing a rack bar was developed by H. Julius Smith in the year 1878. In the year 1884, a large scale use of ammonium nitrate, commonly abbreviated as AN, as an additive to dynamite began. In the year 1895, two-component explosives were used for the first time in New York Harbor. In the year 1902, detonating cord was introduced in Europe. In the year 1913, explosive was used by M. Kinley for extinguishing a wildly burning oilwell in California. In the year 1914, construction of the Panama Canal began. Even today, i.e. after a century, it is considered to be the largest ever engineering project. In the year 1917 L. Mintrop, a German scientist, invented the seismograph.

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A devastating explosion of ammonium nitrate took place in Oppau, Germany in the year 1921. This explosion brought the explosive nature of ammonium nitrate to the notice of scientists. In the mid 1920s liquid oxygen-based explosives made their appearance in the USA. In the year 1935 a machine for measuring blast vibrations was developed by the US Geological Survey. In the mid 1950s high speed photography was introduced in mining for analyzing a blast. In the year 1955 Bob Akre developed “Akremite”. In the year 1956 a mixture of ammonium nitrate and fuel oil was used for the first time in US Steel Corp.’s Oliver Mining Division for rock blasting. This two-component explosive later got the acronym ANFO. In the same year Mel Cook introduced slurry explosives in mining practice. A specialized vehicle which could mix the two components of ANFO in precise proportions inside the blasthole was developed in Canada in the year 1957. By the end of the 1950s prilled AN and fuel mixture became so popular that it shunted out dynamite. In the year 1967 shock tube type non-electric detonators were introduced in Sweden. Emulsion explosives first appeared in the year 1969. Computer modeling of blast design was developed by Land & Favreau in the year 1972. In the early 1970s electronic recording seismographs were introduced by Dallas Instruments. In the year 1973, electronic sequential blasting was introduced at the first Kentucky Blaster’s Conference. Shock tube type non-electronic delay detonators were introduced in the year 1974. The International Society of Explosives Engineers was formed in 1974 in Pittsburgh, Pennsylvania, USA. In the year 1977 glass “bubbles” were introduced in mining blast practice. Digital sampling seismographs were developed in the 1980s. In this decade the earliest computer programs for improving blast timing patterns and blast plans were introduced in mining practice. The technique of laser profiling for blast design and analysis, that was developed in Britain, was introduced in USA in the year 1988. In the late 1980s electronic delay detonators were introduced. Nearly 700 oil well fires in Kuwait after the Gulf War 1n 1992, were put out by means of explosives.

20.3 TYPES OF EXPLOSIVES IN A BLASTHOLE A blasthole may contain two to four of the following entities of explosives. 1 2

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Detonator i.e. Blasting Cap Primary Explosive

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3 4


Booster Explosive Main Explosive

A blasthole charged with them is shown in Figure 20.1. Their primary functions are as under.

20.3.1

Detonator (Blasting cap)

A detonator is the device that starts the process of detonation in a blasthole. It is made from a very sensitive explosive such as lead azide or mercury fulminate. Initial energy delivery to the detonator is by means of a spark/heat generated through two electric wires or fire propagated up to it through a safety fuse. When a detonator is detonated by means of a safety fuse it is called a blasting cap. Modern explosives used in mines are very safe. In other words they do not detonate unless a very high quantum of energy is imparted to them. This is the reason for the need of a detonator. Detonation in a blasthole is always started from the bottom of the blasthole. Hence, the detonator is always placed in the bottom of the blasthole after connecting it to a safety fuse or electric wires. A detonator always consists of explosives that are sensitive to heat.

Stemming

Detonator and Primer Explosive Being Lowered in a Blasthole

Pair of Electric Wires or Safety Fuse Main Explosive Booster Explosive Primer Explosive Detonator or Blasting Cap

Booster Explosive Being Lowered in a Blasthole

Figure 20.1 A charged blasthole.

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20.3.2

475

Primary explosive

Many explosives used in large surface mines are very insensitive. As they do not get effectively detonated even by using blasting cap no. 8, they are called cap insensitive. They need more powerful explosive to start the detonation of the main explosive in the blasthole. Primary explosives are made for this purpose. Primary explosives come in the form of a cylindrical-shaped cartridge and contain a detonator embedded within them. The explosives used for making primer are PETN, TNT, RDX, Tetryl, etc. The properties of these are presented in Table 20.1.

20.3.3

Booster

A booster is also made by using PETN, TNT, RDX, Tetryl, Picrite etc. It is used anywhere in the blasthole where there is a need for generating higher energy. Since both these are cap sensitive, they easily detonate by the detonation wave traveling in the main explosive. Therefore, they do not need any embedded detonator. The most common place for a booster is at the bottom of the blasthole where the burden is usually more than in the rest of the blasthole. At this place much higher energy is required to be generated so as to ensure that no stumps are formed and the rock mass in the wider burden gets well fragmented.

20.3.4

Main explosive

The main explosives used in mining blasts are usually multi-component type. They contain fuel, oxidizer and in many cases a sensitizer as well. Fuel is meant to burn and generate heat. The oxidizer ensures proper burning of the fuel without formation of poisonous gases. It also accelerates the process of burning. Where high energy output is needed a sensitizer is added to the explosive mixture. The fuel, oxidizer and sensitizer combination of Dynamite, ANFO, Slurry and Emulsions is given in Table 20.2. The main explosives used for mine blasts are made from many components. Some of these are 1 2 3 4 5 6 7 8

Nitroglycerin Nitrocellulose Ammonium Nitrate Sodium Nitrate Fuel Oil Wood Pulp Sulfur Antacid

Of these nitroglycerin and nitrocellulose are explosives, ammonium nitrate, fuel oil, wood pulp and sulfur are combustible, sodium nitrate is an oxygen carrier and antacid as the name indicates is an anti-acidity compound. The properties of these components are as follows.

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Table 20.1 Properties of some primary explosives.

Explosive

Chemical formula

Full or other name

Appearance

C5H8N4O12 Pentaerythritol White Crystalline tetranitrate Solid TNT C7H5N3O6 Trinitrotoluene Pale Yellow Solid Block After Casting White Crystalline RDX C3H6N6O6 Research Development Solid Explosive Yellow Crystalline Tetryl C7H5N5O8 Nitramine or Tetralite Solid Picric Acid C6H3N3O7 Phenol Colorless to trinitrate Yellow Solid White Powder Lead Azide Pb(N3)2 Grey Crystalline Mercury Hg(CNO)2 Solid Fulminate

PETN

Molar Weight Density Melting in g/mol in g/cc Point °C

Sensitivity (shock, heat)

Pure water resistance

Velocity of Detonation detonation in pressure in m/s GPA

316.1342

1.77

141.3

Excellent

8400

33.5

227.1315

1.654

80.35

Medium, Medium Insensitive, Insensitive

Excellent

6900

22.7

222.1164

1.82

205.5

8750

33.9

287.1435

1.73

129.5

229.1037

1.763

290.24 284.624

4.71 4.43

Low, Low

Sensitive, Very Good 7570 Sensitive 122.5 Sensitive, Very Good 7350 Sensitive Explodes 350 Very Sensitive Low 5180 Explodes 180 Very Sensitive 4250

Unknown 25.1

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477

Table 20.2 Composition of main explosives used in mine blasts. Explosive type Oxidizer

Fuel

Dynamite

Solid Carbonaceous Oil

ANFO Slurry Emulsion

20.3.4.1

Solid Nitrate Salts

Sensitizer

Liquid Nitroglycerin, Voids, Bubbles Solid Ammonium Nitrate Liquid Diesel Oil Voids, Friction Solid or Liquid Salt Solutions Solid or Liquid Carbonaceous Solid or Liquid TNT, of Nitrates Fuel, Aluminum Fine Aluminum, Bubbles Liquid Nitrate Salt Liquid Carbonaceous Oils Bubbles

Nitroglycerin

Nitroglycerin is a clear yellowish/colorless liquid. It has chemical formula C3H5(NO3)3, molar mass of 227.0845 g/mol, density 1.6 g/cc at 15°C and melting point of 13.2°C. Maximum detonation pressure generated by nitroglycerin is about 26.9 GPa. Nitroglycerin is highly sensitive to shocks and heat. It detonates with even the slightest shock. It decomposes through detonation at low temperatures of 50 to 60°C, which means it is also highly sensitive to heat. Due to extreme sensitivity nitroglycerin is never used in pure form for mining applications. By mixing it with other components its sensitivity reduces considerably and then it can be used with care. Such products are called dynamite and many other trade names. 20.3.4.2

Nitrocellulose

Nitrocellulose is a yellowish white cotton-like filament. It has chemical formula (C6H7 (NO3)3O2)n, molar mass of (297.1313)n, density of about 0.69 g/cc at 15°C and melting point between 160 to 170°C. Nitrocellulose is an explosive. It is rarely used without addition of other components to the explosive mixture. In its non-compressed form, called smokeless powder, at a density of 0.69 g/cc, it has a velocity of detonation of 4492 m/s. If it is compressed to attain a density of 1.2 g/cc it attains a velocity of detonation of 7300 m/s. In this form it is called guncotton. Maximum detonation pressure generated by nitrocellulose is guessed to be between 16 to 27 GPa. Nitrocellulose is a dangerous material to store. It is hygroscopic and has a very high detonation temperature. 20.3.4.3

Ammonium nitrate

Ammonium nitrate is a white solid. It has a chemical formula (NH4)(NO3) with molar mass of 80.043 g/mol, density 1.725 g/cc at 20°C, and melting point of 169.6°C. By itself ammonium nitrate is not considered as an explosive because it does not detonate even with blasting cap no. 8. It has very low shock sensitivity as well as very low heat sensitivity. When engineered to detonate with a combustible, it attains detonation velocities up to 5250 m/s.

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Ammonium nitrate is a most widely used component because it is an excellent oxidizer. It generates 60% oxygen by weight. Due to this property and its low price it is most commonly used in making explosives with a suitable fuel. Ammonium nitrate is also used in many non-ANFO explosives. Ammonium nitrate is highly soluble in water. In 100 ml water its solubility at 0, 20, 40, 60, 80 and 100°C is 118, 150, 297, 410, 576 and 1024 g respectively. Apart from being an abundantly used component in mining explosives, ammonium nitrate is also used as a fertilizer. Both fertilizer and explosive grade ammonium nitrate is in prill form as shown in Figure 20.2. However, there is a big difference in the porosity of these two types. The fertilizer grade ammonium nitrate is meant to slowly dissolve in ground water and moisture. Therefore, it has low porosity on the microscopic scale and a density of about 1.40 to 1.45 g/cc. Prills meant for explosive use have much higher porosity and lower density of about 0.7 to 0.85 g/cc. Such high porosity of explosive grade AN prills is needed because they can absorb the fuel oil when mixed and form uniform ANFO. 20.3.4.4

Sodium Nitrate

Sodium nitrate is a white powder or colorless crystals. It has chemical formula NaNO3 with molar mass of 84.994 g/mol, density 2.257 g/cc at 20°C, and melting point of 308°C. Sodium nitrate does not detonate. It is also highly soluble in water. As the price of sodium nitrate is low it is used as an oxidizer in many mining explosives. It generates nearly 56% oxygen by weight.

2 mm

Figure 20.2 Explosive grade prills of ammonium nitrate.

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479

Fuel oil

During the distillation of crude mineral oils, apart from petroleum many other products are formed. Most of them are in the form of hydrocarbon carbon chains and have a chemical formula (CH3)2(CH2)n, where n indicates the number of links in the chain. There are six classes of fuel oil, numbered 1 to 6 according to their boiling points. The boiling points of the fuels in these six classes range between 175 to 600°C and the length of a chain ranges between 9 to 70. The most common fuel oil used in explosive ANFO is diesel oil because it has a relatively high flash point. 20.3.4.6

Wood pulp

Wood pulp is made by chemically or mechanically separating fibers from wood. Usually softwood trees like spruce, pine, fir, larch or some hardwoods such as lime, birch or eucalyptus are chosen for the purpose. Wood pulp has a chain structure with a common chemical formula (C6H10O5)n, where n is the number of links in the chain. Because of the need of forest conservation the use of wood pulp has decreased drastically over the years. 20.3.4.7

Sulfur

Sulfur is a lemon yellow crystalline solid. It is one amongst 92 naturally occurring elements. It burns with a blue flame that emits sulfur dioxide, carbon disulfide, and hydrogen sulfide. It has a molar weight of 32.066 g/mol, density between 1.92 to 2.07 g/cc depending upon its form and melting point of 115.21°C. For more than a thousand years it has been used as an active combustible component of gunpowder. Sulfur is insoluble in water. 20.3.4.8

Antacid

Acidity of an explosive mixture is an undesirable property because it considerably reduces the shelf life of the product by corroding the container. Zinc oxide is the most commonly used antacid component in an explosive mixture. 20.4

EXPLOSIVE MIXES USED IN MINE BLASTS

As stated in chapter 19, dry blasting agents, slurry, emulsion and dynamite are the main classes of explosives used in mine blasts.

20.4.1

Dry blasting agents

Dry blasting agent is a term given to components of an explosive which themselves are not classified as explosives but when mixed together they form a mixture that can explode.

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In large surface mines the most commonly used explosive mixture is called ANFO because it is formed by mixing ammonium nitrate with fuel oil. Even after mixing these two components the final product remains fairly dry because the percentage of fuel oil in the mix is very small and moreover the fuel oil is absorbed in the pores of the prills of ammonium nitrate. Use of standard diesel oil no. 2 as a fuel in ANFO has become almost universal. Other used oils, such as engine oil or compressor oil etc. should not be used because they are not well absorbed into the prills. Moreover, they also form harmful gases, and particularly metallic free radicals which remain unchanged in the atmosphere over a long period of time. A lot of research has been carried out about ANFO because of its extensive use in mining practice. The following are some of the outcomes of the research. In terms of chemical equations, if three molecules of ammonium nitrate react with one link of fuel the end products are harmless, as shown by the following equation. 3NH4NO3 + CH2 ⇒ 3N2 + 7H2O + CO2 + 940 kcal/kg The above means that for a perfect reaction as above the weight of ammonium nitrate should be equivalent to molecular weight of its three molecules and the weight of fuel oil should be equivalent to its one link. From this it can be concluded that with each 94.3 kg of ammonium nitrate about of 5.7 kg of diesel oil should be used. If the weight proportion of AN and FO is 92 to 8, the equation happens to be 2NH4NO3 + CH2 ⇒ 2N2 + 5H2O + CO + 820 kcal/kg Similarly, if the weight proportion of AN and FO is 96.6 and 3.4, the equation happens to be 5NH4NO3 + CH2 ⇒ 4N2 + 11H2O + CO2 + 2NO + 610 kcal/kg. The energy output as well as the velocity of detonation from the explosion is maximum when the weight proportion of the AN and FO is 94.3 to 5.7, as illustrated by Figure 20.3. The output of poisonous gases is also lowest when the proportion of AM to FO is 94.3 to 5.7 as illustrated by Figure 20.4. Even the sensitivity of ANFO also changes with the proportion of AN and FO, as can be seen from Figure 20.5. As has been stated in chapter 18, the velocity of detonation of ANFO increases as the diameter of the blasthole increases. Data presented in Figure 20.3 are sufficient to prove this fact but data in Figure 20.6 do it more convincingly. There are two disturbing aspects about use of ANFO in large surface mines. The first is the quick evaporation of diesel oil and the second is the high solubility of ammonium nitrate in water. When the atmospheric temperature and humidity are high it becomes essential to add extra fuel in the ANFO mix to take care of the degree of evaporation likely up to the time of detonation. How much extra can be determined by technical common sense.

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Maximum Energy Output kcal/kg

1000

5500

900

5000

800

4500

700

4000 3500

600

3000

500

2500 400 1

2

3

4 5 6 7 8 Percentage of Fuel Oil

9

481

Velocity of Detonation m/s


10

Figure 20.3 Variation of velocity of detonation and energy output with percentage of fuel oil in ANFO mixture.

Gases Evolved mol/100 kg

0.25 0.20 CO 0.15 0.10

NO + NO2

0.05 0.00 0

2

4 6 Fuel Oil Percentage

8

10

Figure 20.4 Variation of quantity of gases evolved by explosion of ANFO with percentage of fuel oil in ANFO mixture.

Relative Difficulty of Initiation with Blasting Cap No. 6

3.0 2.5 2.0 1.5 1.0 0.5 1

2

3

4 5 6 7 Fuel Oil Percentage

8

9

Figure 20.5 Variation of sensitivity of ANFO with percentage of fuel oil in ANFO mixture.

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Velocity of Detonation in m/s

5000 4500 4000 3500 3000 2500 2000 1500

0

50

100 150 200 250 Blasthole Diameter mm

300

350

Figure 20.6 Variation of velocity of detonation with diameter of blasthole.

If the humidity of the atmosphere is very high the ammonium nitrate, being highly hygroscopic, absorbs a large quantity of water and in the process becomes less effective. The decrease in velocity of detonation with percentage of water in the ANFO mix is shown in Figure 20.7. In this context it is worth noting that after about 9% water content, the ANFO mix becomes insensitive and fails to detonate. Heavy ANFO is the name given to an ANFO mix that contains aluminum as a sensitizer. Such mixture is also called ALANFO. When ALANFO detonates some molecules of AN chemically react with the molecules of FO and evolve energy as per the chemical equation mentioned earlier. Due to sufficient energy generated in such reaction, some molecules of AN also react with aluminum in the mixture and evolve even higher energy as shown by following equations. 2Al + 3NH4NO3 ⇒ 3N2 + 6H2O + Al2O + 1650 kcal/kg If the proportion of aluminum in the mix is higher then the equation is as under: 2Al + NH4NO3 ⇒ N2 + 2H2 + Al2O3 + 2300 kcal/kg The above equations clearly indicate that the energy yield of ALANFO is higher than ANFO. The variation of energy yield with percentage of aluminum in ALANFO is presented in Figure 20.8. ALANFO is particularly useful for blasting hard rock masses. The most commonly used percentage of Al in ALANFO is between 10 to 15%. The aluminum powder to be mixed with ANFO for making ALANFO must have particle size between 20 to 150 mesh i.e. 0.853 to 0.104 mm. The purity of the aluminum should be more than 94%. As this purity is easily attainable, even the low grade

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3600 3300 3000 2700 2400 2100 1800 0

Detonation Failure 2

4

6

8

10

Percentage of Water in ANFO Figure 20.7 Variation of velocity of detonation with percentage of water in ANFO mix.

Relative Energy Yield ALANFO/ANFO

1.5

1.4

1.3

1.2

1.1

1.0

0

5

10 15 20 Aluminum Percentage

25

Figure 20.8 Variation of energy yield with aluminum percentage in ALANFO.

aluminum can also be used for mixing in ANFO provided the impurities do not yield harmful products. Table 20.3 gives a general idea about the properties of ANFO. For any calculations or taking any type of decisions, similar data should be obtained from the manufacturers.

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Table 20.3 Properties of ANFO explosives. Property

Poured ANFO

Package ANFO

Heat of Explosion in kJ/kg Gas Volume in L/kg Temperature of Explosion in °C Oxygen Balance in %O2 Velocity of Detonation in m/s (65 mm non confined) Velocity of Detonation in m/s (95 mm Blasthole) Detonation Pressure in GPa Brisance According Hess in mm Strength According Trauzl Test in cm3 Relative Working Ability in % Gap Sensitivity in mm for 30 mm Cartridge Gap Sensitivity in mm for 65 mm Cartridge Bulk Density in kg/m3 Density in Cartridge in kg/m3 Detonation by Blasting Cap Water Resistance Fume Quality

4079 928 2749 0.17% 4000

4079 928 2749 0.17% 4000

4100

4100

2–6 15 400 89 -

2–6 15 400 89 40

-

20

Temperature Resistance −18 to −38°C Smallest Diameter of Blasthole in mm Shelf Life in Months Availability

>800 No. 8 Poor Good (Depends on Conditions) Poor Above 32°C 60 In Truckloads or 25 kg Bags

1050 No. 8 Very Good for Intact Package Good to Very Good (When Cartridges are Intact) Poor Above 32°C 30 6 Cartridge L × D mm 30 × 135, 65 × 585, 75 × 585, 90 × 585

The forms in which ANFO is supplied are as follows: 1 2 3

Poured ANFO Packaged ANFO Heavy ANFO

20.4.1.1

Poured ANFO

Where the quantity of explosives is large – such as in larger surface mines, this form of ANFO is supplied in separate component containers on a truck, mechanically mixed at the worksite and poured into blastholes. Hence, it is also called bulk ANFO. Where the requirement is low it is usually supplied in nylon bags. Poured ANFO proves more effective than the packaged form, as it fills the entire cross section of the blasthole whereas the package leaves a gap between the walls of the blasthole and external diameter of the package.

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20.4.1.2

485

Packaged ANFO

If a blasthole has a significant quantity of ground water seeping into it, it cannot be charged by poured ANFO. In such cases the packaged form offers a good alternative. In this form ANFO is premixed and packed into thick cylindrical plastic bags sealed at both the ends. Packaged ANFO evolves larger quantities of poisonous gases from their blasts due to plastic packing. To reduce this cylindrical textile or cardboard tubes with plastic liners are used. Charging a blasthole with packaged ANFO is more tedious and time-consuming. 20.4.1.3

Heavy ANFO

Heavy ANFO is so called not only because it has higher density but it is also meant for “heavy duty” blasts in hard rock in large mines. It is a mixture of ANFO and emulsion explosive in which the interstitial space between the ANFO prills is filled by the emulsion. When the proportion of emulsion in the mix increases, the strength, velocity of detonation, and water resistance increase. With a normal range of compositions, density increases with emulsion contents up to a maximum of 1.3 kg/L. The charge sensitivity varies inversely with density and emulsion contents. Like bulk ANFO this form is also mixed at the bench and immediately loaded into a blasthole. The proportion of emulsion and ANFO in the mix can be chosen while the products are being mixed so the mix has the desired properties. The cost of heavy ANFO rises with increasing amount of emulsion.

20.4.2

Slurry

Slurries are also called water gels. They are made from ammonium nitrate partly in an aqueous solution. Their further classification into a blasting agent or an explosive depends upon other ingredients used with them. As slurry explosive is fluid, it is pumpable and miscible with water. They get good water resistance from the gaur gum mixed with them. “Slurry boosting” is practiced when slurry and a dry blasting agent are used in the same blasthole. Most of the charge will come from the dry blasting agent. Boosters placed at regular intervals may improve fragmentation. Slurries are available in highly viscous tooth paste-like form as well as in cartridge form. Slurries are also made dimensionally stable and water-resistant by adding crosslinking agents. Slurries cost more, give somewhat unreliable performance and deteriorate with long storage. Water gel explosives, a special form of slurry explosive, contain significant amounts of water and separate oxidizer and fuel components, making them less sensitive than water-free nitroglycerin dynamites. Water gels are made up of oxidizing salts and fuels dispersed in a continuous liquid phase. The addition of gelants and cross-linking agents thickens the mixture and makes it water-resistant. Ammonium nitrate, sodium nitrate, and calcium nitrate are the commonly used oxidizing salts, whereas aluminum, coal, sugar, ethylene glycol, and oil are common fuels. The addition of nitrate salts of

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organic amines, nitrate esters of alcohol, perchlorate salts, fine particle size aluminum, or other explosives may adjust the sensitiveness of the water gels. Also, physical sensitizers such as plastic bubbles or glass ‘micro balloons’, may be blended into the gel either alone, or in combination with other chemical sensitizers. The density of most water gels ranges from 1.0–1.35 g/cc. Water gels may be formulated to be either cap sensitive, or insensitive. In either case they are less sensitive to abusive impact, shock, or fire than dynamite. They are explosives, however, and should never be abused. The sensitivity of water gels is affected by temperature, with higher temperatures increasing the sensitivity of the products. Slurry explosives are supplied in two forms viz. bulk slurry and cartridged slurry. 20.4.2.1

Bulk slurry

Bulk slurry explosives are supplied in component form contained in tanks on an offhighway truck. They are thick and appear as shown in Figure 20.9. They can be pumped into a blasthole through tubes. Their critical diameter is relatively small and can be used for charging blastholes of diameters ranging between 50 to 150 mm. For blastholes of larger diameter it is more economical to use ANFO with addition of slurry in certain portions of the blasthole depending upon the need dictated by the site conditions. 20.4.2.2

Cartridged slurry

When slurry explosives are to be used in blastholes of smaller diameter, pumping it inside the blasthole is not advisable. In such case cartridges filled with slurry are used. Cartridges are made from hard plastics or cardboard and are rigid. Since they do not deform easily they can be lowered in a blasthole very easily. Cartridges are available in different diameters and lengths to suit blastholes of different diameter.

AN Oxidizer Gel Ring Liquid Phase

Solid Fuel

Figure 20.9 Emulsion explosive.

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20.4.3

487

Emulsions

Emulsion explosives are formed by four components viz. an oxidizing agent (usually AN), water, combustible material (mineral oil, paraffin or microcrystalline wax) and gas-filled micro balls of glass or special plastic. The water droplets, in which ammonium nitrate is dissolved, are covered by a very thin film of fuel oil. An emulsifier is used for stabilization of these oxidizer droplets. These droplets are dispersed in the liquid phase. The ratio of oxidizer to fuel in an emulsion is typically 9:1. The liquid phase also contains solid fuel. Voids in the form of microballs make the emulsion more sensitive. The contents of a typical emulsion are shown in Figure 20.9. Sensitivity of emulsions decreases with increasing density. Dry products like aluminum powder are used to adjust strength. Gasifying products are used for reducing density. Emulsions have excellent water resistance, are relatively insensitive to temperature changes, have high energy and give dependable performance. Emulsions provide increased explosive efficiency because both oxidizer and fuel phases are liquid and the dispersed nitrate solution droplets are microscopically small i.e. about 0.001 mm. As they are tightly packed within the fuel phase the contact surface is very large. This ensures very rapid reaction. Properties of a particular variety of emulsion are presented in Table 20.4. The direct cost of an emulsion explosive is higher but this is offset by time saved in loading.

Table 20.4 Properties of an emulsion.

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Property

Emulsion

Heat of Explosion in kJ/kg Gas Volume in L/kg Temperature of Explosion in °C Oxygen Balance in %O2 Velocity of Detonation in m/s (30 mm non confined) Velocity of Detonation in m/s (65 mm non confined) Detonation Pressure in GPa Brisance According Hess in mm Relative Working Ability in % Gap Sensitivity in mm in 50 mm non Confined Cartridge Density in Cartridge in kg/m3 Detonation by Blasting Cap Water Resistance in Small Dia. Cartridge Fume Quality Temperature Resistance –18 to –38°C Smallest Diameter of Blasthole in mm Shelf Life in Months Availability

2800 800 1800 + 0.5 >4400 5100 4–9 20 63 30 >1050 No. 8 10 hrs/0.35 MPa Good to Very Good Good 30 12 Cartridge D × L mm 30 × 680, 38 × 393, 50 × 540, 65 × 660, 75 × 520, 90 × 450

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20.4.4

Dynamite

The original dynamite made by Alfred Nobel was a mixture of nitroglycerine and kieselguhr i.e. diatomaceous earth. The diatomaceous earth absorbed the oily nitroglycerine and the mixture became quite insensitive to shock. It could be used far more safely than nitroglycerine. Over the years formulations of dynamite have changed but nitroglycerine has still remained the main detonating component. Other components of modern dynamites are nitrocellulose, ammonium nitrate, sodium nitrate, wood pulp, carbonaceous fuel, sulfur, antacid etc. Three basic types of dynamites are granular, gelatin and semi-gelatin. 20.4.4.1

Granular dynamite

Granular dynamite is in granular form but is still very sensitive and cannot be poured like ANFO. Granular dynamite is available in cartridges of different diameters and lengths. There are two types of granular dynamite viz. straight dynamite and extra dynamite. Straight dynamite has all the common components such as nitroglycerin, sodium nitrate, carbonaceous fuel, sulfur, antacid and moisture but is without ammonium nitrate. Extra dynamite additionally contains ammonium nitrate. Both straight and extra dynamite have a pungent, sweet odor because they contain nitroglycerin. Inhalation of the nitroglycerin fumes will usually cause a severe and persistent headache. Straight dynamite generally has a light tan to reddish-brown color. In extra dynamite the reddish tint is lesser. The texture of straight dynamite can be described as a loose, slightly moist, oily mixture. Extra dynamite, in addition, has a pulpy appearance. Ammonia dynamite is less sensitive to shock and friction than straight dynamite because a part of the nitroglycerin content has been replaced with ammonium nitrate and nitroglycol. Compositions of some of the straight and extra dynamites are given in Table 20.5 and Table 20.6. The properties of straight and extra dynamite presented in Table 20.7 are indicative. In any case when authentically confirmed data on properties is needed, the manufacturer should be contacted.

Table 20.5 Composition of some straight dynamites. Ingredient proportion by weight

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Nitroglycerin

Sodium nitrate

Carbonaceous fuel

Sulfur

Antacid

Moisture

56.8 49 39 29 20.2

22.6 34.4 45.5 53.3 59.3

18.2 14.6 13.8 13.7 15.4

2 2.9

1.2 1.1 0.8 1 1.3

1.2 0.9 0.9 1 0.9

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489

Table 20.6 Composition of some extra dynamites. Ingredient proportion by weight Nitroglycerin

Sodium nitrate

A. nitrate

Carbonaceous fuel

Sulfur

Antacid

Moisture

22.5 16.7 16.5 12.6 12

15.2 25.1 37.5 46.2 57.3

50.3 43.1 31.4 25.1 11.8

8.6 10 9.2 8.8 10.2

1.6 3.4 3.6 5.4 6.7

1.1 0.8 1.1 1.1 2.2

0.7 0.9 0.7 0.8 0.8

Table 20.7 Properties of straight and extra dynamite. Property

Str. dynamite

Extra dynamite

Heat of Explosion in kJ/kg Gas Volume in L/kg Temperature of Explosion in °C Oxygen Balance in %O2 Velocity of Detonation in m/s (65 mm non confined) Velocity of Detonation in m/s (28 mm non confined) Detonation Pressure in GPa Brisance According Hess in mm Strength According Trauzl Test in cm3 Relative Working Ability in % Gap Sensitivity in mm in 90 mm cartridge Density in Cartridge in kg/m3 Detonation by Blasting Cap Water Resistance in Large Dia. Cartridge Water Resistance in Small Dia. Cartridge Fume Quality

4680 717 3400 +1.5 6400

>4100 858 >3000 +2.2 >6000

>6000

2400

4–7 22 380 70 100 1450 No. 8 24 hrs/0.8 MPa 10 hrs/0.3 MPa Poor to Good (When Cartridges are Intact) Poor to Good 22 12 Cartridge D × L mm 22 × 220, 25 × 220, 28 × 220, 38 × 320, 50 × 440

3–6 >14 >385 >78 >40 >1300 No. 8 4 hrs/0.3 MPa 2 hrs/0.3 MPa

Temperature Resistance –18 to –38°C Smallest Diameter of Blasthole in mm Shelf Life in Months Availability

20.4.4.2

Poor to Good 28 9 Cartridge D × L mm 28 × 220, 38 × 320, 50 × 440, 65 × 550, 70 × 450, 80 × 450, 90 × 480

Gelatin dynamite

Gelatine is a name added to dynamite when the composition of dynamite includes nitrocellulose. Nitrocellulose is of cotton-like form with significant volume of voids. It can absorb the nitroglycerine mixed with it to attain a very viscous gel form.

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Gelatin dynamite is also made in three forms viz. straight gelatin, extra gelatin and semi gelatin. Straight dynamite contains nitrocellulose in higher proportion but lacks ANFO or other oxidizers. Here the principal source of energy is nitroglycerin. In ‘ammonia’ or so-called ‘extra’ dynamites, ammonium and sodium nitrates along with carbonaceous fuel replace a large portion of the nitroglycerin to create a less expensive and more impact-resistant dynamite. In these dynamites the ammonium and sodium nitrate along with carbonaceous fuel act as the principal source of energy, while nitroglycerin acts as a sensitizer. Compositions of some varieties of straight and extra gelatin dynamite are given in Tables 20.8 and 20.9 respectively. Semi-gelatin dynamites do not contain any ammonium nitrate prills. Therefore, they are in semi-gel form rather than granular. Compositions of some varieties of semi-gelatins are given in Table 20.10. Properties of straight, extra and semi-gelatins are presented in Table 20.11. The data presented in the table are from different manufacturers. Hence, the values may not be very compatible.

Table 20.8 Composition of straight gelatins. Ingredient proportion by weight Nitroglycerin

Nitro cellulose

Wood pulp

Antacid

Moisture

91

7.7

0.3

0.8

0.2

Table 20.9 Composition of some extra gelatins. Ingredient proportion by weight Nitroglycerin Sodium nitrate

Nitro cellulose A. Nitrate Carbonaceous fuel Sulfur Antacid

Moisture

35.3 29.9 26.2 22.9

0.7 0.4 0.4 0.3

1.7 1.6 1.4 1.5

33.5 43 49.6 54.9

20.1 13 8 4.2

7.9 8 8 8.3

3.4 5.6 7.2

0.8 0.7 0.8 0.7

Table 20.10 Compositions of some semi gelatins. Ingredient proportion by weight

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Nitroglycerin

Sodium nitrate

Nitro cellulose

Carbonaceous fuel

Sulfur

Antacid

Moisture

49.6 40.1 32 25.4 20.2

38.9 45.6 51.8 56.4 60.3

1.2 0.8 0.7 0.5 0.4

8.3 10 11.2 9.4 8.5

1.3 2.2 6.1 8.2

1.1 1.2 1.2 1.2 1.5

0.9 1 0.9 1 0.9

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491

Table 20.11 Properties of gelatin explosives. Straight gelatin

Extra gelatin

Semi gelatin

Energy kcal/kg

1055

1055

955

Energy kcal/L

1510

1510

1240

Gas Volume in mol/kg

32

32

37

Detonation Pressure in kbar

130

106

60


6000

5300

4300

Relative Bulk Strength wrt ANFO in %

210

210

172

Relative Weight Strength wrt ANFO in %

120

120

109

Fume Class

IME1

IME1

IME1

Density in Cartridge in kg/m3

1.5

1.51

1.3

Detonation by Blasting Cap

No. 8

No. 8

No. 8

Water Resistance

Excellent

Excellent

Good

Smallest Diameter of Blasthole in mm

32

25

29

Availability

Cartridge D × L mm – (25, 29, 32, 40, 45, 50, 75) × 200, (32, 50, 60, 65, 70, 75) × 400

Cartridge D × L mm – (25, 29, 32, 40, 45, 50, 75) × 200, (50, 60, 65, 70, 75) × 400

Cartridge D × L mm – (25, 29, 32, 40, 50) × 200, (50, 60, 65, 70, 75) × 400

The use of dynamite as a main explosive has declined considerably ever since ANFO entered mining as an explosive. With rare exceptions it is hardly ever used in large surface mines.

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Chapter 21

Tests on explosives

21.1

INTRODUCTION

All explosives are extremely dangerous. Even one kilogram of many explosives is sufficient to kill hundreds of persons. To ensure that no disaster happens, Governments of most countries lay down very stringent requirements about various aspects related to the properties of explosives. Manufacturers of explosives, in turn, have to carry out many tests on their products to ensure that the characteristics of explosives are within the limits permitted by the governing authorities. This chapter is devoted to giving details of many tests usually identified as essential to resolve many issues related to the hazards analysis of existing or proposed facilities, process, or products. There are literally hundreds of tests meant to measure the properties of explosives. Only a few important ones have been briefly described here.

21.2 TESTS FOR MEASUREMENTS OF VOD VOD is the acronym of Velocity of Detonation. Velocity of detonation is measured by using many different techniques. Tests commonly used for laboratory measurement of velocity of detonation are: 1 2

D’Autriche method Chronograph method

Velocity of detonation measured in the laboratory differs significantly from the one experienced in a blasthole. Measurements of velocity of detonation in a blasthole are carried out by the following methods. A B C D

Fiber optic sensor method SLIFER method Optical measurement method CORRTEX method A brief overview of all these methods is as under.

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21.2.1

D’Autriche method

Actually this method compares the velocity of an explosive with that of a detonating cord that has a known velocity of detonation. From the result the detonation velocity of the explosive can be found precisely. This test is carried out in a laboratory. In actual tests a detonating cord having a precisely known velocity of detonation is inserted in a tube loaded with explosive to be tested at points A and B as illustrated in Figure 21.1. Midpoint C of the cord length is kept on a lead plate and its position on the plate is marked. When the explosive tube is detonated from one end and the detonation zone reaches point A, the cord also gets detonated. A detonation wave in the cord starts traveling towards the other end. After some time interval the detonation zone in the explosive tube travels a distance d and reaches point B. Here the other end of the detonation cord is also ignited. A detonation wave in this end of the cord also starts traveling towards the first end. Since the plate is of lead, which has high malleability and low melting point, a mark is easily visible at the point D where the two detonation zones in the cord meet. By precise measurements of the distances dak, and length AC and the velocity of detonation in the cord, the velocity of detonation in the explosive tube can be calculated by using the following equation. Ve = Vc * d/(2 * k) where Ve = Detonation velocity of explosive (m/s) Vc = Detonation velocity of cord (m/s) d = Length of explosive between two ends of the cord (m) k = Distance between point of collision of detonation in cord and midpoint of cord (m)

21.2.2

Chronograph method

In this test, carried out in a laboratory, two electric probes, that detect detonation and which send a signal through a wire attached to them, are placed at a certain distance d Tube Filled with Explosive

B D

Detonating Cord k

d

C

A

Lead Plate Point of Detonation Initiation

Figure 21.1 D’Autriche method of detonation velocity measurement.

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495

in an explosive-filled tube. The other ends of the probe wires are connected to a special instrument, called a chronograph. A chronograph measures the time interval between two signals sent to it to an accuracy of one microsecond. This way the time interval t, for the travel of the detonation zone over a distance d is known. Most such instruments are able to directly calculate and show the velocity of detonation when the value of d in terms of appropriate units is fed through an input device. In modern instruments, fibre optic cables are used for better accuracy. These cables transmit the signals in the form of light rather than electric current.

21.2.3

Fiber optic sensor method

In this method a specialized fiber optic cable of 1 mm diameter with smaller holes drilled across it at specific uniform distances, is used. This cable is placed in the column of the explosive to be tested. When the explosive is detonated and the detonation reaches the hole the air confined in the hole is compressed and ionized. This generates a light pulse in the cable. The timing recorded for the first pulse is 0.000. Later as the detonation zone travels in the explosive column, the light pulsed generated by second and third and subsequent holes are sent to the recorder where their arrival timings are measured. Since the uniform distance, d, between two consecutive holes is known, the velocity of detonation can be calculated from the recorded timings.

21.2.4

SLIFER method

SLIFER is an acronym of Shortened Location Indication by Frequency of Electrical Resonance. In this case a standard coaxial cable RG6U with a resistance of 75 ohm is placed centrally in the column of explosive in a blasthole. This cable forms the part of an oscillator circuit where the frequency measured by the instrument is dependent upon the length of the coaxial cable. When the detonation zone in the blasthole Advances, the cable burns and the effective length of the cable reduces. With this, there is an increase in the frequency measured by the instrument. As there is a well-defined relation between length of the cable and the frequency, the continuous measurement of frequency and time gives velocity of detonation in the blasthole. Instruments capable of giving a graphical display of velocity of detonation as shown in Figure 21.2 are available. Since the measurements of frequency are taken continuously the instrument can also display the instantaneous velocity of detonation throughout the explosive column in the blasthole. Currently this method of measurement of velocity of detonation is most widely used in surface mining practice.

21.2.5

Optical measurement method

This method is mainly used for measuring the speed at which a shock wave travels through a shock tube and the time taken from initiation of the shock tube to detonation of the blasting cap at the end of the shock tube.

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Figure 21.2 Graphical display of velocity of detonation.

A shock tube is a 3 mm diameter plastic tube lined with a coating of explosive powder. It is used as an accessory in blasting. The nanosecond light pulse generated, when the explosive in the shock tube detonates, is used for starting the timer in the measuring instrument. When the detonator at the other end of the shock tube explodes the wire wound around it breaks and the timer stops. The reading of the time recorded can be used for calculation of the speed of shock wave travel and time for initiation of the blasting cap. Apart from the above, some more techniques, such as CORRTEX, are also used to measure the velocity of detonation of an explosive.

21.3 TESTS FOR MEASUREMENT OF STRENGTH Several tests have been used in the past for evaluation of the strength of an explosive. In these tests a parameter related to the strength of explosive is measured. Most commonly used tests are 1 2

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Ballistic mortar test Cratering test

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Tests on explosives

3 4 5 6

497

Underwater test Plate dent test Traulz lead block test Cylinder compression test These tests are very briefly explained hereunder.

21.3.1

Ballistic mortar test

In this test a small charge of explosive is detonated in a chamber. The exhaust of the chamber is used for oscillation of a pendulum. The angle in which the pendulum oscillates depends upon the energy released by the detonation as also the weight and location of center of gravity of the pendulum. The explosive energy can be calculated from known values of the weight of the pendulum and its CG (Center of Gravity) position and measured value of angle of oscillation.

21.3.2

Cratering test

The depth of a crater formed in a rock mass by the detonation of an explosive placed at its surface is related to the energy released by the detonation. Based on this principle, an explosive of specific weight is detonated on the surface of the rock to measure the critical and optimum depth of the crater. Critical depth is the maximum depth at which an explosive can be placed in the rock and yet a crater is formed. Optimum depth is the maximum depth at which an explosive can be placed in the rock and maximum volume of crater is achieved. These depths and the dimensions of crater formed are then used to get the strength of explosive.

21.3.3

Underwater test

In this test a specific mass of explosive is detonated in a large pond filled with clear water. The explosive, while it is being detonated, is placed at sufficiently great depth of the water so that air bubbles do not form and escape to the surface of water. Upon such detonation a pressure wave is generated and travels spherically in all directions. The pressure of this wave is measured at a suitable distance. From a score of tests carried out, equations have been developed to find the energy output of the explosion. This test gives more accurate results than most other tests.

21.3.4

Plate dent test

In this test a specific weight of explosive charge is kept on the top of a plate and detonated. Diameter, depth and volume of the dent caused by the detonation of explosive is measured and is then correlated with the strength of the explosive. In the case of low strength explosives, or where a very small quantity of explosive is to be used, an aluminum plate is used instead of the usual steel plate.

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21.3.5 Traulz lead block test The Traulz lead block test has been commonly used for measuring the strength of explosives for more than a century. The set-up for this test is shown in Figure 21.3. In actual tests a cylindrical lead block having a diameter and length of 200 mm, as shown in the figure, is taken and a hole of diameter 25 mm is drilled in its center to a depth of about 125 mm. The weight of the explosive to be used for the test is 10 g. The explosive is wrapped in an aluminum foil and is placed centrally in the hole. After filling the remaining space with sand the explosive is detonated electrically. Detonation causes the volume of the cavity to increase. The increase in volume is arrived at after deducting the original volume of the cavity and the volume of explosive from the newly measured volume. The result is expressed in cm3 and is called the Traulz number of the explosive. In the case of explosives that yield a huge quantity of gases or those which have a high shattering effect, the lead block cracks and gives incorrect readings. In such cases an aluminum block is used. Traulz numbers for some of the explosives are given in column 2 of the Table 21.1. 200 mm

Hole of Dia. 25 mm

Detonating Wires

Lead Block 125 mm 200 mm

Electric Detonator

Figure 21.3 Traulz lead block test. Table 21.1 Relative strength measured by lead block test.

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Explosive

Gas volume in cc

Relative strength

Blasting Gelatine (8% Collodion Cotton) Gun Cotton (13% N) Ammonal Gelatine Dynamite Tetryl Dynamite No.1 Picric Acid Trinitrotoluene Collodion Cotton (12% N) Mercury Fulminate Ammonium Nitrate

520 420 400 400 350 325 290 260 250 150 130

100 81 77 77 67 63 56 50 48 29 25

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Tests on explosives

21.3.6

499

Cylinder compression test

This test is meant to measure relative strength. In the cylinder compression test a solid lead cylinder of diameter 40 mm and height 60 mm is placed upon a rigid steel plate. On the top of this a steel plate of 4 mm thickness and 40 mm diameter is placed. Explosive of weight 100 g, contained in a shell of 40 mm diameter, is placed on the top of this steel plate. All this set up is shown in Figure 21.4. Reduction ΔH in the height of lead cylinder after detonation is measured in mm. If ΔHR is the distance measured in the cylinder compression test for the reference explosive, and if ΔHt is the distance measured for the explosive being tested, the relative strength of the explosive under test is ΔHt/ΔHR

Explosive 100 g

Original Height 60 mm 4 mm

Reduction ΔH

Steel Plate 60 mm Height Lead Block 40 mm Dia.

Figure 21.4 Cylinder compression test.

Table 21.2 Relative strength of some explosives measured by cylinder compression test.

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Explosive

Relative strength

Explosive Gelatine (Made by Vogue’s Process) Hellhoffite Nitroglycerin (Made by the Old Process) Noble’s Smokeless Powder Gun Cotton (Made as per procedure used in 1889) Dynamite Amide Powder Silver Fulminate Mercury Fulminate Mortar Powder

106.17 106.17 100 92.38 83.12 81.31 69.87 50.27 49.91 23.13

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500


The relative strength of some explosives on the basis of cylinder compression is given in Table 21.2. Explosive strength can also be calculated from some other easily measurable parameters. The equations proposed for this purpose are empirical in nature. Some equations are given here below.

21.4 TESTS FOR MEASUREMENT OF SENSITIVITY Safety is the most important aspect in the context of explosives. Many tests are carried out on the explosives to ascertain their safety from the viewpoint of different properties mentioned earlier. Some of the tests are: 1 2 3 4 5 6

Shock sensitivity test Heat resistance test Cap sensitivity test Gap sensitivity test Friction sensitivity test Electrostatic discharge sensitivity test

Some details of these tests are presented in the following subsections.

21.4.1

Shock sensitivity test

This test measures the shock sensitivity of an explosive. Shock sensitivity is the reciprocal of impact resistance of an explosive. In actual tests a weight of either 2 kg or 5 kg is freely dropped on a sample of explosive weighing 100 mg placed on an anvil. Initially the drop height is kept at 10 mm. If the explosive does not explode by the impact, the drop height is increased by ten mm. The sequence is repeated and the height of the drop at which the explosive explodes is noted. Different explosives give different heights of drop. These are indicators of their impact resistance. The set up of the drop weight test is shown in Figure 21.5. The impact resistance of some explosives measured with 2 kg weight are given in Table 21.3.

21.4.2

Heat resistance test

This test is carried out to investigate the mode of behavior of an explosive subjected to a gradually increasing thermal environment. In actual tests a sample of an explosive is subjected to a temperature of the surrounding air slowly increasing at the rate of 3.3°C per hour until a reaction occurs. Time elapsed and temperature of the surroundings is continuously recorded. What type of reaction takes place is also noted. In cases where the explosive explodes, whether any fragmentation of material has taken place, or whether a crater has formed and what the dimensions of the crater are etc., are noted.

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Figure 21.5 Drop weight test for measuring shock sensitivity.

Table 21.3 Impact resistance measured with 2 kg weight.

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Explosive

Falling distance in cm

Mercury Fulminate Nitroglycerin Nitrocellulose (Dry) Dynamite (25% kiselguhr) Blasting Gelatine (7% Collodion Cotton) Picric Acid Trinitrotoluene Black Powder Dinitrobenzene

2 4 5 to 10 7 12 25 47 to 90 70 200

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502


To reduce the time required for the test a sample is preheated to about 55°C below the previously known reaction temperature of the explosive.

21.4.3

Cap sensitivity test

This test is used to determine susceptibility of explosives to detonation from the energy delivered by an electric blasting cap. The most commonly used blasting cap is no. 8, which comprises 1600 mg of mercury fulminate and 400 mg of potassium chlorate. In the actual test about 1 L of explosive is placed in a cardboard tube. In the case of slurry explosives or emulsions a polypropylene bottle is used. The tube or cartridge is placed on the top of a witness plate made of steel. The detonator (blasting cap no. 8) is inserted in the explosive sample and detonated. If detonation of the sample explosive takes place, the witness plate is torn or penetrated. This can conclusively prove if the sample explosive is cap sensitive or not. The sample explosive is considered a high explosive if it fails the No. 8 cap test and a low explosive if it passes.

21.4.4

Gap sensitivity test

This test is used to find the minimum distance required to prevent sympathetic detonation i.e. the detonation of a test explosive by the shock wave generated by a donor explosive. There are many versions of the test. Only the simplest among them is described below. A cartridge of explosive of length 200 mm is divided into two by a cut in the middle. One of these parts is used as a donor and other is a receptor. A no. 6 detonator is inserted in the donor. The receptor is kept in the same line as the donor at a certain gap distance and the donor is detonated. If the receptor also detonates by the shock, the test is repeated with a larger gap and the maximum gap at which the receptor detonates is determined. In cases where the test is to be carried out in confined conditions, the explosive cartridge parts are inserted in a steel tube having a thickness of 6 mm and internal diameter 5 to 6 mm more than the diameter of the cartridge. The length of the tube must be at least 200 mm more than the total length of the two cartridges including the gap kept between the cartridges. To mimic field conditions the space in the tube at the end of donor and receptor explosive is filled with wetted sticky earth and the gap between the cartridge and tube is filled with sand. Here also the test is repeated to find out the maximum distance over which the receptor cartridge detonates by the detonation of the donor cartridge. In the case of a few explosives for a certain gap distance, the receptor does not detonate but deflagrates. Noting such distance from the observations in the test is also important.

21.4.5

Friction sensitivity test

Two versions of tests for friction sensitivity of an explosive are recognized. In the first version, called the ABL Friction Test, a sample of explosive is placed on a sliding anvil. The rotating wheel on top of the anvil exerts a well-defined normal

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503

force on the explosive in a vertical direction. The anvil is made to move at different speeds in a horizontal direction. Instances of deflagration or detonation are noted after each slide of the anvil. Analysis of the observation gives the friction sensitivity of the explosive. In the second version, called the BAM Friction Test, a sample of the explosive is placed on a porcelain disk. A pin with a spherical shaped end exerts pressure on the explosive. The pressure can be varied by adjusting the weight kept on the lever arm. An electric motor attached to the plate moves the plate to and fro. This generates friction and the explosive sample either deflagrates or detonates. Analysis of the observation gives the friction sensitivity of the explosive.

21.4.6

Electrostatic discharge sensitivity test

This test is meant to determine the response of an explosive when subjected to various levels of electrostatic discharge energy. A capacitor with known electric voltage (usually 5000 V) stored in it is connected to a needle. The sample of explosive is placed on the top of a conductor plate. The needle is then moved towards the explosive sample. At a certain distance between the explosive and the needle the explosive deflagrates or detonates. Analysis of the observations in the test gives the electrostatic discharge sensitivity of the explosive.

21.5 TESTS RELATED TO STORAGE OF EXPLOSIVE The various types of sensitivities discussed above are very important from the viewpoint of the handling of the explosives. In many instances explosives have to be stored for long periods of time. Certain properties of the explosives become very crucial in such conditions. Tests related to these are covered in the following subsections.

21.5.1

Effect of wetness test

In this test a 5 g sample of the explosive is kept in surrounding air having a certain degree of humidity for a period of one hour. Any reaction such as degradation, deflagration or detonation that takes place during such storage is noted. The air in the container of the sample is also tested to measure the gases evolved during the period. The test is repeated with higher humidity and higher sample weight. If the explosive does not ignite spontaneously or does not release more than 1 L of inflammable or toxic gas per kg weight of the explosive it is not considered as dangerous when wet.

21.5.2

Internal ignition test

The objective of this test is to find out the response of an explosive to rapidly rising temperature and pressure. These conditions are experienced in adjoining fires or stormy weather. The sample, with an initial temperature of 25°C, is loaded into a schedule 80 pipe with forged steel end caps. A black powder bag igniter is inserted into the center of the pipe and the leads are sealed with epoxy resin. The igniter is fired and the results

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are assessed. Either the pipe or at least one of the end caps must be fragmented into at least two distinct pieces for a positive result. The test is considered negative (the material passes) if the pipe is merely split open or the caps are blown off in one piece. Three trials are performed unless a positive result occurs earlier. The test determines if a material will explode or detonate when ignited under confinement.

21.5.3

Material compatibility test

This test is performed to determine the ability of a material to coexist in intimate contact with an explosive without adverse reaction, for a certain acceptable period of time. This test measures heat increase or heat loss. It is used when looking for an exothermic or endothermic reaction. This test may be also used for composite explosives.

21.5.4 Vacuum stability test When explosives are to be transported through, or are to be stored in, a very low pressure region, this test becomes important because in such an atmosphere apart from the container, the explosive must also be stable. In this test a sample of explosive is placed in surroundings with air pressure of 5 mm Hg for 48 hrs at constant temperature. The volume of gas liberated during this period is calculated after measurement of pressure. If the ingredients of a mixture of explosive are separately tested and then the explosive made by the same weight of each ingredient in a mixture form is tested, and if the gas volume evolved from the mixture is more than the sum of gas volume evolved individually, it means that the ingredients are reacting with each other at a certain rate.

21.6

MISCELLANEOUS TESTS

In the manufacture and use of explosive or other activities closely connected with explosives, failures of the explosives or related disasters can happen. To avoid such occurrences many other tests are performed with the objective of evaluating a specific property of an explosive. Important amongst them are those described below. Some of these are very similar to those described in earlier sections.

21.6.1

Critical diameter test

This test is conducted to measure the minimum diameter of an explosive column that can satisfactorily propagate an explosive detonation. The test uses black seamless steel tubes of schedule 40, class B, type A-53 variety of several different diameters and lengths. The length of the tube is kept at about 3 times the diameter. Explosive filled in the pipe is detonated at one end. A witness plate placed at the other end indicates if the detonation has propagated to the other end. Alternatively, such verification can also be done by velocity probe. If detonation travels to the other end in three of the five tests for the same diameter pipe, the test is

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repeated with a smaller diameter pipe. When detonation does not travel to the other end in three consecutive tests the critical diameter is considered to be the internal diameter of that pipe plus 5 mm.

21.6.2

Critical height test

The purpose of this test is to determine the height of explosive in a metal tube at which the burning reaction of an explosive changes to detonation. The test also uses black seamless steel tubes of schedule 40, class B, type A-53 variety of several different diameters and lengths. The pipe is held in a vertical position and filled with explosive. A 12 gram igniter is kept at the bottom of the explosive column with wires connected to it for initiating the burning reaction. After closing he bottom of the pipe with a threaded cap, the burning is initiated. One of several devices is used to find out the velocity of propagation of the reaction in the explosive. A sudden increase in the velocity of propagation indicates that the burning reaction has transformed to detonation. The height at which such transformation has taken place is treated as the critical height. By repeating the test three times the height reading is reconfirmed. Observations are made by using pipes of different diameters. A plot of internal diameter of the pipes and the critical height for that internal diameter can be used to find the critical height for other diameters.

21.6.3

Bullet impact test

The impact of a bullet is a means of transfer of energy to an explosive. The consequent reaction can be none, deflagration or detonation depending upon the type of explosive and the energy transferred by the impact. In the actual test a 0.3 caliber (7.62 mm diameter) bullet is fired at an explosive mass from a distance of 30 yards (27.431 m). The firing device used in the test is a rapid action gun like an AK 47 capable of firing at the rate of 600 rounds/min at a speed of 856 m/s. Three rounds are fired at the explosive in three different orientations so the penetration of the bullets is in the most sensitive locations of the explosive mass. A post-test inspection of the test film and hardware is performed to evaluate the reaction. If the explosive does not detonate it is given a good rating; if no reaction takes place it is excellent.

21.6.4

Koenen test

This test is used to determine the sensitiveness of a material to the effect of intense heat under vented confinement. In this test, the material is placed in a steel container with an orifice plate. The test apparatus is then placed in a protective steel box, and heated at a specified rate. A series of trials is conducted using different sizes of orifices. A “go” reaction is determined by examining the container. Conducting three successive “no-go” reactions with an orifice plate size above that which produced a positive result concludes the test. This orifice is called the limiting diameter. The limiting diameter may be used to evaluate the degree of venting required to avoid an explosion in the process.

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Chapter 22

Blasting consumable and accessories

22.1

INTRODUCTION

To engineer a blast in a mine, apart from the main explosive, many additional items are required. Collectively all these are referred to as accessories. These can be divided into three categories viz. consumables, instruments and tools. Very careful selection of the accessories is essential so as to perform the blasting operation safely, smoothly and with minimum hazards. In this chapter the various accessories needed in surface blasting operations have been described to the necessary depth. It is to be clearly understood that the variety available in blasting accessories is so wide that it is almost impossible to give details of everything available. Manufacturer’s representatives do give information about their own and occasionally their competitor’s products. This information must be used intelligently in proper selection.

22.2

BLASTING CONSUMABLES

Consumables used in a blast are: a booster cartridge, a primer cartridge, initiation transmission line (ITL), a detonator, and stemming material. Placement of these items in a blasthole is as shown in Figure 22.1. Of these, the primer and boosters are meant to amplify the energy generated by the detonation of the detonator. The most commonly used explosive in primer or booster cartridges is Pentolite which is a mixture of PETN and TNT, usually in the proportion of 50%:50%. Primers and booster cartridges typically have a density of 1.55 to 1.7 g/cc. They have excellent water resistance and detonation velocity ranging between 6000 to 7600 m/s. Often other explosives such as dynamite, water gel/slurries or emulsions are also used in primer or booster cartridges. Both primers and boosters are available in cartridge form as shown in Figure 22.2. Such cartridges are made from casting the explosives in the pre-formed cartridge shells of lined paper. A primer cartridge typically has one through hole and one well within it. The well is meant to accommodate the detonator and the through hole is meant to carry the ITL from the bottom side of the detonator to the ground, where it is connected to the main ITL that leads to an initiating device.

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Top of Blasthole Stemming

Detonating Cord

Main Explosive Booster Explosive Primer Explosive

Detonator or Blasting Cap

Figure 22.1 A charged blasthole.

Primer Cartridge

Booster Cartridge

Figure 22.2 Primer and booster cartridges.

Some manufacturers make a primer cartridge with embedded detonator. In such cases the cartridge may have only one through hole or no hole at all because the ITL can be taken up through the gap between the cartridge and the blasthole wall. Depending upon the design of a blast, multiple primers can be used in a blasthole to detonate different sections of intentionally separated explosive charges. A detonator should be inserted in a primer just before insertion of the assembly in the blasthole for the sake of safety.

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509

If a hole is to be made in a primer cartridge, only non-sparking implements must be used for the purpose so the danger is minimized. Such a hole must be at the center of the cartridge cross section. Booster cartridges may have only one through hole to carry the ITL through them. Boosters to be used in large blastholes do not have any through hole at all because there exists sufficient gap between the booster cartridge and the wall of the blasthole for the ITL to pass through. Since the booster is invariably placed in the vicinity of the primer, detonation of the primary cartridge automatically detonates the booster also. Thus, the booster cartridge does not need an additional detonator. Typical placement of a detonator in a primer cartridge is shown in Figure 22.3.

22.2.1

Initiation transmission line

A primer cartridge is always detonated by an initiating device, placed at a very great distance for the sake of safety. It is, therefore, very essential to transmit the initiating pulse from the initiating device to the detonator through a line called the Initiation Transmission Line and abbreviated as ITL. Three types of ITL are used for engineering a blast in mines. They are: 1 2 3

Safety Fuse for Transmitting Fire Electric Transmission Wires Detonating Cord for Transmitting Shock

22.2.1.1

Safety fuse for transmitting fire

This is the oldest ITL device. It consists of black powder wrapped in a textile tube so it remains in continuity and is protected from mechanical shocks and abrasion. For protecting it from water or moisture, a waterproofing layer surrounds the textile tube. A flame is used to ignite the safety fuse at one end. The fire travels towards the other end at a burning rate of about 7.5–10 mm/s. Safety fuse is used when only a few blastholes are to be blasted by using low explosive, as in secondary blasting. The propagating fire does not have sufficient energy to cause detonation of modern high explosives. Igniter cord is very similar to safety fuse but burns at a rapid rate of 1 to 3 m/s. The core of igniter cord is a pyrotechnic powder mix.

Lead Line / Leg Wire

Primer Cartridge

Lead Line Passed through Cord Well

A detonator inserted in a Cap well

Figure 22.3 Placement of detonator in a primer cartridges.

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22.2.1.2

Electric transmission wires

As the name indicates this ITL transmits electric impulses through conductor wires. The conductor wires are normal copper wires used in house electrification. The electric current required to be passed through the wire to cause ignition is called the ignition current. It can be as much as 25 A. Electric circuit analysis has to be done to ensure that the blast takes place as planned. To ensure that the wire transmits the electric current surge without burning, wires of adequate gage are chosen. Table 22.1 gives details of some conductor wires used in large surface mine blasts. Values of resistance mentioned in this table are required to be used in the absence of more authentic values specified by the manufacturer. 22.2.1.3

Detonating cord for transmitting detonation

Detonating cord is meant to transmit a detonation wave. It is made up of a plastic tube with 3 mm outside diameter. The inner diameter of the tube is 1 mm. It is usually filled with PETN in uniform density. When a detonation is initiated at one end of the tube it travels towards the detonator at a speed of 6500 m/s. The plastic material of the tube has sufficient tensile strength to ensure that it can be easily lowered into a deep blasthole without breaking and has sufficient radial stiffness so it usually does not burst when the shock wave is traveling within it. However, such things depend upon the material density per meter in the detonating cord. Therefore, the manufacturer should be consulted in this regard. Some bulk explosives, when used in small diameter blastholes, are rather sensitive to the compression caused by the detonating cord. If such is the case, shock tubes described in the next section can be used in the blasthole, in conjunction with detonating cords laid on the ground surface. Detonating cord as described above is meant to be attached to a detonator. Detonation in the detonation cord at the free end can be initiated through a No. 6 Detonator. Detonating cords with higher explosive contents must be used for trunk lines that run on the ground surface towards the initiating device. Specifications of some

Table 22.1 Conductor wires used in large surface mine blasts.

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Gauge no.

Diameter of wire in mm

Resistance in mΩ/m

Maximum permissible current in A

6 8 10 12 14 16 18 20

4.1148 3.2639 2.58826 2.05232 1.62814 1.29032 1.02362 0.8128

1.295928 2.060496 3.276392 5.20864 8.282 13.17248 20.9428 33.292

37 24 15 9.3 5.9 3.7 2.3 1.5

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such detonation cords, made by Explosia S. A. of the Czech Republic are given in Table 22.2. Table 22.3 gives the recommendations for use of detonating cords with various explosive contents. The detonation cord makes a lot of noise when the detonation wave travels through the tube. It is therefore necessary to cover the detonating cord laid on ground surface by sandy earth. To ensure complete detonation, a double trunk line or loop system must be used to connect holes in the blast. Cutting detonation cord with a knife is safer than shearing it off by using pliers or wire stripper. 22.2.1.4

Detonating cord for transmitting shock

This is more often called a shock tube. In this type of detonating cord a tube made from very special plastic is used. The inner surface of the tube is coated with a fine layer of a high explosive, usually HMX, and a very fine powder of aluminum. The explosive is held on the tube wall by a static charge. When sufficient shock and ignition is delivered to the tube, the dust explodes and the detonation is propagated through the tube in a fashion similar to a coal dust explosion in an underground mine. The tube is very insensitive to impact and atmospheric heat. The shock wave can be caused only by a special pistol-like device. The shock wave travels through the tube at a speed of about 2000 m/s.

Table 22.2 Specifications of commonly used detonating cords.

Cord name

Color

Explosive contents in g/m

Starline 6 Starline 12 Starline 15 Starline 20 Starline 40 Starline 80 Starline 100

Red Green Blue Yellow Orange Ultraviolet Ultraviolet

6.0 ± 1.0 12.0 ± 2.0 15.0 ± 2.0 20.0 ± 2.5 40.0 ± 4.0 80.0 ± 8.0 100.0 ± 10.0

Velocity of detonation m/s

Outside diameter in mm

Tensile strength in kg

6500 6500 6500 6500 6500 6500 6500

min 3 5.0 ± 1.0 5.2 ± 1.0 6.6 ± 1.0 8.7 ± 1.5 11.5 ± 2.0 13.0 ± 2.0

50 60 60 70 75 75 75

Table 22.3 Recommendations for use of detonating cord.

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Cord explosive contents g/m

Recommended application

1.5–3 6 12–20 40 80–100

Initiation of Primers and Very Sensitive Explosives Trunk Lines Connecting Blastholes Initiation of Conventional and Low Sensitivity Explosives Seismic Exploration Contour Blasting and Demolition

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Non-electric detonating cords, particularly the shock tube, must be handled very carefully to ensure that they are not subjected to any shock, excessive heat, impact etc., as however insensitive the tubes may be, they must be handled carefully to avoid any inadvertent explosion. They should also not be pulled or stretched unduly. 22.2.1.5

Hercudet tube for transmitting fire

In this system the transmission tube is a simple plastic hollow tube, about 4 mm outside and 2 mm inside diameter. It does not contain any explosive or metal. By means of special straight or elbow connectors the tube forms an airtight circuit, much similar to household piping. Hercudet detonators are a part of such a circuit. When the circuit is completed it is first checked for air-tightness by pumping compressed air into it. Once it is confirmed that the circuit is airtight for pressure up to about 300 kPa, gaseous oxidizer and gaseous fuel mixture is fed into the circuit. When the circuit is completely filled with the mixture the gas is ignited. The fire travels throughout the circuit at a speed of about 300 m/s to detonate all the detonators in the circuit.

22.2.2

Detonator

A detonator is often referred to as an initiating device because it initiates the detonation process in a blasthole. It contains a primer and a secondary explosive. The primer is usually a mixture of lead azide, lead styphnate and aluminum powder abbreviated as ASA. Some manufacturers use DDPN i.e. diazo-dinitro-phenol so as to reduce air pollution by lead. PETN or Pentolite is used as the secondary explosive. With such sensitive explosives, detonators become sensitive and are more prone to accidental detonation. They must be handled very carefully. There are three basic types of detonators as under: 1 2 3

Electric Non-electric Electronic

The above classification is based on the source of energy used for starting detonation in the detonator. All three types of detonators can be instantaneous, or with a delay element built into them. Usually detonators are made to have a strength level of 6 or 8. Both these contain 0.2 g of ASA. The No. 6 contains 0.22 g of PETN, whereas the No. 8 contains 0.45 g of PETN. The length of No. 6 detonator is about 35 mm, whereas for No. 8 detonator it is about 42 mm. 22.2.2.1

Delay element

A delay element is in the form of a small tube filled with densely packed pyrotechnic material, usually antimony and potassium nitrate or red lead and silicon. One end of the element is in contact with the primary explosive and the other end is in contact with the initiating element such as an electric spark/heat generator or shock wave

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transferred by the detonating cord. The pyrotechnic material transfers the fire from one end to the other at a specific speed, which is either 3.3 s/m or 33 s/m (303 mm/s to 30.3 mm/s). Thus, depending upon its length, the delay element delays the detonation of the explosives in the detonator by few milliseconds. Delay elements are also available for insertion between detonation cords. Commonly used delays are either from short delay series or long delay series. Information about delays is given in Table 22.4. Delays available may differ from manufacturer to manufacturer. Consideration must be given to factors such as length of explosive column, blasthole diameter, burden, spacing, type of rock and whether the blast is for bench, underground mine, tunnel, shaft etc. 22.2.2.2

Electric detonators

Electric detonators cause the initiation of detonation by an electric current passed through the detonators by electric wires. They have an outer aluminum or copper shell that contains primary and secondary explosives, insulation material, two wires and a delay element if applicable. Use of an aluminum shell is prohibited in gassy underground mines because it generates a very high quantity of heat and causes high brisance. Most electric detonators have a diameter of 7 mm. Table 22.4 Timings for available delays for electric detonators. Nominal time in seconds for

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Delay number

Short delays

Long delays

0 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20

0 0.5 1.0 1.5 2.0 0.5 3.0 3.5 4.0 4.5 5.0 5.5 6.0

0 25 50 75 100 125 150 175 200 225 250 300 350 400 450 500 550 600 650 700 750

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Many developments have taken place in the technology of electric detonators ever since the early modern type detonators made by Gardiner and Smith in 1875. Three different types of electric detonators used in mining blasts, as described here below, have evolved through such developments. 1 2 3

Solid Pack Electric Detonator Fusehead Electric Detonator Exploding Bridgewire Detonator

All these use electric current in a different manner to initiate detonation. Technical specifications of a typical electric detonator are presented in Table 22.5. They are self explanatory and hence not described. 22.2.2.2.1

Solid pack electric detonator

The construction of a solid pack electric detonator is shown in Figure 22.4. As shown there, these detonators contain both primary and secondary explosives. A thin metal

Wires for Electric Connection

Insulating Material

Heating Primary Secondary Explosive Explosive Wire

Figure 22.4 Construction of a solid pack electric detonator.

Table 22.5 Technical specifications of an electric detonators. Specification category Classification Shell Specification Detonator Specifications

Aspect of the detail

Detail

Indian Explosive Rules 1983 UN Number Material Length Strength Leg Wire Color

Class 6, Division 3 0030 Aluminum 42 mm No. 8 White/ White (For all wire lengths) 1.8, 3.0, 4.0, 5.0 Steel 25.5 SWG 0.5 to 0.8 1.6 to 2.2 ohm 3.2 milliwatt.second/ohm 1.2 A DC for Series 0.8 A DC for Single 0.18 A applied for 300 seconds

Leg Wire Lengths in m Leg Wire Material Leg Wire Diameter Leg Wire Resistance in ohm/m Fuse Head Resistance Firing Impulse Firing Current No-Fire Current

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wire, called a bridgewire, embedded in the primary explosive, is used for passing electric current. Bridgewire is made from an alloy comprising nickel, chromium and iron. The current heats up the bridgewire and causes detonation of the primary explosive. This in turn causes detonation of the secondary explosive. Later, the main explosive in the blasthole is detonated by the energy released. Solid pack electric detonators can have a built-in delay element. 22.2.2.2.2

Fusehead electric detonator

In this type of detonator the two ends of the current-carrying wires have a small sheet of insulating material between them for added safety against static electric currents. The bridgewire is soldered across the sides of the insulating sheet. The complete assembly is then dipped in a mixture of potassium chlorate, nitrocellulose and charcoal to form the fusehead. The fusehead is inserted in a container that has an adequate quantity of primer, and booster explosives and a delay element if required. The mouth of the cap is then sealed. In this manner the fusehead, delay element and cap filled with explosives can be made separately to reduce danger. A typical fusehead electric detonator is shown in Figure 22.5. 22.2.2.2.3

Exploding bridgewire detonator

The construction of an exploding bridgewire detonator is very similar to that of a solid pack electric detonator. However, the bridgewire in this case is very thin, i.e. about 0.04 mm in diameter. These detonators are to be detonated by high voltage electric charge. Such high voltage electric charge quickly vaporizes the wire rather than heating it. For this reason the exploding bridgewire detonators are far more instantaneous than the solid pack or fusehead type where a little time in some ms is needed for their heating to the requisite temperature. Apart from being instantaneous in their action, the exploding bridgewire detonators also require high voltage for detonation. Since such voltage is not generated by static electricity, their field use is safer. Construction of a typical bridgewire detonator is shown in Figure 22.6. As stated earlier, to ensure that sufficient electric current passes through the electric detonators to cause their detonation, electric circuits containing detonators and conductor wires have to be analyzed. Details of resistance to electric current flow offered by conductor lines has been given through Table 22.1. The resistance offered by electric blasting detonators and the wire attached to them can be found from Table 22.6.


Insulating Material

Fusehead

Primary Secondary Explosive Explosive

Figure 22.5 Construction of a fusehead electric detonator.

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Insulating Material

Primary Explosive

Secondary Explosive

Figure 22.6 Construction of an exploding bridgewire detonator.

Table 22.6 Resistance of electric blasting detonators. Copper wire conductor

Iron wire conductor

Length of conductor wire in m (ft)

Instantaneous detonators

Delay detonator

Instantaneous detonators

Delay detonator

(4) (6) (8) (10) (12) (14) (16) (20) (24) (30) (40) (50) (60) (80) (100)

1.26 1.34 1.42 1.50 1.58 1.67 1.75 1.91 2.07 2.15 2.31 2.42 2.59 2.71 3.11

1.16 1.24 1.32 1.40 1.48 1.57 1.65 1.81 1.97 2.06 2.21 2.32 2.49 2.61 3.01

2.1 2.59 3.09 3.59 4.09 4.58 5.08 6.06

2.0 2.49 2.99 3.49 3.99 4.48 4.98 5.98

22.2.2.3

Non-electric detonators

Non-electric detonators are also called plain detonators. They are fired by detonating cord instead of electricity. As shown in Figure 22.7, a non-electric detonator consists of a plastic shell filled with primary explosive, secondary explosive and a delay if applicable, and a certain length of detonating cord. The construction of a non-electric detonator is shown in Figure 22.8. The shell has a long empty space, to accommodate one end of the detonating cord. To ensure that the maximum area of the core touches the delay element, the detonating cord is cut in a plane perpendicular to its length. Upon this the aluminum shell is crimped onto the cord so the detonator becomes waterproof and firmly fixed onto the detonating cord.

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Figure 22.7 Non electric detonator with detonating cord.

Outer Covering Reactive Powder Core

Aluminum Crimp Shell Marks

Delay Secondary Element Explosive Primer Explosive

Figure 22.8 Construction of a non electric detonator.

Rubber Cover Two Hollow Tubes

Aluminum Shell

Delay Element

Secondary Explosive

Crimp Igniting Primer Marks Explosive Explosive

Figure 22.9 Construction of an hercudet detonator.

The Hercudet detonator to be used in The Hercudet system is somewhat special. As shown in Figure 22.9 it contains primary and secondary explosive and additionally a very sensitive igniting explosive. Two gas tubes go inside the detonator. These two tubes facilitate filling the complete circuit with gas.

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22.2.2.4

Electronic detonators

As the microelectronic circuits became available easily and cheaply, electronic initiating systems were made by using them. Like electric initiating systems, these systems also use a detonator and electric wires but the detonator in an electronic system works on the basis of digital signals emanating from the initiation device located at a long distance on the ground rather than by electric current. A typical electronic detonator looks like the one shown in Figure 22.10. The most important feature of an electronic initiation system is that it can be tested in the field without causing actual detonation. Appropriate measures eliminate the possibility of any misfire i.e. explosive in a blasthole not getting detonated along with other blastholes. In an electronic detonator when the current from the wires enters first, the voltage of the current is evaluated by the over-voltage protection circuit. When the current is within the acceptable range of voltage it is allowed to go further to the integrated circuit chip which actually controls the further sequence of actions. In the case of higher current the protection circuit burns but no detonation is caused by such burning. Once the current is allowed to pass to the integrated chip, the signals coming from the blasting machine located at far distance are interpreted and only upon receiving a specific code is the capacitor in the detonator allowed to become charged. When the capacitor is charged adequately, the integrated circuit sends a ready signal to the blasting machine. The machine gives an indication about readiness of all the detonators for blasting, and the blasting machine operator can then send a very specific signal to initiate the blast by the release of charge from the capacitor after a preset delay. The electronic initiation system is considered to be the safest amongst all the initiating systems. Of the numerous advantages of the electronic initiation system, the following are a few: 1 2 3 4

5

The detonators do not have any energy of their own and therefore no accidental detonation can take place. The integrated circuit must receive specific signals to start a further chain of actions. Pre-detonation testing of the complete blast is possible. Delay timing of any of the detonators can be changed by means of a computer program contained in the blasting machine without actual replacement of the detonators. The initiation system operates on low voltage – usually less than 50 V. It is thus intrinsically safe as the danger of current leakage is low. Secondary Primary Integrated Open Voltage Lead in Explosive Explosive Circuit Protection Circuit Wires

Outer Shell

Spark Head

Capacitor

Sealing Plug

Figure 22.10 Construction of an electronic detonator.

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6 7

519

The wired round won’t fire until all detonators in the circuit are properly accounted for with respect to the current blasting plan layout. Because of the unique design and construction of electronic blasting systems, each must be used according to the manufacturer’s instructions.

Electronic initiation systems (electronic detonators) cannot be initiated by a conventional blasting unit, nor can they be activated without entering proper security codes. However, electronic detonators are still susceptible to initiation by lightning, fire, and impact of sufficient strength. Therefore, they must be properly transported, stored and handled as an explosive.

22.3

BLASTING INSTRUMENTS

Apart from the consumables discussed in the previous Section, many instruments and tools are required to be used by a blaster to connect and test the blast circuits in the mine bench. Collectively these are called accessories. Blasting instruments are of three types as under. 1 2 3

Testing Instruments Initiating Instruments Measuring Instruments Details of these are given here below.

22.3.1 Testing instruments Every non-electric, electric or electronic circuit, whichever the mode of blasting circuit may be, must be thoroughly tested before initiating the blast. Besides this, the area of blasting must also be surveyed for extraneous current, static voltages and voltages in the nearby power lines. Two very commonly used instruments for this purpose are a Blaster’s multimeter and Blaster’s ohmmeter. 22.3.1.1

Blaster’s multimeter

A Blaster’s multimeter is shown in Figure 22.11. It is used to measure voltage, resistance and current in various parts of the blasting circuit. Blaster’s multimeters have better accuracy than commonly used multimeters. Their current delivering range is also wider so they can give higher current when all the detonators in the circuit need it and give very small current when only one detonator is to be tested safely i.e. without detonating it. 22.3.1.2

Blaster’s ohmmeter

Sometimes an ohmmeter is preferred by blasters as they are usually more accurate than multimeters for measurement of resistance. Such measurements of the following aspects are of vital importance.

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Blaster’s Multimeter

Blaster’s Ohmmeter

Figure 22.11 Blaster’s multimeter and ohmmeter.

1 2 3

To determine if the bridgewire of an individual detonator is intact To ensure that the electric detonator circuit has continuity To locate broken wires and connections in a series, or series-in-parallel, circuit A Blaster’s ohmmeter is shown in Figure 22.11.

22.3.1.3

Blaster’s tagger

A tagger is a testing instrument used in conjunction with electronic blasting systems. It is shown in Figure 22.12. It is used to send a signal created through pushing its buttons, much like a calculator. The signal is sent through the electric circuit. It is picked up by a particular electronic detonator in the circuit for which it is meant. The tagger can be used to test an individual detonator, or a part of the blasting pattern, or the entire blast circuit. The tagger, together with blast box, enables initiation of the blast from a long distance. Easy-to-follow screen menus lead the blaster through all on-bench and firing operations.

22.3.2

Initiating instruments

An initiating instrument is an instrument that actually causes an action that leads to detonation of the detonator in the main explosive. In the primitive days of mine blasting it was a simple match stick. As the main explosives used for mine blasting became more and more insensitive for the sake of safety, the need arose for more sophisticated devices. An initiating device naturally depends upon the ITL used for the blasting circuit. Generally five different types of initiating devices are used. 1 2 3

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Safety Fuse Initiator Electric Detonation Initiator Detonation Wave Initiator

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521

Figure 22.12 Tagger used for testing electronic blast circuits.

4 5

Shock Wave Initiator Hercudet System Initiator Details of these are as under.

22.3.2.1

Safety fuse initiator

Instances of the use of safety fuse are very rare. The blasts designed with safety fuse are also small, and usually limited to one or two small diameter very shallow blastholes. The safety fuse, even the fast varieties, transmit fire so slowly that a blaster can use special entertainment pyrotechnic fire sticks for starting the burning of the fuse and easily move away from the blast site. 22.3.2.2

Electric detonation initiator

Initiators used for initiating the detonation of electric detonators have the capability of imparting electric current to the blast circuit. This can be done in two ways. The first is the transfer of electricity generated by a small electric generator connected to the blasting circuit. The second is the discharge of electricity stored in capacitors and given to the blasting circuit.

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22.3.2.3

Generator type blast initiator

This initiator is actually a small hand-driven generator that produces a direct current pulse to energize the detonator. In actual operation a handle in the machine is either pushed down or twisted. This action must be done very rapidly so the current generated is sufficiently high. The generator-type machines are usually rated by the number of instantaneous, or delay, caps that they will successfully fire in a straight series. Under certain conditions this type of machine may be used to detonate series-in-parallel circuits, but should never be used for straight parallel circuits. A typical generator-type initiator looks like the one shown in Figure 22.13. 22.3.2.4

Capacitor discharge type blast initiator

Capacitor Discharge, (CD) type machines have one large, or a bank of many small, capacitors that store electric energy. Charging of these capacitors can be done by using a high voltage battery or through an oscillator connected to a low voltage battery. A switch on the machine is used for discharging the charge into the circuit very rapidly – within a few milliseconds. CD-type blast initiators are small and yet powerful. They are very reliable in firing electric blast circuits. These initiators are rated in terms of voltage and energy. It is therefore essential to verify if the initiator is sufficient for the purpose. The CD initiators must always be used as per manufacturer’s recommendations. Frequent testing of these initiators from an approved tester is also necessary to ensure that the machine delivers its full output of energy. Some CD-type initiators can energize many separate circuits one after another. They have a timer circuit that ensures the energization of each blast circuit after a time lag. For this reason the unit is called a sequential timer. This facility enables the

Generator Type

Capacitor Discharge Type

Figure 22.13 Electric detonation initiators.

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blaster to give a longer delay interval than that possible with the delays inserted in the detonators or circuits. They are very useful where it is necessary to limit the amount of explosives per delay in order to control noise and vibration. Discharge type blast initiators must be used very carefully because the voltages can be in excess of 1000 V and can prove lethal. A typical discharge type initiator is shown in Figure 22.13. 22.3.2.5

Detonation wave initiator

A detonation cord transmits a detonation wave. It must be initiated at one end. For such purposes a starter gun, as shown in Figure 22.14, is used. The gun has an integral safety device and uses Shot Shell Primers No. 20 as a primer cap. It is a complete blasting machine, no other equipment being needed to initiate a Nonel tube. 22.3.2.6

Shock wave initiator

A shock wave initiator is more commonly called DynoStart, for it works with Dyno Nobel’s Nonal System. It is shown in Figure 22.15. The DynoStart blasting machine consists of an energy source, a voltage converter, a capacitor for energy storage, a voltage supervision circuit, an electrode and switches for control. A common 9 V battery is used as a source of energy. Electronic energy is converted into a strong shock wave of high temperature which is applied inside the Nonel tube, by means of the electrode, giving reliable initiation. 22.3.2.7

Hercudet system initiator

In the Hercudet system it is very essential to check the air circuit before actually initiating the detonation. For this purpose a special machine is used. It is connected to three separate cylinders of nitrogen, oxygen and gaseous fuel.

Figure 22.14 Nonal starter gun.

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Figure 22.15 Nonel shock wave initiator.

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For the purpose of testing, nitrogen gas is released into the circuit. It displaces the air and builds up pressure. If the circuit is leaky, the pressure reduces as the nitrogen escapes. After taking necessary remedial action and ensuring the circuit is leak proof, oxygen and fuel gas are mixed in appropriate proportions and are circulated through the circuit. The initiator then ignites the gas. The ignition travels through the circuit at a speed of about 2400 m/s and starts the detonation at the detonator. 22.3.2.8

Electronic blast initiator

An electronic blast initiator is used to compose a computer program-like listing that controls the complete blast. Almost every aspect of the blast such as the day, time etc. of the blast, the detonation sequence, the delay intervals, instructions to the operator at every stage of the program, automatic alarm signals etc. can be fed to the initiator. As the machine has password protection, and a very specific physical as well as computer-coded key the level of safety in blasting is highest.

22.3.3

Measuring instruments

Different types of instruments are required to be used on a blast site for measuring various parameters related to a blast. The following are a few. 1 2 3 4 5

Hot Hole Meter Burden Measuring Instrument Seismograph VOD Meter High Speed Camera Details of these are as under.

22.3.3.1

Hot hole meter

On some occasions steam flows into the blasthole from the ground. By how much the temperature of the air in the blasthole rises depends upon many factors. However, the temperature at the point of intrusion can be as high as 100°C or so. Some explosives, particularly ANFO, are very sensitive to heat and can detonate at such temperatures. A hot hole meter is a small measuring instrument as shown in Figure 22.16, which when lowered into the blasthole, keeps on measuring the temperature through its thermocouople. It gives warnings at preset temperatures so the blaster can decide in advance about the additional precautions to be taken in charging the blastholes. 22.3.3.2

Burden measuring instrument

Measurement of burden as done by the laser measuring instrument, as described in chapter 6 of this book, was rather indirect because it was computed from the observations made through the instrument. A sonic device can do direct measurement.

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Figure 22.16 Hot hole meter.

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The sonic device has two components viz. a probe and a detector. The probe is lowered in the blasthole to a desired depth and switched on to emit sonic waves. The detector is kept in contact with the bench face at the appropriate point. It receives the sound waves. From the characteristics of the sound wave received the distance between probe and detector can be calculated. 22.3.3.3

Seismograph

A seismograph system is meant to measure the properties of ground vibrations as well as the sound waves caused by a mine blast. A typical seismograph system looks like the one shown in Figure 22.17. It consists of a microphone, a geophone and main receiver and analyzer unit. The geophone is buried in the ground to a depth of about 200 mm in a particular way at the point of measurement, so the ground vibrations are received clearly. The microphone is also kept at the point of measurement with as clear a passage as possible for the sound waves. Once all the connections to the receiver unit are made, the system is switched on two to five minutes before the expected time of blast. When the blast occurs the resulting ground vibrations and sound waves are recorded. The data gathered can be subsequently analyzed to determine the safe distance for construction of houses etc., or the effects of blast on existing structures.

Figure 22.17 A seismograph.

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22.3.3.4

VOD meter

Different techniques for determining the velocity of detonation have been amply described in chapter 21 of this book. An actual instrument used for one type of VOD measurement is shown in Figure 22.18.

Figure 22.18 A VOD meter.

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22.3.3.5

529

High speed camera

A high speed camera is shown in Figure 22.19. Such cameras are capable of filming at a speed of even 1 million frames per second. For most of the applications in mining a speed of 10000 frames per second is sufficient. In other words a camera with this speed photographically records events at every ten thousandth of a second. Such photographs are of great help in many different types of measurements.

22.4

BLASTING TOOLS AND MISCELLANEOUS ITEMS

Once all blastholes are drilled, first a blaster has to check many things related to charging and then connect all the blasting components required to form a circuit. For this purpose he has to use many tools. Most of such tools have been described in Table 22.7 and some of them are shown in Figure 22.20.

Figure 22.19 High speed camera.

Table 22.7 Blasting tools and miscellaneous items. Name of the tool

Brief description

Air Horn

Air horn makes sufficiently harsh sound to draw attention of persons at long distance. This proves useful on mining bench where giving adequate warning to people is very important.

Binoculars

Binoculars prove very handy in viewing distantly placed objects.

Ceramic Knife

This knife is to be used for cutting a cartridge of explosive. The knife must be made from ceramic materials so it does not emit any spark and eliminates the chance of inadvertent detonation. (Continued)

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Table 22.7 (Continued) Name of the tool

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Brief description

Crimper

This is a plier-like tool used for firmly crimping detonating cord or shock tube into a detonator.

Depth Gauge or Water Level Indicator

This consists of a hard plastic tube bound to a flexible measuring tape in such a way that graduations on the tape indicate the depth at which the bottom of the tube lies. At the bottom of the tube an electric circuit closes when the bottom touches the top of the level of water in the blasthole. This gives alarm to the user and he can measure the depth of water level. The markings on the tape are covered by abrasion resistant transparent material.

Goggles

These prove more comfortable than safety glass in bright sunlight.

Helmet or Hard Hat

One of these two items is very essential by a blaster and particularly driller or his assistant. It protects head from falling or flying objects.

Leather Gloves

Leather gloves are used by the blaster to ensure that his hands are protected from the injuries that may arise while working in the field.

Level

This is much like a surveyor’s level but somewhat less sophisticated. It is used to measure vertical angles. Due to vernier caliper arrangements they can read angles correctly up to 0.1°.

Lighting Forecast System

Many systems, that reliably give early warning of stormy weather and lighting, are available. In most of the mining projects this system is installed at the site office.

Measuring Tape

Special abrasion resistant measuring tape made from unstretchable nylon has to be used for measuring distances like burden, spacing etc.

Plexi Mirror

This is a simple aluminum polished or mercury-backed mirror with transparent abrasion resistant coating. It is used to reflect sun rays inside the blasthole. Inner side of the blasthole becomes distinctly visible by the bright sunlight.

Portable Weather Radio

It is very important to know the weather forecast before laying the blasting circuits on a bench because rain, storm etc. can greatly affect the blasting operations. Usually portable weather radio gives weather forecast in more details than predicted through the video news channels. Before relying upon the information broadcast one must verify if the information is applicable for the mine site as it may by far away from the broadcast station.

Safety Cones

Safety cones are required by a blaster to mark the prohibited area when no one is allowed to enter as some operation related to explosive loading is in progress..These safety cones have a very wide base and are the same as those used for temporarily marking road divides.

Sound Meter

A sound meter measures the sound level of a blast by using more than one sound meters it is possible to extrapolate the sound level at a farther distance away from the origin of the blast.

Stemming Rod

This is a hollow tube of hard plastic with hard plastic caps at both the ends. It is used for tamping explosive in blasthole.

Wood Pricker

When it is necessary to make a hole in the explosive cartridge a non metallic wood pricker is used. It looks like a screw driver and has a ceramic rod on the front. When pushed inside a cartridge it makes a hole without sparking.

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Ceramic Knife

Hand Gloves

Level

Respirator

Air Horn

Welding Helmet

Figure 22.20 Some blasting tools and miscellaneous items.

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Chapter 23

Hazards to and from blasting

23.1

INTRODUCTION

The purpose of blasting in mines is not only to somehow fragment the rock mass, but to fragment the rock mass in such a way that a b

It yields breakage of maximum volume of rock mass. The fragments formed by blasting are of such size that even the largest piece formed by blasting can easily be lifted by loading equipment and carried by the hauling equipment.

Further, while achieving the above objectives the blast must not cause damage or unbearable disturbances to the surroundings. In this context the word hazard has so far been looked upon to mean the exposure to danger. This can be considered in two ways. The first is the risk to which the process of blasting is exposed to, by some external factors. The second is the risk to which the external entities are subjected to, as a result of blasting operations. This chapter deals with blasting hazards from both the meanings.

23.2

HAZARDS TO BLASTING PROCESS

Many conditions make it difficult to engineer a blast in appropriate manner and so must be considered to be hazardous to the blasting process. Some of these conditions are as under. 1 2 3 4

Presence of some ground structures such as hard rock boulders, voids etc. Heavy seepage of ground water into the blastholes Blastholes that have high temperature within them Deviation in drilling of blastholes Details of these are given below.

23.2.1

Presence of ground structure

Ground structures such as boulders, voids, beds of soft formation do affect the process of blasting and efficiency of a blast. This topic is discussed in detail in chapter 24.

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23.2.2

Seepage of ground water

The presence of ground water in a blasthole is undesirable. This is particularly so in large surface mines because in these mines the most commonly used explosive is ANFO and the AN component of the explosive dissolves in water. In such an event the blast may occur very feebly or many not occur at all. In such situations the alternatives are: 1

2 3

Dewatering the blasthole and then using a thin (usually 0.5 mm thickness) plastic tube inside the blasthole. When such a tube extends from the bottom of the blasthole to ground surface the explosives filled inside the tube do not come in contact with water. It can therefore be safely blasted. Using water resistant explosives such as slurry or emulsion. These explosives are much costlier than ANFO, resulting in increased cost of blasting. When the location of the bench is such that water seeps into many blastholes, one or more deep waterwells can be drilled at appropriate locations and ground water is pumped from such wells. This results in decreasing the ground water table in the area and the flow of water in the blasthole stops or reduces.

23.2.3

Hot blastholes

On many occasions blastholes have relatively high temperatures within them. This can result from the inflow of steam or a layer of burning coal or some similar phenomenon. Depending upon the temperature prevailing inside the hole and the temperature gradient, certain explosives can deflagrate or even detonate either immediately or after some time. For this reason it is very essential to find out if the blasthole is hot. The likelihood of the presence of a hot blasthole is very high in coal mines, where the coal layers may actually be burning. In such circumstances it is compulsory to lower the hot hole meter and determine the temperature gradient in every blasthole.

23.2.4

Hole deviation

In designing a blast it is always presumed that the blastholes will be drilled exactly at the intended collaring point and they will lie perfectly along their intended path. In actual practice the position and the alignment of the blastholes deviates from the ideal for many reasons described earlier in this book. Situations of hole deviation like this can be dealt with by using a higher quantity of explosive or using more powerful explosive in the blasthole in proportion to the hole deviation.

23.3

HAZARDS OF BLASTING PROCESS

Hazards caused by blasting process can be due to many factors as under. 1 2

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Misfires Ground vibrations

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3 4 5 6

535

Air blast Fly rocks Air pollution Environmental changes

All these are of great importance because they cannot be totally wiped out. Corrective measures to ensure that their effect is on an acceptably reduced level, are required to be taken from time to time.

23.3.1

Misfires

Misfire means missed fire. When a charged blasthole does not explode or when a part of explosive column in a blasthole has not exploded along with other blastholes in the round, a misfire is said to have occurred. If the danger of misfire is neglected and further mining activities are allowed the explosive in the blasthole can suddenly explode during such activity. This can prove disastrous to the surroundings and can lead to huge monetary loss as well as fatalities. Therefore, in every blast a blaster must take utmost care to ensure that no misfire occurs. After the blast, he must also verify that no misfire has actually occurred and if it has occurred he must take corrective measures before further activities in the mining sequence are allowed. Misfires can occur in one of the following ways. 23.3.1.1

Faulty safety fuse installation

In safety fuse blasting circuits a misfire occurs when an improperly cut safety fuse has been inserted and crimped to a detonator, or when a safety fuse is not sufficiently inserted into a detonator before crimping. These are shown in Figure 23.1. Loose crimping may separate the fuse and the detonator while they are being lowered into the blasthole. Misfire can also occur if a safety fuse that does not have sufficient water resistance, or a fuse that has a puncture within it, is used in watery condition. 23.3.1.2

Faulty electric blasting circuits

In electric blasting circuits a misfire occurs when the detonator is not properly connected to the lead wires, or when the lead wires are not properly connected to the Correctly Cut and Sufficiently Inserted Incorrectly Cut but Sufficiently Inserted Correctly Cut but Insufficiently Inserted

Figure 23.1 Faulty safety fuse installation.

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main wires on the ground, or due to use of internally damaged/short circuited wire or current leakage. Sometimes the sheathing of the lead wire gets punctured or damaged during the process of filling the stemming length of the blasthole by drill cuttings. This can lead to a misfire. If the exploder chosen for the initiation is of insufficient capacity, inadequate current reaches the detonator and becomes a cause of misfire. 23.3.1.3

Faulty detonating cord circuits

Misfires in detonating cord circuits can occur by improper joint, branch line failure, use of a detonator with a very long delay interval, or incorrect sequence of firing. Besides these, incorrect method of limiting the detonating fuse, loop cross-over, and approach of a different branch of detonating fuse can also cause a misfire. 23.3.1.4

Faults of exploder or faulty operation

More often than not, misfires are caused by a faulty exploder or its improper use. The following are the ways in which misfires are caused for this reason. Choosing an exploder of inadequate capacity Exhausted batteries Faulty indicators of the battery status Faulty cranking mechanism Defective generator. These factors are greatly reduced by proper maintenance of the exploders. 23.3.1.5

Unnoticed ground water inflow

The likelihood of sudden inflow of ground water into the blasthole is unlikely if the blasthole is in a solid intact rock mass because water always percolates through cracks and joints in the rock mass. It is always necessary to check the inflow of water into the blasthole before charging. For such observation a blasthole camera is ideal but even a plexi mirror or a beam of searchlight is sufficient. If it is noticed that there is no or negligible flow of water inside the blasthole, it is reasonable to assume that such sudden water flow will not take place after the blasthole is filled with explosive. In any case to reduce the chance of such an unexpected occurrence, blasting should be carried out as quickly as possible after the blasthole is charged. Finally, a misfire can still occur. Very careful on-site investigation about misfires must be made by the blaster about half an hour after the blast. This time must lapse because if an explosion of the explosive in the blasthole is to take place, it will take place within this time period and the blaster will not be exposed to danger. If the blaster notices the misfire there are different ways to deal with the situation as under.

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After removing the stemming material from the blasthole with utmost care a detonator can be put to a certain depth inside the blasthole. After filling the stemming material again the explosive column can be blasted. One or more blastholes can be drilled around the misfired blast. These newly drilled holes can be charged and blasted. It may perhaps be worthwhile to remove the stemming material and then circulating water inside the blasthole so as to wash out the explosive or make the explosive totally insensitive. Appropriate guidelines to deal with such situations are often given by the Governing Authorities through their publications. A blaster must proceed according the recommendations.

23.3.2

Ground vibrations

When an explosive is blasted in a blasthole, the chemical reaction evolves a huge quantity of energy. This energy starts propagating away in a radial direction. Initially the intensity of the energy is so high that matter near the walls of the blasthole gets crushed and displaced radially. When the intensity of energy decreases the energy continues to travel through the rock as an elastic ground vibration. While the ground vibration radiates out from the blasthole, the intensity continues to reduces and at long distance it becomes too low to be perceived. At nearby distances the ground vibrations have sufficiently high energy to shatter many structures firmly embedded into ground. Such shattering of the structure can cause damage to the structure which can be very extensive as well. Human beings or other things not firmly fixed to the ground also experience these vibrations but adjust and usually do not experience any permanent damage. Since it is essential to reduce the likely damage to structures by blasting, it is essential to study the nature of these waves, how they propagate and decay in their intensity, and what structures are damaged at what intensity. 23.3.2.1

Nature of ground vibration waves

There are three distinct types of vibration waves. They are compression wave, shear wave and Rayleigh wave. They are usually recognized by their shortened name, P wave, S wave and R wave respectively. When the waves travel radially the particles of the ground matter move in different ways. The P wave is a compression wave. Here the particles of the matter move back and forth in the radial direction. Before the wave generates, the particles of the matter are in steady state. As the wave begins, the particles swiftly move in a radial direction and push the particles on the front side in a radial direction. This gives rise to compression. Immediately upon passing the energy to the next particles, the particles rebound back again in the radial direction and create a tension wave. This is elucidated through Figure 23.2. In the shear wave the movement of the particles is at right angles to the direction of wave propagation, as shown in Figure 23.3.

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First compression wave has just formed from particles at rest Direction of Wave Propagation Compression wave has moved and tension wave has formed Direction of Particle Movement Cycle of Compression + tension waves is moving

Figure 23.2 P wave propagation. Upward shear movement has just taken place from particles at rest Upward shear movement has moved and downward shear movement has taken place.

Direction of Wave Propagation

Direction of Particle Movement

Figure 23.3 S wave propagation.

In the Rayleigh wave the direction of wave propagation is radial but the movement of the particles is in an elliptical manner in a plane that is perpendicular to the direction of wave propagation. The P wave is robust and travels over a very long distance at a speed that is dependent upon the medium in which it travels. Vibrations are measured in terms of amplitude and frequency. Amplitude relates to the movement of particles in either direction from the stable position. It is measured in mm. Frequency is the number of repetitions of the movement pattern. It is measured in number of cycles per second. Excessively high ground vibration levels are known to damage structures. Moderate to low levels of ground vibration are irritating. Seismographs are used to record the ground vibration waves experienced at a place. The recordings are called a wavetrace. Figure 23.4 shows the wavetrace at a point approximately 1800 meters from a single hole blast of 1000 kg of explosives fired in the overburden in the Hunter Valley coal mine. The P wave arrives first followed by the various reflected and refracted vibrations associated with the P wave. This vibration gradually reduces until approximately 800 ms later, when the S wave arrives with its associated reflected and refracted waves. Approximately 900 ms later, the R wave arrives with its associated reflections and refractions. The vibration gradually decreases and returns to zero sometime after the 4 second time window. Actually the waveform recorded by the seismograph is the combination of many waveforms each having a different frequency. Studies have shown that high frequency wave energy is absorbed more readily than low-frequency wave energy so that the energy content of stress waves at large distances is concentrated at low frequencies. The velocity is referred to as particle velocity in order to distinguish this quantity from

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Figure 23.4 Ground vibration wavetrace of a blast in single blasthole.

propagation velocity. The peak particle velocity of ground vibration depends on the maximum weight of the charge per delay of eight milliseconds or more and not on the total charge weight of the blast. The most significant ground motion parameter is the maximum radial particle velocity, commonly referred to as peak particle velocity which is usually the maximum of the three components. The amplitude of ground vibrations arising out of a mine blast increases with the energy released by the explosion, the confinement of explosive in the ground and to a small extent the density of rock. Ground movement induced by blasting has some important characteristics. Usually the frequencies of vibration are in excess of 100 Hz in the close surroundings of the blast area. The vibrational frequencies and the maximum particle velocity decrease with increasing distance. If the rock mass is very hard more energy is absorbed in the vicinity of the blast. 23.3.2.2

Prediction of ground vibration levels

Research on accurate prediction of ground vibration levels through mathematical formulae has been carried out by many researchers. Some of the better known formulae are given here below. 23.3.2.2.1

Langefors formula

The formula proposed by Langefors is as under: v = k * (Q/D1.5)0.5

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where v = Peak particle velocity in mm/s K = Rock transmission factor Q = Instantaneous charge mass in kg D = Distance from blast point in m Langefors and Kihlstrom in their research in Sweden found that for the Swedish granite rock in which the research was carried out the rock transmission factor K was 400. Specific attention must be given to the value Q i.e. instantaneous charge. Instantaneous charge is the charge that has blasted within a time period of 8 milliseconds. For example, in a blast of 96 blastholes each containing 300 kg of explosive, if the timings of the delay interval were as given in Table 23.1, then the value of q will be 26 * 300 = 7800 because within a time span of 8 milliseconds (from firing time 30 to 35) 7 + 19 = 26 blastholes have been fired. 23.3.2.2.2

Scaled distance formula

For prediction of peak particle velocity, the scaled distance formulae are used more often than the Langefors formula. The general form of these formulae is as follows. v = k * (D/Q(1/x))-e where v = Peak particle velocity in mm/s k = Constant related to site conditions Q = Instantaneous charge mass in kg D = Distance from blast point in m e = Exponent related to site conditions In the above formula the term (D/Q(1/x)) is called the scaled distance. Table 23.1 Timings for a mine blast.

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Firing time in milliseconds

No of blastholes fired

0 25 30 35 75 100 125

5 12 7 19 20 18 15

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541

For x = 2 the scaled distance equals (D/√Q) and is called the square root scaled distance. This value was recognized through the investigations carried under USBM. This value is also adopted by US and Australian Standards. For x = 3 the scaled distance equals (D/3√Q) and is called the cube root scaled distance. This value was recognized through the research carried out by Ambraseys, Hendron and Oriand. The square root scaled distance is most commonly used and is based on the observation that the charge is distributed in a long cylinder (the blasthole), therefore the diameter of the hole is proportional to the square root of the charge weight. However, it can be argued that as the hole length shortens in relation to the diameter, the charge mass approaches a spherical shape, in which case the diameter is proportional to the cube root of the charge weight. For estimation of the vibration level the value of k is taken as 1140 and the value of e is taken as = 1.6. With these values the peak particles can be easily found, for different instantaneous charge masses and different distances, from the chart in Figure 23.5. The site condition constant will differ depending upon the conditions. Values of this constant are: for highly structured or hard rock k = 500, for average conditions k = 1140, for heavily confined blasting near field k = 5000. The values of exponent e for different rock masses are as given in Table 23.2. 23.3.2.3

Damage by ground vibrations

Ground vibrations affect every living and nonliving entity. Damage directly caused to human beings or other living entities by ground vibrations is virtually unheard of. This is because they are very loosely connected to the ground mass. More importantly they have inherent capability of responding to different vibrational frequencies by internal adjustments. However, human response to ground vibrations can be perceptible, unpleasant or intolerable depending upon the peak particle velocity and the frequency of vibrations. The chart in Figure 23.6 gives a fair idea of the human response to peak particle velocities and frequencies of ground vibrations. Structures that are firmly fixed in the ground have only one natural frequency, which depends upon their structural properties. Natural frequencies of different types of structures are given in Table 23.3. Damage to structures occurs due to differential particle movement as shown by Figure 23.7. It can be divided into three categories as stated in Table 23.4. The significance of natural frequencies lies in the fact that when the ground vibration frequencies match with the natural frequencies of the structures, a phenomenon called resonance takes place. In such circumstances the amplitude of the vibrations keeps on increasing and the structure absorbs a higher quantum of energy. Eventually the stresses at certain points in the structure exceed the strength limit and the structure fails.

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10000

3000

Distance from Blast Centre m

1000

Peak Particle Velocity in mm/s

0.1 300

0.2 0.5 1

100

2 5 30 10 20 10

50 100 200

3 0.3

1

3

10

30 100 Charge Weight Per Delay kg

300

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Figure 23.5 Chart for estimation of ground vibration level.

1000

3000

10000


543

10 6

Peak Particle Velocity inches per Second

4 2 1 Intolerable 0.6 0.4 Unpleasant 0.2 0.1 0.06

Perceptible Level

0.04 0.02 0.01

1

2

4

6 10 20 Frequency CPS

40

60

100

Figure 23.6 Response of humans to levels of ground vibration. Table 23.2 Values of exponent e for different rock masses. Rock mass type

Value of exponent e

Rhyolite Granite Limestone Ordovician Sediments Overburden in Coal Mines Massive Basalt

2.2–2.5 2.1–2.4 2.1 2.8 1.5–1.8 1.9–3.0

Table 23.3 Natural frequencies of some typical structures. Structure or element

Natural frequencies (NF) in Hz

Multistory Building

NF = 0.1N (N = Number of Stories) 3.8 1.2 0.6 12–20 7.0 (Standard Deviation 2.2)

Radio Towers 100 ft Tall Petroleum Distillation Towers 65 ft Tall Coal Silo, 200 ft Tall Building Walls Wood Frame Residences

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Failure by P Wave i.e. Compression Wave

Failure by S Wave With Movement in Vertical Direction

Failure by S Wave With Movement in Horizontal Direction

Figure 23.7 Effect of different types of ground vibration waves on structures.

Table 23.4 Three categories of damage to structures by ground vibrations. Category

Description of damage

Threshold

Formation of new minor cracks in plaster or at joints in wallboard, opening of old cracks and dislodging loose objects.

Minor

Superficial, not affecting the strength of the structure; for example, loosened or fallen plaster, broken windows, significant cracks in plaster, hairline cracks in masonry.

Major

A significant weakening of the structure, large cracks, shifts of the foundation, permanent movement of bearing walls, settlements which cause distortion of the structure or walls out of plumb.

On the basis of research carried out, mainly in the USA, the maximum limits set for ground vibrations are as shown in Figure 23.8. A maximum limit of peak particle velocity of 2 in/sec. was suggested by the U.S. Army Corps of Engineers, whereas a maximum limit of 1 in/sec. was suggested by USBM in 1977. Both these values were independent of the frequencies.

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Peak Particle Velocity inches per Seconds

10 6 4

Limits as Suggested by U.S. Army Corps of Engineers

2 Limits as Suggested by USBM in 1977 1

7.5

0.6 0.4

11 Limits as Suggested by USBM in1983

0.2 0.1

1

2

4

6 10 20 Frequency Hz

40

60

100

Figure 23.8 Limits of ground vibration levels to avoid damage to structures.

The latest limiting values as applicable from 1983 are based on more rational research by USBM and are dependent on ground vibration frequencies. This need arose because the tall structures are very vulnerable to ground vibrations with frequencies less than 10 Hz. This is quite evident from the values of natural frequencies given in Table 23.3. It is usually possible to design the blasts by adjusting the mass of instantaneous charge in such a way that the peak particle velocities lie below these limits. More detailed description can be found in many research papers and books.

23.3.3

Air blast

Air blast is often called air overpressure because the rapid movement of rock mass, and the ejection of gases under extremely high pressure, cause pressure waves by successively increasing and decreasing the pressure of atmospheric air. The pressure waveform recorded from the blast of a single blasthole is shown in Figure 23.9. Air pressure waves are essentially compression waves. Like ground vibration waves they also have amplitude and frequency. Higher amplitude means louder sound. High and low frequency sound waves form sharp or coarse sounds. Human beings are not able to hear sounds of all frequencies. A few individuals may have the ability of hearing sounds of frequencies as low as 20 Hz and as high as 20 kHz, but for most persons the frequency of audible sound is limited to 60 Hz to 15 kHz. The absolute air pressure that can be barely heard by a human with sharp ears is measured to be about 0.00002 i.e. 2 * 10−5 Pa. Similarly the air pressure at the point of blast has been estimated at 20000 Pa. Thus the range of pressure in the spectrum

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Figure 23.9 Air vibration wavetrace of a blast in single blasthole.

is very high. Therefore, air pressures are most usually described in terms of decibel, shortened to dB. The relation between actual pressures and dB is dB = 20 * log10 (P/P0) where dB = Sound level value P = Pressure as measured in Pa P0 = Reference pressure in Pa 23.3.3.1

Prediction of air blast pressure

The formula used for calculation of pressure of an air vibration wave is P = A * (D/W0.33)a where P = Peak air pressure in kPa A = First site constant a = Second site constant D = Distance from blast in m W = Instantaneous charge mass in kg As per ICI “Handbook of Blasting Tables”, airblast overpressure for an unconfined charge may be estimated through the following equation: P = 185 * (D/W0.33)-1.2 Extensive tests carried out by Terrock Pty Ltd. at a military firing range confirmed the above equation with slight variation as under.

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P = 182.5 * (D/W0.33)-1.2 For fully confined blasthole charges, the formula set forth by ICI is P = 3.3 * (D/W0.33)-1.2 However, as per tests carried out by Terrock this changes to P = 12.7 * (D/W0.33)-1.2 The above formulae can be used for the purpose of estimation of airblast pressures. Air pressure of a blast diminishes as the pressure wave travels away from the blasthole. Tests carried out by Terrock revealed that when the distance from the blasthole was doubled, a pressure drop of about 7 to 10 dB was experienced. The most common value was 8.6 dB. On the basis of data collected, the graph as shown in Figure 23.10 was constructed. It is to be noted that air blast pressure waves travel from the blastholes towards the bench face and then away in the same direction without much pressure drop. However, the pressure drop towards the rear side is much higher. 23.3.3.2

Damage by air blast pressure

For quick apprehension of the levels of sound i.e. air blast pressure waves, Table 23.5 mentions different sound levels and the corresponding sound commonly heard. Usually below sound pressure level of 90 dB no hearing loss occurs even if it is heard for 8 hrs per day. Above this level hearing loss occurs. The hearing loss is greater when sound pressure is higher and/or duration of hearing is longer. When sound pressure level is about 100 dB, to avoid significant hearing loss the maximum duration should not exceed 2 hrs in a day.

Figure 23.10 Reduction in air blast pressure with distance.

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Table 23.5 Air blast pressure wave examples. Pressure level dB

Value of exponent e

180 176 171 155 151 133 130 110 100 90 60 40 25 5 0

Most Structures Severely Damaged Plaster of Walls Crack Most Windows Break Space Rocket Some Windows May Break Most Commonly Adopted Maximum Allowable Value Jet Airliner Large Orchestra with Electric Amplification Pneumatic Rock Drill without Silencer Loud Noise by Human Conversational speech Normal Conversation Whisper Rustling Leaves Threshold of Hearing

When sound pressure level is about 115 dB to avoid significant hearing loss the maximum duration should not exceed 15 minutes in a day. Ear plugs or ear muffs are required to be worn to reduce hearing loss in conditions of harsh and loud sound. These devices have been found to be useful in reducing sound levels entering the ear by about 20 dB. Sound pressure level of 133 dB is the maximum allowed even for a very short period of time. Beyond this level the air blast pressure wave causes pain to the ears. Injury may occur to the ears even with protection. Damage to structures starts at about 133 dB and increases as indicated in Table 23.6. There are many means by which damage from air blast can be minimized. Some of them are as under. 1 Ensure that all blastholes are adequately stemmed with appropriate stemming material. To prevent blowouts from blastholes a stemming length of 30 times the diameter of the blasthole should be used 2 Ensure that the front row has adequate burden. Adjust the charge according to the available burden 3 Adopt appropriate timing delays and sequence 4 Keep the rate of detonation across the face less than the speed of sound 5 Ensure that all explosive materials like detonating cords etc. are buried at least one foot in the ground 6 Do not carelessly leave any unused pieces of detonating cord or other explosive items on the bench top 7 Do not use more explosive in a delay than permitted 8 When possible, delay the blast to direct movement away from critical areas

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Table 23.6 Damaging effect of air blast pressure waves. Pressure level dB

Value of exponent e

<153 165 166–173 <173

Severe shaking of windows and doors but no damage One out of twenty properly mounted window glass may just break Most of the window glass break depending upon the pressure All glass will break and additionally intact plaster of the wall may also break

9 Avoid adverse environmental conditions such as blasting when wind is blowing toward residential areas, or under an atmospheric temperature inversion 10 Do not blast more often than necessary 11 Time blasts to coincide with peak ambient noise levels 12 Give adequate warning messages about a blast in the offing to all those who are likely to be within 3 to 5 km distance from the blast bench. Use a warning horn, loudspeaker messages etc.

23.3.4

Fly rocks

In a mine blast the build up of excessive pressure within the blasthole exerts a propelling force on the pieces of rock and makes them fly like a projectile. These are called flyrocks. The throw of flyrocks is very unpredictable as far as their precise direction is concerned. However, it can be said that flyrocks usually move from the blastholes in the first row towards the free face as shown in Figure 23.11. Flyrock can be attributed to many reasons. Some of them are as under. 1 2 3 4 5

Inadequate burden Use of high quantity of explosives charge Insufficient stemming height Improper firing sequence Use of heavily loaded snake holes.

Flyrock studies have been conducted only recently on a large scale when high speed cameras became easily available. Their use in the field enabled visual tracking of the trajectories of the flyrocks. It has been observed that when the specific charge in the blasthole is less than about 0.2 kg/m3, there is hardly any throw of flyrocks. As the specific charge becomes higher, the distance to which flyrocks fly becomes greater. An empirical relation between different associated variables is L = 5.63 * d * (q – 0.2) where L = Maximum throw of flyrocks in m q = Specific charge of the blast in kg/m3 d = Diameter of blasthole in mm

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Large pieces of rock ejected with low velocity

Small pieces of rock ejected with high velocity

Figure 23.11 Flyrocks at the position of reduced burden.

Thus for a typical value of 0.5 kg/m3, the equation becomes L = 1.689 * d From this a flyrock throw distance of 525 m can be estimated for a 311 mm dia. blasthole. Two more empirical equations have been set forth after analyzing the problem with the help of computer simulations. They are as under. L = 260 * d0.66 φ = 0.1 * d0.66 where L = Maximum throw or flyrocks in m d = Diameter of blasthole in inches φ = Diameter of boulder in m From this equation for a 10 inch diameter blasthole, the throw distance of 1188 m can be calculated for a boulder of diameter 0.47 m. More research is needed in respect of flyrocks so as to formulate safe limits. Till then blasters have to rely on the factors stated above and more importantly their experience in blasting.

23.3.5

Air pollution

Air pollution is probably the most devastating hazard of mining. Unfortunately it is never considered so because effects of air pollution are not immediate like air blast, fly rocks or ground vibrations.

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Following are some of the happenings during or after the blast. The formation of toxic fumes in chemical reactions that take place during blasting was described in chapter 18 of this book. The noxious gases formed in the process, viz. carbon monoxide, nitrogen oxide and nitrogen dioxide, mix with air. Their concentration quickly reduces below the danger levels. Sometimes they remain entrapped below the heap of broken rock and can make their presence felt even after some days but they too eventually mix with atmospheric air and their concentration reduces below danger level. If someone has to be in the midst of a blasted bench immediately after the blast he must wear an oxygen mask and special suit. In a mine blast more serious pollution of the atmospheric air is caused not by noxious gases but by other factors. The first of these factors is the dust. Rational guesswork in this regard indicates that at least 500 g of dust is being mixed with the atmospheric air for each m3 of blasted material in a mine blast. Silicates being the most abundant minerals in the earth’s crust, a very large part of the dust is siliceous. Millions of microscopic silica particles go into the lungs with every breath. As they do not have any role in the body they are not absorbed into the blood. They remain glued to the alveoli through mucus and are not excreted with exhalation. This way silica particles accumulate in the lungs. Over time scar tissue develops in the lungs, which damages the lungs’ ability to work properly. The inhalation of silica particles has also been linked with lung cancer, bronchitis and tuberculosis. The silicosis itself may lead to other conditions including lung fibrosis and emphysema. The disease is also linked to a fatal pulmonary tuberculosis actually called silico-tuberculosis. As the disease progresses, the person experiences increasing difficulty in breathing and may die. Workers working near the points of pollution are at greater risk. The medical community considers that these diseases are incurable and irreversible and may progress even after the exposure ends. This is not entirely true. Silicosis kills thousands of people around the world every year. Variants of silicosis are only a part of the story. Apart from silica, every blast mixes several million metallic free radicals with atmospheric air. These metallic free radicals are mainly of lead, mercury, aluminum and copper. Depending upon the mine, nickel, beryllium, uranium, cadmium, iron etc. are also mixed in atmospheric air. Molecular or atomic size particles of these elements are very active because they are deficient in their electronic balance. They lack one electron and hence are termed free radicals. When they enter into the human body, they try to grab electrons from the atoms in their contact. This proves very harmful to human body as it accelerates ageing. Persons exposed to such an environment succumb to several degenerative diseases in their forties or fifties rather than eighties. High blood pressure, coronary artery disease, Alzheimer’s disease and Wilson’s disease are some examples of such diseases. These diseases are also considered to be incurable, irreversible and progressive. This is also not entirely true. If not curable, these diseases are reversible to a very great extent with appropriate medical treatments. I have studied these aspects and later given treatments to some patients and verified this.

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In any case the treatments for these diseases are very long drawn if not very expensive.

23.3.6

Environmental changes

Hazards of drilling and particularly blasting have very long term effects on the surroundings. They not only affect the human population in the vicinity but also affect all types of animals and plants. When a mining project is to be started, a large area is acquired by the authorities. It runs into several thousand hectares. Such a large area is required for the actual mining site and also the processing plants, township of mine workers etc. The area is often a forest land where trees and shrubs have grown for centuries. Several generations of animals have spent their entire lifespan there. Birds have sang morning and evening songs. Reptiles have hissed. As the mining project progresses greenery reduces when trees and shrubs are cut. Birds fly away as they have no place to make their dwelling. Animals run away when their food resources have dried up. Reptiles come out of their underground holes and are often killed on the makeshift roads. In a typical mine scenario, the work site is almost invariably situated far away from the locality. Even the workshop is located a few kilometers away. As the mine starts, the burning sun and movement of heavy equipment on the dry ground make the atmosphere so dusty that the color of the surroundings gets rapidly converted into the color of dust layers deposited on such surroundings. Harsh rattling noise of the machines often becomes unbearable. Visitors unaware of such surroundings get frustrated. Their frustration turns into annoyance when they realize that even for a glass of water they have to tread a kilometer of uneven dusty road. In the last few decades the grave consequences of such changes in scenario have been noticed, and now every mining project, whether government-owned or private, has to take necessary steps towards preserving or restoring the environment.

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Chapter 24

Properties of rock masses

24.1

INTRODUCTION

As has been said earlier, the main objective behind blasting a rock mass is to fragment it in such a way that the fragments created in the process are small enough to be easily removed from the blasting site. To achieve this objective the following necessities arise. 1 2

3

Choose proper blasthole parameters viz. diameter, depth and inclination. Determine such a layout for the blasthole positions, that after appropriately charging and blasting them the fragments of rock mass are easily loaded and hauled by the equipment chosen for the purpose. Choose apt explosive, efficient method and pattern of charging the blastholes, the sequence and timing of detonating explosives and appropriate accessories for the blast.

As stated earlier, blasting is a very hazardous and dangerous operation. It can become the cause of hundreds of deaths if attention is not given towards minimizing the hazards. Obviously, all the measures to reduce such hazards have become a part of every blasting program. Several properties of rock specimen were described in chapter 3 of this book. The background behind various tests carried out to determine the magnitude of the properties and the methodology of the test was also elaborated. The reason to take such account early on in this book was that the properties had great influence on the process of rock breakage in drilling the blastholes. In the first part of this chapter the influence of rock specimen properties on blasting is considered. When it comes to blasting, the properties of the rock mass are of greater importance than rock specimen properties. This is due to the fact that in drilling, only a small portion in the alignment of the blasthole is to be fragmented into very small pieces, whereas in blasting a very large mass of rock is to be fragmented into relatively large pieces. How the properties of rock mass can be taken into consideration while designing a blast has been elaborated in the second part of this chapter.

24.2

ROCK SPECIMEN PROPERTIES AND BLASTING

Rock specimen properties that affect blasting are strength, density and porosity.

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24.2.1

Influence of rock strength on blasting

For long it was an observed fact that rocks with higher compressive strength need more explosive energy for fragmentation. It was later noticed that actually the tensile wave reflected from the free surface influences rock breakage to a greater extent. In the year 1959 Hino postulated that the tensile rock fracture occurs in the form of slabs that are parallel to the free surface. The extent of tensile fractures and the number of slabs so produced depends on the tensile strength of rock (σt), amplitude (σa) and length (L) of the compressive wave. He concluded that the number of slabs (n) produced by tensile slabbing due to reflected shock waves may be given by n ≤ σa/σt or n ≤ L/2t where t = thickness of slab. Hino also noticed a linear relationship between the compressive strength of rock (σc) and the amplitude of the compressive stress wave (σa) propagated through the rock. Therefore, σa ∝ σc and hence, n ∝ σc/σt. Since σc/σt is proportional to the blasted rock, he proposed it to be called a blasting coefficient. In the year 1979, Kutuzov correlated the powder factor with compressive strength of rock and gave the tabulated form, as in Table 24.1.

24.2.2

Influence of rock density on blasting

Normally the density of rock is well correlated with its compressive strength. High density rocks require more energy for their deformation and fracture. Explosives that release a higher volume of gases exert higher pressure on the blasthole walls, and the consequent higher bubble energy is capable of fracturing a larger rock mass.

Table 24.1 Rock mass classification on the basis of joint spacing and bed thickness. Powder factor in kg/m3

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Range

Average value

Mean distance between natural fractures in rock in m

0.12–0.18 0.18–0.27 0.27–0.38 0.38–0.52 0.52–0.68 0.68–0.88 0.88–1.10 1.10–1.37 1.37–1.68 1.68–2.03

0.150 0.225 0.320 0.450 0.600 0.780 0.990 1.235 1.525 1.855

<0.10 0.10–0.25 0.20–0.50 0.45–0.75 0.70–1.00 0.95–1.25 1.20–1.50 1.45–1.70 1.65–1.90 >1.85

Uniaxial compressive Strength MPa

Density of rock kg/m3

10–30 20–45 30–65 50–90 70–120 110–160 145–205 195–250 235–300 >285

1400–1800 1750–2350 2250–2550 2500–2800 2750–2900 2850–3000 2950–3200 3150–3400 3350–3600 >3550

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Blasthole pressure is also dependent upon the velocity of detonation of the explosive, and the velocity of detonation is in turn dependent upon the blasthole diameter. For all these reasons it is prudent to use blastholes of larger diameter and use an explosive with higher bubble energy while blasting a denser rock mass.

24.2.3

Influence of rock porosity on blasting

Porosity of rocks is the spread of small inter-particle pores that are created while the rocks are being formed, either by plutonic or sedimentation activities. These pores are spread throughout the rock mass. In many cases such rocks do not require blasting. They can be fragmented by ripping or other methods. However, when blasting has to be carried out, explosives with higher bubble energy and low shock energy are found to be more suitable for the fragmentation. Increasing bubble energy and reducing shock energy can also be accomplished by decoupling the charge and initiation system. More than normal stemming is also a method through which pressure built up in the blasthole can be increased.

24.2.4

Specimen blastability

This blastability index has been proposed by the Norwegian University of Science and Technology (NTNU). It takes into account many variables such as sonic velocity, anisotropy, density of rock, charge density etc., involved in the process of blasting. The NTNU Equation for blastability index is as under. S = (0.7364 * Ia 0.61 * (T)0.72)/((C/1000)0.4 * (W/C)0.25 * ρ0.19) where S = Rock blastability index Ia = Anisotropy index = Cy/Cz Cy = Sonic velocity of dry rock parallel to the foliation in m/s Cz = Sonic velocity of dry rock normal to the foliation in m/s C = (Cy + Cz)/2 ρ = Dry density of rock in kg/L T = Charging density of explosive in kg/L W = Detonation velocity of explosive m/s. Good, medium and poor blastability is indicated by index values of 0.38, 0.47 and 0.56 respectively.

24.3

PROPERTIES OF ROCK MASSES AND BLASTING

In a large area to be excavated, the rock mass is hardly ever homogeneous. Different portions of rock mass have varying mineral contents.

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Rock masses also have inconsistencies like voids, folds, unconformities, bedding planes, faults and joints. All these defects result from volcanic, plutonic and tectonic activities, and other processes at the surface of the earth, through which rocks are formed.

24.3.1 Voids Two types of voids can be found in rock masses. In some types of rocks, particularly in the softer varieties, very small size voids are homogeneously spread over a very large area. These voids make the rock mass very weak – so weak that there is no necessity of blasting. However, if such a rock mass contains a layer of hard rock or some very large boulders, blasting may be unavoidable. Sometimes voids of large volume are formed in a rock mass during volcanic or tectonic activities, or by erosion. In such circumstances the gases formed in the explosion rush into the void as shown in Figure 24.1 A. In the process these gases absorb quite a bit of energy. Thus, the energy left for fragmentation of rock mass is reduced and large pieces of rocks are formed. Similarly, if an easily blastable ground mass contains a large boulder of significantly hard rock as shown in Figure 24.1B, during the fragmentation the rock mass breaks but the boulder does not get sufficient energy to break. This results in the need for secondary blasting.

24.3.2

Folds, unconformities and bedding planes

Sedimentary rock masses cover a considerable part of the earth’s surface. When formed, they are in layers lying one upon another. These layers are called beds or strata. The thickness of these beds varies from a few millimeters to several meters. The bounding planes of a bed are called bedding planes. During the deposition of the sediments that give rise to the beds, often there is a time interval during which no deposition takes place. Therefore, a surface formed during this time interval separates the old beds from new beds. On many occasions, due to a change in the mode of deposition or a change in the type of sediment, such a surface is very distinct and is called an unconformity. The properties of the rock mass on two sides of an unconformity can differ considerably. A

B

Figure 24.1 Presence of voids and boulders.

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Horizontal beds of sedimentary rocks are often distorted by physical forces exerted on them. Such geological activities are termed tectonic activities. Tectonic i.e. structural activities are mainly caused by plutonic activities taking place in the earth’s crust and mantle. In tectonic activities beds are often compressed and distorted in such a way that they take the shape of a waveform. Such structures are called folds. Strike and dip are two terms used to indicate the direction and magnitude of inclination of the bedding planes of the sedimentary layers. Whatever the inclination of a bedding plane may be, it is always possible to draw a straight line on each of the bedding planes in such a way that the line will be horizontal. The direction of such a line is called the strike. The strike can be defined by noting its bearing with north or any other well-defined direction. If the bedding plane is really a plane and not a curved surface, all the lines parallel to the strike and lying on that bedding plane will also be horizontal. A line lying in the direction perpendicular to the strike and also lying on that bedding plane, will have maximum inclination with the horizontal. The angle of inclination of this line with the horizontal is called the dip. The direction of such a line is called the direction of dip. Any line on the bedding plane, but not lying in the direction of dip will have lesser inclination than the dip. It is, therefore, termed as apparent dip. The strike and dip of a bedding plane are shown in Figure 24.2. The orientation of bedding planes with respect to the blasthole alignment can have considerable influence on the outcome of blasting because the bedding planes are weak, and parting of the rock mass along these bedding planes is rather easy. Based on the relationships between the orientations of bedding planes and blastholes, three cases, as under, are usually considered. 1 2 3

Shooting with the dip Shooting against the dip Shooting along the strike An elaboration of these is given below.

Sedimentary Layers

Bedding Plane

Direction of Strike

90°

Direction of Dip

Figure 24.2 Strike and dip of a bedding plane.

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24.3.2.1

Shooting with the dip

In this situation, shown in Figure 24.3, the lines of intersections of the bedding planes and bench floor or bench top are parallel to the bench crest. For the following reasons, the blast results are neat. Toe problems are less hence the resultant bench floor is smooth. As planes of weakness parallel to the bench face already exist, the failure by flexure is easier. For this reason the fragmentation tends to be more satisfactory and the throw of the fragments is farther away from the bench face. Such a muckpile is easy for loading operations. However, a blast in such circumstances tends to give backbreak problems. The inclination of blastholes in the direction of dip (as shown in Figure 24.3) is one of the remedies in reducing backbreak problems. 24.3.2.2

Shooting against the dip

In this situation also the lines of intersections of the bedding planes and bench floor or bench top are parallel to the bench crest, as shown in Figure 24.4.

Figure 24.3 Shooting with the dip.

Figure 24.4 Shooting against the dip.

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The results of blasting can have one or more of these undesirable features. The booster charge at the bottom forms cracks along the bedding planes and the magnitude of breakage of rock on the two sides of such planes is often different. For this reason there can be many ups and downs in the newly formed bench floor after blasting. The toe can have large size stumps. There are more oversize rock pieces. The backbreak may be less but there are more cracks on the top near the crest of the newly formed bench. Large overhangs increase the possibility of a rock fall from the high wall. The blasted fragments are not thrown farther away from the high wall. This results in a tall heap of broken material. The following measures may reduce the problems to some extent. Extended subdrilling length Higher booster charge Use of explosive with higher brisance at bottom Decking of explosives in a blasthole Use of small diameter satellite holes near the crest 24.3.2.3

Shooting along the strike

In this situation the lines of intersection of the bedding planes and bench floor or bench top are perpendicular to the bench crest, as illustrated in Figure 24.5. From the viewpoint of blasting this is the worst situation for the following reasons. Since different types of rocks outcrop on the floor of the newly formed bench floor, due to the differences in their properties the bench floor can be highly uneven. Backbreak resulting from the blast can be very irregular. Field experience has shown that this situation is the worst amongst the three. Reorienting the bench face, and blasthole inclination, may be the only viable solutions.

Figure 24.5 Shooting along the strike.

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24.3.3

Faults and joints

Both faults and joints are fractures in the ground mass. Their size is very large. They are essentially planes of separation formed during plutonic or tectonic activities. When there has been observable movement of the rock mass on the two sides of the fracture plane, they are termed as faults, otherwise they are called joints. Joints are usually small in thickness and may have intrusion of fine clayey particles. Faults are usually much thicker. During formation of the fault, when two sides of the rock mass rub against each other, many pieces of rock are formed and remain inside the fault zone. Later the intrusion of fine particles into the fault zone takes place and rocks like breccia or conglomerate are formed. As far as blasting in large surface mines is concerned there is no distinction between faults and joints. These fracture surfaces can be found in almost any inclination and direction. The frequency of their occurrence is often very high. Fracture surfaces i.e. joints, require to be given great attention, not only in the realm of blasting but any other type of excavation and construction in or above the ground. For this reason very much research has been done in respect of the effects of joints in the rock mass during blasting or excavation. The next sections of this chapter have been devoted to some details of the effects of joints on blasting.

24.4

CLASSIFICATION OF ROCK MASSES

Rock masses are classified either through visual observations or indexes proposed on the basis of diverse parameters of the rock mass.

24.4.1

Classification by visual observation

Terzaghi proposed a rock mass classification in the context of tunnel support design in 1946. The concepts behind the classes are very important. Classes and their description proposed by Terzaghi are given in Table 24.2. In 1963, Deere proposed a classification of rock mass based on the spacing between joints found in the cores obtained in exploratory diamond core drilling and the thickness of the beds encountered in the rock mass. This conceptual description is presented in Table 24.3. Both these classifications are now superseded by other classification schemes that are more specific. They are based on absolute factors and do not leave much scope for different interpretations. Many such schemes take into account many more factors than UCS of rock or joint frequency.

24.4.2

Classification by index

Some systems of rock mass classification that are based on statistical or empirical indices include:

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Table 24.2 Rock mass classification based on field and core sample observations. Rock mass

Description of the rock mass

Intact Rock

It contains neither joints nor hair cracks. Hence, if it breaks, it breaks across sound rock. On account of the injury to the rock due to blasting, spalls may drop off the roof several hours or days after blasting. This is known as spalling condition. Hard, intact rock may also be encountered in the popping condition involving the spontaneous and violent detachment of rock slabs from the side of roof.

Stratified Rock

It consists of individual strata with little or no resistance against separation along the boundaries between the strata. The strata may or may not be weakened by transverse joints. In such rocks the spalling condition is quite common.

Moderately Jointed Rock

It contains joints and hair cracks, but the blocks between joints are locally grown together or so intimately interlocked that vertical walls do not require lateral support. In rocks of this type, both spalling and propping conditions may be encountered.

Blocky and Seamy Rock

It consists of chemically intact or almost intact rock fragments which are entirely separated from each other and imperfectly interlocked. In such rock, vertical walls may require lateral support.

Crushed but Chemically Intact Rock

It has a character of crusher run. If most or all of the fragments are as small as fine sand grains and no re-cementation has taken place, crushed rock below the water table exhibits the properties of water-bearing sand.

Squeezing Rock

This type of rock slowly advances into the tunnel without perceptible volume increase. A prerequisite for squeeze is a high percentage of microscopic and sub- microscopic particles of micaceous minerals or clay minerals with a low swelling capacity.

Swelling Rock

This type of rock advances into the tunnel chiefly on account of expansion. The capacity to swell seems to be limited to those rocks that contain clay minerals such as montmorillonite, with a high swelling capacity.

Table 24.3 Rock mass classification on the basis of joint spacing and bed thickness.

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Joint spacing

Bed thickness

Spacing in inches

Very Close Close Moderately Close Wide Very Wide

Very Thin Thin Medium Thick Very Thick

Less than 2 in. 2 in. to 1 ft. 1 ft. to 3 ft. 3 ft. To 10 ft. More than 10 ft

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Rock Quality Designation i.e. RQD Rock Mass Rating i.e. RMR Rock Tunnel Quality Index i.e. RTQI. Descriptions of these systems are given below: 24.4.2.1

RQD based classification

RQD was proposed by Deere in 1969 and is defined on the basis of core samples of the rock mass obtained in diamond core drilling carried out during exploration stage. It is defined by a very simple formula as under. RQD = (100 * L)/D where L = Total cumulative length of core pieces obtained in a diamond core drilling run, each with 100 mm or longer length. D = Total length drilled in the run. Since L is dependent upon the diameters of drill rods, core barrel and the type of core barrel used in drilling, these parameters must be as follows: 1 2 3 4

The cores must be NW or larger size in diamond core drilling i.e. the core diameter should be at least 54.7 mm or 2.15 inches. Double tube swivel core barrel must be used. RQD value should be calculated immediately after recovering the core before the pieces break in subsequent rehandling. Core drilling operations should be carried out with the intention of getting maximum core recovery. From this viewpoint the drill rods used should be at most one size smaller than the size of core barrel. If the size of drill rods used for drilling is smaller than such size difference, heavy vibrations are caused in the drill string and the core samples get broken easily. Table 24.4 Standard dimensions in diamond core drilling. Size designations

BW

HW

HW

Diameter of the Drilled Hole mm Outside Diameter of Drill Rod mm Outside Diameter of Core mm

59.6 54.1 42

75.3 66.8 54.7

98.8 89.1 76.2

Table 24.5 Classes of rock quality designation.

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RQD value

Rock quality class

Less than 25% 25%–50% 50%–75% 75%–90% 90%–100%

Very Poor Poor Fair Good Excellent

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Relevant sizes of diamond core drilling equipment are given in Table 24.4. Rock quality classes according to their RQD values are given in Table 24.5. 24.4.2.2

RMR based classification

Classification based on RMR, proposed in 1973 by Bienawski, is also called the Geotechnical Classification. It is based on a few basic parameters relating to the geometrical and mechanical properties of the rock mass as listed below: 1 2 3 4 5 6

Uniaxial compressive strength of intact rock Rock Quality Designation Spacing of discontinuities in the rock mass Condition of the surfaces at the discontinuity Groundwater conditions Orientation of discontinuities relative to the engineered structure

Of these, the rating values for the first five parameters are to be appropriately chosen from Table 24.5. To the sum of these values an adjustment value for the orientation of discontinuities, chosen from Table 24.6, is to be added. The resulting sum will give the RMR rating that can be interpreted from Table 24.7 to give the class assigned to the rock mass. It is easy to observe that, whenever the rating value of a parameter is higher, the condition is favorable. Whenever conditions are very near the boundary of a particular criteria, the values can be interpolated or the graphs presented by Bienawski can be used. One such graph is shown in Figure 24.6. Table 24.6 Values of variables to be chosen for calculating blastability index proposed by Ghose.

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Parameter

Range of variation for the variable and value to be chosen for the range

Density of the Rock Mass in t/m3

<1.6

1.6–2.0

2.0–2.3

2.3–2.5

>2.5

Value for Density Ratio

20

15

12

6

4

Discontinuity Spacing in m

<0.2

0.2–0.4

0.4–0.6

0.6–2.0

>2.0

Value of Discontinuity Spacing Ratio

35

25

20

12

8

Point Load Strength Index in MPa

<1

1–2

2–4

4–6

>6

Value for Point Load Strength Index Ratio

25

20

15

8

5

Joint Plane Orientation

Dip into Face

Strike at an Angle to the Face

Strike Normal to the Face

Dip Out of Face

Horizontal

Joint Plane Orientation Ratio

20

15

12

10

6

Adjustment Factor 1 – for Highly Confined Condition

−5

Adjustment Factor 1 – for Reasonably Free Condition

0

Adjustment Factor 2 – for Hole Depth/Burden Ratio >2

0

Adjustment Factor 2 – for Hole Depth/Burden Ratio 1.5–2

−2

Adjustment Factor 2 – for Hole Depth/Burden Ratio <1.5

−5

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Table 24.7 Relationship between blastability index proposed by Ghose and the powder factor. Range of Blastability Indices Powder Factor in kg/m3

30–40 0.7–0.8

40–50 0.6–0.7

50–60 0.5–0.6

60–70 0.3–0.5

70–85 0.2–0.3

0.6

Powder Factor in lb/ton

0.5

0.4

0.3

0.2

0.1

0.0

0

1

2

3

4 5 6 7 8 9 10 11 12 13 14 Sonic Velocity in ft/s (x 1000)

Figure 24.6 Relationship of powder factor with sonic velocity.

Many other correction values and subsequent interpretations are required to be applied to these RMR ratings to get a more purposeful picture of the rock mass from the civil engineering viewpoint. For this purpose additional tables have been presented in books written on this subject. Presenting all such information is beyond the scope of this book. 24.4.2.3

RTQI based classification

Rock Tunnel Quality Index was proposed by Barton and others at the Norwegian Geotechnical Institute on the basis of several case histories of tunnels. Therefore, it is referred to as NGI Q Index. Since this index is applicable for tunnel construction, where a long stretch of rock mass in limited cross section is under consideration, it does not have much relevance on blasting in large surface mines. Therefore, it has not been discussed at length in this book.

24.4.3

BI index based classification

Blastability index is applicable to rock mass as well as rocks. Blastability index for rock mass aims at its direct correlation to fragment size distribution resulting from a

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blast. Therefore, when it comes to blasting, it is far more important than the basis of rock mass classification described earlier in this chapter. Several researchers have attempted to correlate many different properties of rock and rock mass to arrive at an index that will give an indicative idea of the results of blast in the rock mass. Some formulae have been described below. 24.4.3.1

Index proposed by Hansen (1968)

After carrying out experiments at Morrow Point Dam, Hansen proposed an equation, as below, for estimating the quantity of explosive required for optimum fragmentation of the rock mass. Q = B2 * (0.236 * (h/B + 1.5) + 0.1984 * C * (h/B + 1.5)) where Q = Total charge in a single blasthole with free burden in kg B = Burden in m H = Height of free face on m C = A rock constant to be estimated by tests He also proposed that the total charge Q computed by the above equation be corrected by the following equation. Qc = 0.8 * (F/E) * (S/B) where F = Fixation factor depending upon blasthole inclination E = Explosive factor depending upon the explosive S = Spacing B = Burden

Powder Factor for ANFO in kg/m3

0.8 0.7 φ + ι = 45°

0.6

φ + ι = 40° 0.5

φ + ι = 35° φ + ι = 30° φ + ι = 25°

0.4 0.3 0.2 0.1

0

10

20

30 40 50 60 70 Fracture Frequency No/m

80

90

Figure 24.7 Relationship between powder factor and fracture frequency.

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24.4.3.2

Index proposed by Hainen and Dimock (1976)

While working in a copper mine in Nevada in the USA, Heinen and Dimock investigated the correlation between powder factor and the velocity of sound wave in the rock mass. For the investigation they studied several blasts of rectangular patterns measuring 18 × 21 ft, 21 × 24 ft, 24 × 27 ft, 27 × 30 ft and 30 × 33 ft. The results obtained by them are plotted in Figure 24.6. In the figure the dotted line represents the mean of the rock mass sonic velocities for which the powder factor is valid. 24.4.3.3

Index proposed by Ashby (1977)

Ashby proposed an empirical equation for powder factor from his observations in Bougainville Copper Mine. The variables in his equation were fracture frequency which, in effect, meant the density of fractures in the rock mass, and the friction angle which is related to the joint shear strength. The equation is as follows: Q = (0.56 * tan(φ + ι))/D0.333 where Q = Powder factor for ANFO in kg/m3 φ = Friction angle in ° ι = Roughness angle in ° D = Frequency of fractures in number/m The graph constructed for the above equations is shown in Figure 24.7. 24.4.3.4

Index proposed by Langefors (1978)

As per the concept proposed by Langefors, for every rock mass there is certain powder factor C0 for which there is no appreciable throw. He then proposed that for satisfactory breakage of the rock mass the powder factor Q should be taken as Q = 1.2 * C0. For the brittle crystalline granite rock in which he carried out his research the value of C0 was found to be 0.17 kg/m3. For other rocks the value of C0 lies between 0.18 to 0.35 kg/m3. 24.4.3.5

Index proposed by Lilly (1986)

Lilly developed an equation for the blastability of rock mass based on five parameters related to the site conditions. This equation is as under: BI = 0.5 * (RMD + JPS + JPO + SGI + H) where BI = Blastability index RMD = Rock mass description

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JPS = Joint plane spacing JPO = Joint plane orientation SGI = Specific gravity influence H = Rock hardness on Moh's scale The values to be used for the above variables are as below. ROCK MASS DESCRIPTION

The RMD values are 10 – for powdery/friable rockmass 20 – for blocky rockmass 50 – for totally massive rockmass JOINT PLANE SPACING

The JPS values are 10 – for joints with spacing <0.1 m 20 – for joints with spacing 0.1 to 1 m 50 – for joints with spacing >1 m JOINT PLANE ORIENTATION

The JPO values are 10 – when the joint orientation is less than 10° with the horizontal plane. 20 – when the absolute difference between joint dip angle and face dip direction is less than 30°. 30 – when the absolute difference between joint dip angle and face dip direction is more than 60°. 40 – when the absolute difference between joint dip angle and face dip direction is between 30° and 60°. SPECIFIC GRAVITY INFLUENCE

The SGI values are to be calculated as: SGI = 0.25 * SG − 50 where SG = rock mass specific gravity in kg/m3 ROCK HARDNESS ON MOH’S SCALE

The value to be used for H is the hardness value lying between 1 and 10 on Moh’s scale of hardness. These values are given in an appendix at the end of this book.

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From the value of blastability index proposed by Lilly calculated by formula given above one can calculate the powder factor Q and energy E factor by using equations as under. Q = 0.004 * BI E = 0.015 * BI where Q = Quantity of ANFO in kg/ton of rockmass E = Energy required in MJ/ton of rockmass Similarly, the rock factor to be used in predicting fragment size distribution in the Kuz Ram model, proposed by Cunningham, can be calculated by multiplying BI (i.e. Lilly’s blastability index) by 0.12. 24.4.3.6

Index proposed by Ghose (1988)

Ghose proposed a blastability index based on many properties of rock and rock mass. His approach was somewhat similar to that of Lilly. The equation for the blastability index is as under. BI = (DR + DSR + PLR + JPO + AF1 + AF2) where BI = Blastability index DR = Density ratio DSR = Discontinuity spacing ratio PLR = Point load index strength ratio JPO = Joint plane orientation ratio AF1 = Adjustment factor 1 AF2 = Adjustment factor 2 The values for these factors are to be chosen from Table 24.6. Once the value of blastability is determined, it can be used to find out the powder factor from the correlation between the two as given in Table 24.7. 24.4.3.7

Index proposed by Gupta (1990)

Gupta et al suggested the following equation for charge factor based on their field observations. CF = 0.278 * B-0.407 * F0.62 where B = Effective burden in m F = Protodyakonov strength index

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The Protodyakonov strength index is to be calculated from the following equation. F = 1.06 * C2/E where C = Compressive strength in kg/cm2 F = Modulus of elasticity in kg/cm2

24.4.3.8

Index proposed by JKMRC (1996)

The Julius Kruttschnitt Mineral Research Center in Australia has developed a blast fragmentation model through the efforts of many researchers working in the organization. A rock factor used in the fragmentation analysis can be used for prediction of the powder factor. Exact details and the formulae used for the calculation are not divulged but are built into software prepared by the organization. Factors taken into consideration while calculating the powder factor include: strength, density and Young’s modulus of the rockmass, average in situ block size, influence of structure, target fragment size, heave desired, confinement provided, scale of operation and groundwater. For appropriate input in the software the values of the variables are obtained as below: 1 2 3 4

Strength, density and Young’s modulus by laboratory tests. Block size through field measurements of exposed rock surface. Target fragment size is the desired value. For all other factors the input value is to be chosen from a scale of 1 to 9. A value of 5 is to be treated as neutral, values 4 to 1 are progressively favorable and 5 to 9 are progressively adverse.

Cast blasting needs the highest heave. A front end loader needs larger heave to have a spread (instead of a heap) of the blasted material. For a bench with an open free face the confinement is neutral. The JLMRC fragmentation model is gaining wider acceptability. 24.4.3.9

Index proposed by Han, Weiya and Shouvi (2000)

These researchers have used an Artificial Neural Network approach for determining rockmass blastability through a computer program. The logic of the equation is based on the following Expression: K = f {dcp, L, S, Rcd, Ed, Pc}

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Table 24.8 Blastability calculation by JKMRC software.

Parameter

Dragline Operation

Dragline Operation with Cast Blasting

60 2.51 12

60 2.51 12

Shovel Operation

Shovel Operation in Wet Conditions

Front End Loader Operation

50 2.47 10

50 2.47 10

40 2.42 10

Rockmass Parameters Rock Strength in MPa Density in g/cc Young’s Modulus in GPA Structural Parameters Block Size in m Structural Favorability*

2 5

2 5

2 5

2 5

0.3 3

0.3 5 5 5

0.3 5 5 5

0.15 7 7 7

1

1

5

1

0.30 0.08 0.51 0.03 0.24 0.61

0.25 0.13 0.26 0.00 0.17 0.42

0.25 0.13 0.26 0.08 0.21 0.52

0.20 0.06 0.36 0.02 0.16 0.39

Design Parameters Target Fragment Size in m Heave Desired* Confinement of Blast* Scale of Operation*

0.5 5 5 3

0.5 10 5 3

Environmental Parameters Groundwater Presence*

1

Output Given by Software Strength Breakage Heave Modifier Powder Factor in kg/ton Powder Factor in kg/m3

where K dcp L S Rcd Ed Pc

= = = = = = =

0.30 0.08 0.25 −0.02 0.18 0.44

Output parameter i.e. blastability Index Mean fragment size in mm Total length of fractures in a block measuring 2 × 2 m Mean distance of fractures in the 2 × 2 m block Dynamic compressive strength of rock in MPa Dynamic elastic modulus of rock in GPa Percentage of unqualified blocks in %

Nowadays, many rotary blasthole drills are equipped with a system to measure various rock characteristics based on the penetration rates and the power required in terms of rotary speed, weight on the bit and torque, while the drilling is being carried out. Analysis of the data obtained through such systems by means of software enables planning of a blast in a much more effective and accurate manner than the use of the different equations for blastability mentioned above.

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Chapter 25

Methods and patterns of charging blastholes

25.1

INTRODUCTION

After all the blastholes to be detonated in one blast are drilled, they are loaded with explosive. Even with the use of modern safe explosives, this has remained most dangerous operation in a mining cycle. In large surface mines most of the blasthole charging is done through mechanical means. The equipment used for the purpose differs depending upon the type and quantity of explosive. Some idea of the manner in which a blasthole is loaded with explosive, has been given in an earlier part of this book. This chapter deals with such topics in greater detail.

25.2

MECHANIZED BLASTHOLE CHARGING SYSTEM

Even the smallest amongst large surface mines use blastholes of diameter 200 mm or larger. When such a blasthole of 10 m depth is loaded with ANFO the weight of the explosive works out to about 267 kg. Some of the largest blasts in the largest mines have detonated more than 1000 tons of explosive within a short time span of 1 min. Such huge quantities of explosives can only be handled by mechanical means. Further, charging blastholes through mechanical means is advantageous in several ways, some of which are: 1 2 3 4 5 6 7

Fewer persons required for the operation Very fast charging operation Ability of using bulk explosives Use of full blasthole volume Possibility of attaining highest charge density Possibility of increasing charge density in certain parts of the blasthole. Ability of adding certain sensitizers in certain parts of the blasthole.

All these advantages lower the cost of the charging operation so much that it easily offsets expenditure in investment and maintenance of the fleet of bulk loading equipment.

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Usually the storage depot of the explosives is situated at a long distance from the mine bench. The only means of bringing the explosive to the mine bench is through trucks. The most commonly used explosives in large surface mines are: ANFO and Heavy ANFO Bulk Slurries and Emulsions. The equipment for charging blastholes with the above explosives differ from each other. The following sections gives details of the equipment.

25.2.1 Trucks for loading ANFO or heavy ANFO In the very early stages there may have been trucks that merely transported ammonium nitrate bags and fuel oil drums to the work site. However, modern trucks have the facility to carry all the components of the explosives safely in different compartments, enable their mixing at the site and pump them into the blasthole. A typical truck used for charging a blasthole with ANFO is shown in Figure 25.1. The internal construction of the charging truck is as shown in Figure 25.2. Ammonium nitrate, in the prill form, is stored in three or more compartments on the truck. A fuel tank on the truck stores an adequate volume of fuel oil. All these compartments or tanks are coated with specialized materials for long corrosion-free life. Ammonium nitrate is pushed into the mixing chute by the main augur. The rotary speed of the main augur is controlled and set by the operator depending upon the requirement of the explosive. Thus, ammonium nitrate flows at a predetermined rate. The fuel oil jet sprays fuel at the appropriate rate to match the flow of ammonium nitrate. As the mixture moves into the discharge arm by use of the discharge augur the fuel is well absorbed into the AN prill.

Figure 25.1 Charging of a blasthole by use of explosive mixed by a delivery truck.

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Figure 25.2 Internal construction of a charging truck.

The discharge arm can be swung by a hydraulic cylinder provided for the purpose so as to bring the discharge point above the blasthole mouth. This ensures that all the ANFO is poured directly into the blasthole. Stemming of ANFO-filled blastholes can be done immediately after the blasthole is charged.

25.2.2 Trucks for loading slurry or emulsion A typical truck meant for delivering slurry and emulsion explosives is shown in Figure 25.3. Trucks used for loading slurry or emulsion into the blasthole have many chambers to store the ingredients needed for making the explosive slurry. Typically the chambers store fuel, oxidizer, aluminum, sulfur and trace ingredients. These ingredients are pumped by the pumps provided for respective chambers into a mixing chamber in appropriate proportions as set by the blaster on the control panel. After mixing, the slurry explosive formed in the process is made to flow through a long hose by a screw pump. Some slurry explosive charging trucks have two cabs. The charging cab located on the rear side has the main control unit. The panel on the control unit has counters and totalizers to monitor the quantity being delivered. The calibration of the unit is made at the manufacturer’s plant to ensure that the slurry explosive has the advertised properties. Most of the units are capable of making slurry with viscosity up to 60000 centipoise.

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Figure 25.3 Truck for charging a blasthole by slurry explosive.

When the blasthole to be charged has a significant quantity of water, the slurry delivery hose is lowered to the bottom of the hole. This ensures that the slurry explosive starts getting filled from the bottom and the water in the blasthole is lifted out from the blasthole mouth. Slurry explosive pumping units have discharge rates normally ranging to 300 kg/min. For the gassing and gelling to be completed it may take about 5 to 10 minutes in the case of most of the explosives. For this reason stemming of the slurrycharged blastholes is started after waiting for about 10 minutes for each blasthole. The trucks used for charging slurry can also be used for charging emulsions into the blastholes. The delivery settings have to be appropriately adjusted for such changeover. Since emulsions take as much as 40 to 50 minutes for gassing and gelling, it becomes necessary to wait for nearly an hour before stemming of the blasthole can be started. In the case of emulsion explosives, the gassing process can be accelerated by using acetic acid.

25.2.3

Safety features of bulk delivery system

Bulk delivery systems generally have the following safety features. The ingredients of the explosives are nonexplosive in nature. They can, therefore, be transported very safely. The chambers used on the trucks for storage of the ingredients are coated with resins that have high insulation properties. The diameter of the delivery hose is kept below the critical diameter which is of the order of 100 mm. Since explosive is directly charged from the bottom of the blasthole, and on the top of this a considerable depth of the hole is filled with stemming material, it

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Figure 25.4 Lowering of slurry explosive hose to the bottom of the blasthole.

is cannot be easily scooped from the surface of the ground. This ensures that the explosive cannot be stolen. The truck unit is equipped with a few chains that roll on the ground and carry static electricity from the truck components to earth. In case of slurries or emulsions the gassing process is slow and the mixture poured in the blasthole takes some time to acquire the explosive properties.

25.3

BLASTHOLE CHARGING PATTERN

The most fundamental form of charging a blasthole with ANFO is shown in Figure 20.1. As briefly explained there, the detonator, primer, booster and main explosive are the four components of a charged ANFO blasthole. Some more information about these components is also contained in Chapter 22. When the main explosive detonates, the energy released is in two forms viz. strain energy ET and bubble energy EB. The total energy TE released by the explosive is TE = ET + EB. For any explosive, TE depends upon its composition and characteristics and as such it is constant. However, if the confinement of the explosive is very tight and without any air gas in it, the strain energy component is larger. On the contrary, if the confinement is tight but with some air gap, the bubble energy component is larger. Thus, when very strong and brittle rocks are to be fragmented, confinement without any air gap is more suitable. Addition of Al is also desirable. For fragmentation of relatively weak rocks confinement in the blasthole with an air gap is more desirable. In fact recent research indicates that use of an air gap is beneficial in every type of blast.

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The technique of leaving air gaps in explosive columns is called air decking. Since every air gap causes a discontinuity, the explosive column gets divided into more than one zone. When the length of a blasthole is more than 10 m or when the blasthole is in a rock mass that has many layers, there is a possibility of water seeping into the blasthole. Such water inflow makes a part of the explosive column very weak and the detonation may not continue from one side of such a weak zone to the other side. For this reason, it becomes necessary to divide the explosive column into two or more zones. The components of an explosive column are a detonator, a primer, a booster and the column of main explosive.

25.3.1 Type and placement of primer To properly charge a blasthole with ANFO as the main explosive, the arrangement, positioning and other relevant characteristics of the above components must be determined. In this regard the following points are of great importance. The most important objective of a primer cartridge is to start detonation of the ANFO column with maximum attainable velocity of detonation for the diameter of blasthole (i.e. explosive column), right from the point of detonation. This objective can be achieved only by using the primer cartridge with diameter close to the diameter of the ANFO column, and by using the primer that generates maximum detonation pressure in a confined state. Both these hypotheses are supported by Figure 25.5 and Figure 25.6. Table 25.1 gives the velocities of detonation that can be achieved for ANFO densely filled in blastholes of different diameters. As stated earlier, the primer is usually made of pentolite which is a mixture of PETN and TNT in equal proportion. Pentolite primer cartridges are available in different sizes.

Velocity of Detonation of ANFO in m/s

5000 Primer of 75 mm Diameter Primer of 64 mm Diameter 4000 Primer of 51 mm Diameter

3000

Steady State VOD

Primer of 25 mm Diameter 2000

Diameter of ANFO Column = 75 mm

1500 0 100 200 300 400 500 Distance from Initiation Point in mm

Figure 25.5 Effect of diameter of primer cartridge on velocity of detonation of ANFO.

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Velocity of Detonation of ANFO in m/s

5000

Curve Detonation Pressure of Primer

A

4000

577

B C

3000 D

A

24000 MPa

B

13500 MPa

C

5000 MPa

D

4000 MPa

E

700 MPa

Steady State VOD

2000

Diameter of ANFO Column = 75 mm

E 1500 0 100 200 300 400 500 Distance from Initiation Point in mm

Figure 25.6 Effect of detonation pressure generated by primer cartridge on velocity of detonation of ANFO.

Table 25.1 Velocities of detonation for ANFO columns of different diameters. Diameter of ANFO

Column VOD (m/s)

50 100 150 200 250 300 350

3000 3600 4000 4250 4400 4570 4650

For very large blastholes a primer of slurry or emulsion can also be used in conjunction with the primer cartridge. Such primers are made in plastic bag form as shown in Figure 25.7. They must be assembled by correctly inserting a detonator within their body as shown in the figure. The primer cartridge must not be tamped nor dropped into the blasthole. Run up distance is the distance in the explosive column from the point of initiation of detonation to the point where maximum velocity of detonation is achieved. For convenience it is measured in terms of the diameter of the explosive column. As shown in Figure 25.8, it is about 8 diameters for an ANFO column. Therefore the length of primer cartridge or bag should be equal to about 8 times the blasthole diameter. The quantity of primer is also important. As a general rule it can be said the weight of the primer in g should be equal to diameter of blasthole in mm. ANFO is also available in pressure-packed bags. In these bags the density of ANFO is about 1.1 kg/L rather than 0.8 kg/L in the case of poured ANFO. These ANFO bags

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Figure 25.7 Primer made from emulsion in bag form.


10000 8000 6000

PETN+TNT

ANFO

4000 2000 0 0 1 2 3 4 5 6 7 8 9 Number of Hole Diameters from Point of Detonation

Figure 25.8 Run up distance for ANFO and PETN + TNT.

are used in blastholes where water seeps inside the blasthole. Initiation of detonation in such watery blastholes is done by cast primer. While placing the bags one above another in a blasthole with only one primer in the bottom, as shown in Figure 25.9A, one of the bags may rupture and the explosive in the bag may desensitize. In such cases only a part of the explosive column gets detonated. To avoid such a situation a larger number of cast primers are introduced as shown in Figure 25.9B.

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Cast Primer Wet ANFO that may not detonate at all.

Cast Primer ANFO Bags Cast Primer A

B

Figure 25.9 Use of pre packed ANFO bags.

In waterlogged blastholes, use of poured slurry or poured emulsion is most appropriate. When these explosives are detonated by a primer cartridge, as the detonation front travels, the explosive in viscous or semi viscous state gets compressed to the point that it becomes insensitive. Naturally, the detonation does not propagate beyond this zone. In such instance more than one cast primer has to be used. For determining the number of cast primers an empirical equation as under is used. Nb = L/(30 * D) + 0.73 where Nb = Number of primers L = Length of explosive column in m D = Diameter of blasthole in mm Thus, in the case of a blasthole of 200 mm diameter and 20 m depth excluding 1.8 m subdrilling the number of cast primers to be used can be found as below. As will be explained in the next chapter, the stemming length presumed for this blasthole works out to 25D i.e. 25 * 0.200 = 5 m and required subdrilling works out to 8D i.e. 8 * 0.2 = 1.6 m. Therefore, L works out to 20–5.0 + 1.6 = 16.6 m. This gives Nb = 16.6/6 + 0.73 = 3.49. Therefore, to be on the safer side 4 cast primers should be used.

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While placing the primers, the lowermost primer should be kept at a distance of 4D from the bottom of the blasthole and the topmost primer should be kept at a distance of 4D from the bottom of the stemming column. In the rest of the column the primers should be equidistant. Thus in the above example the lowermost primer should be at a height of 0.8 m from the blasthole bottom. The depths of other three primers from the blasthole bottom should be 5.8 m, 10.8 m and 15.8 m. In very heavily waterlogged blastholes, use of emulsions in bagged form is most suitable. In charging these blastholes, bags are lowered into the blasthole and remain in stacked form. In such blastholes the likelihood of desensitization of explosive due to water contamination is virtually non-existent but the danger of desensitization due to explosive compression remains. For this reason, the blastholes do need more than one primer. These primers should be placed at a distance of 30D after the lowermost primer which is to be kept at a height of 4D from the blasthole bottom.

25.3.2

Direction of propagation of detonation

The position of detonator and primer in the column of main explosive determines whether the detonation travels from bottom to top or from top to bottom or in some other manner. The three types recognized on this basis are: 1 2 3

Top Priming Bottom Priming Multi Point Priming

All the three types are important and one of them is to be chosen depending upon the condition of the rock mass on the basis of the following details. In surface mines till some two decades ago top priming was practiced to a larger extent but this trend is now reversing, and in many instances bottom priming is preferred. This, of course, largely depends upon the ground conditions. The most important factor in bench blasting is the pressure exerted at the toe by the blast of explosive in the blasthole. When the detonation propagates from the top to bottom the velocity of detonation is insignificantly less than that when the detonation propagates from bottom to top. With a velocity of detonation of 4000 m/s about 4 ms are required for the detonation wave to travel from bottom to top or top to bottom in a blasthole of 20 m depth and 4 m stemming. In bottom priming the point of initiation of detonation is never at the bottom but at some distance above the bottom. Due to this, actually two detonation zones, one traveling upward and other traveling downward as shown in Figure 25.10A, simultaneously exert pressure at the toe. In the case of bottom priming, where the detonator is some distance below the top of the main explosive column, the upper wave detonation zone quickly reaches the bottom of the stemming column and results in venting off of the detonation pressure built up in the blasthole. The lower detonation wave travels downward and when it reaches the floor level, pressure is exerted in the toe region.

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Burnt Explosive

Upper Component

Resultant

Upper Component

Resultant

Lower Component

Figure 25.10 Top and bottom priming.

Pressure

Bottom Priming Top Priming

Time

Figure 25.11 Curves of pressure at toe for bottom and top priming.

If very precise measurements of the pressure at the toe are taken by some means, the indicator curves for the pressure at the toe, in case of bottom and top priming, will look like those shown in Figure 25.11. 25.3.2.1

Top priming

The main advantage of top priming is that it is safer than other types of priming. This is so because the detonator is introduced in a charged blasthole at much later stage. At the same time the stemming in top primed blastholes is somewhat casual and weak. Therefore, venting of explosion gases starts rather easily. When blasting is carried out in coal mining or similar sedimentary formations where one layer is to be fragmented by leaving the adjacent layer intact, there exists an inherent weak plane at the toe. In such circumstances the likelihood of toe

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problems is very low. For such blasts top priming can be equally as effective as bottom priming. Similarly, in a rock mass of low compressive strength, top priming can be resorted to without any particular disadvantage. 25.3.2.2

Bottom priming

In hard rock formations, where the likelihood of toe problems is greater, bottom priming is practiced. In some instances use of a booster placed in the vicinity of the primer at the bottom of the blasthole is also advisable. It has been estimated that the peak strain level at the toe is about 37% higher in the case of bottom priming as compared to top priming. The need for shorter subdrilling length is also likely in bottom primed blastholes. 25.3.2.3

Multi point priming

When two detonation fronts travel towards each other and collide at a point the pressure built up is highest. It is noticed that the pressure magnitude at the point of collision is as much as 46% higher than that from one-way traveling detonation. This fact can be utilized with advantage in blasting of rock masses that contain many layers with varying weakness. One such example is shown in Figure 25.12, where one pair of primers is placed at the level of each of the two surfaces of hard rock layer. The detonation zones collide within the hard rock layer and create heavy stress in the hard rock layer to shatter and fragment it very effectively. The other place where two primers can be used is at the floor level as shown in the figure. This primer placement reduces toe problems in hard rock.

Primer

Point of Collision Hard Rock Layer

Primer

Detonation Front

Figure 25.12 Multi point priming.

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25.3.3

583

Continuous side initiation

By placing the detonating cord near the wall of the blasthole it is possible to initiate the detonation from one side of the explosive column rather than in the center of it. Such detonation propagates at low velocity and produces more bubble energy as compared to strain energy. For this reason, this type of initiation is more suitable in fissured rock masses where more bubble energy is necessary.

25.3.4

Air decking

Air decking was first introduced in Russian mines as early as the 1940s but remained forgotten until it was revived by Melnikov and Marchenko in the 1970s. The principle behind air decking is that in some parts of the explosive column the explosive is replaced by atmospheric air. This is achieved by using one of the many available devices. The air gaps surely increase fragmentation size but at the same time the quantity of explosive is reduced to a much greater extent. Thus, there is an overall cost saving by adopting the air decking technique. Some idea about the advantage of air decking can be gained by reviewing Figure 25.13. It can be easily concluded from the figure that even if 40% of the explosive column is replaced, the increase in the mean size of the fragments yielded by the blast is only 9% or so. In most cases this increase in fragment size does not affect the loading and hauling operations and affects the crushing only to a very small extent. The above indirectly means that a saving of about 30% cost is possible by using air decking of 40%. The difference of 10% is due to the cost of additional consumables and accessories required for air decking and extra efforts required in charging the blasthole. For creating an air deck in the explosive column, some consumables are needed to be firmly placed at appropriate places in the blastholes. These consumables are available in two forms, viz. gas bags and hard plastic plugs. In the early days of air decking gas bags were lowered into the blasthole to the required depth and were inflated by a compressor or gas cylinders from the surface of the ground. Such types of gas bags did not develop strong friction between their

Increase in Fragment Size in %

Hard Rock

Soft Rock

Rip Rap

Pre Split

50 40 30 20 10 0

0

10

20

30 40 50 60 70 Air Deck Volume in %

80 90

100

Figure 25.13 Advantage of air decking.

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outer surface and the blasthole wall so as to bear the load of the explosive column and stemming material. They could be used only near the surface. In the improved version of gas bags, the bag contains a specific quantity of vinegar in a sachet and sodium bicarbonate. When the gas bag is lowered at desired depth the sachet is broken. As the vinegar comes in contact with sodium bicarbonate, carbon dioxide is evolved. It fills up the bag and expands it to firmly adhere to the walls of the blasthole. In yet another form of gas bag, an aerosol is contained in the bag. Just before the bag is lowered into the blasthole the aerosol cap is pressed down. The aerosol starts slowly releasing air and expanding the bag. A time of about 20 s is available to the blaster to lower the bag to the requisite depth before it catches on the blasthole wall and grips well with it. Some manufacturers have made expandable plastic plugs made from a special type of hard plastic. These can be used in place of gas bags. In general all these consumable accessories allow air decks to be created in a satisfactory manner. In some blastholes water cannot be completely removed and a column of water always remains inside it. In such blastholes, decks are created by using air bags or plastic plugs. The water decks contain water instead of air. Depending upon various conditions of blastholes and surrounding rock mass, many suitable air decking patterns have evolved. Patterns shown in Figure 25.14 are based on the experience of air decking practice described in a technical publication by Mintech, Australia. As will be seen from the patterns shown in all the figures, there are no specific distance markings. This is because it is not possible to definitely define what percentage of explosive can be replaced by an air or water column to give acceptable fragmentation. For coming to a conclusion about the percentages, field trials need to be carried out. The best way to carry out field trials is to reduce the explosive by 10% and replace it with air deck volume without any change in stemming height. If the fragmentation is much below the acceptable size of fragments, then increase the air deck volume to 15% by reducing the explosive to 15%. Some benefits that accrue by using air decks in blasting practice are summarized here below. 1 2 3 4 5 6 7 8 9 10 12 13 14 15 16

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Considerable reduction in explosive quantity Better confinement of explosion gases Less ejection of stemming material Better fragmentation at bench surface Huge reduction in ground vibration level Less air blast Better charging efficiency Reduction in explosive cost Possibility of increasing concentration of explosive in hard rock layers Less backbreak Better heave Consistent fragmentation level Possibility of using ANFO in place of emulsion Less fine generation Clean and stable bench face with no toes or overhangs.

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Dry Blastholes in Soft Rock Mass

Stemming

Dry Blastholes in Layered Rock Mass

Dry Shallow Blastholes in Hard Rock Mass

Dry Deep Blastholes in Hard Rock Mass

Air Bag

Air Deck

ANFO

Figure 25.14 Air deck pattern different rock mass conditions.

Wet Blastholes With Slowly Seeping Water

Water Deck

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25.3.5

Priming under special rock mass conditions

Some geological conditions of rock mass need special considerations while priming the blastholes drilled in the benches. The conditions can be titled as below. 1 2 3 4

Well Defined Ore Layer Heavy Water Seepage in Blasthole Hard Boulder in Soft Bed Cavities in the Rock Mass

The following elaboration deals with the ways in which such conditions can be tackled while priming the blastholes. 25.3.5.1

Well defined ore layer

This condition often arises in mines of the minerals that are found in sedimentary rock masses. In most surface coal mines there is a well defined coal layer separated from the overburden by a surface as shown in Figure 25.15. In such coal mines the overburden is removed by using a dragline, whereas the coal is excavated by using shovels as shown in Figure 26.4. One of the primary needs of blasting in such a rock mass is that by blasting in the overburden, the coal layer should not get disturbed. If coal gets fragmented along with overburden, quite a large portion of coal mixes with overburden and gets wasted. To accomplish this objective, blastholes are always drilled only down to the top surface of the coal layer. While charging the blastholes the bottom one or two meter portion is filled with stemming material before charging the blasthole. The blasthole is charged as per conventional priming practice. 25.3.5.2

Heavy water seepage in blasthole

When water seeps profusely into the blasthole, dewatering fails. In such circumstances a thick but flexible plastic tube can be lowered into the blasthole with some steel

Stemming Strong Bed

Explosive Weak Bed Stemming

Strong Bed

Explosive Backfill Coal

Figure 25.15 Pattern of blastholes charging in coal mines.

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ingots at the bottom of the tube. A charging hose is also placed into the tube while it is being lowered. ANFO can be charged in the blasthole. It may be necessary to use a pneumatic charging method. Such blastholes can also be charged with emulsion explosive because this has excellent resistance to water. 25.3.5.3

Hard boulder in soft bed

On rare occasions large boulders get embedded in a soft clay mass. A blasthole proceeds to the desired depth through this boulder. When explosive is loaded in the conventional manner the portion of the blasthole in the boulder many not have explosive at all as shown in Figure 25.16 A. When such a blasthole detonates, the energy is absorbed by the clay layer and the boulder remains intact. To avoid such situations the portion of the blasthole within the boulder is filled with explosive as shown in Figure 25.16B. This charge is called a pocket charge. In practice a boulder can be detected by careful study of the penetration logs of a group of adjacent blastholes. It can be seen by a blasthole camera. 25.3.5.4

Cavities in the rock mass

Some minerals, concentrated in the rock mass, can get leached to form a cavity in the ground. Such conditions can occur in iron ore or limestone rock masses. In such circumstances poured explosive will keep on filling the cavity as shown in Figure 25.17A and the blast may become a huge devastating explosion. To avoid such consequences the blasthole has to be charged very carefully. The first step is to detect the cavity. This is possible by careful study of the penetration rate log. In most cases the drill string either gets quickly dropped in the cavity or the penetration rate increases to disproportionate magnitude. A cavity can also be seen by using a blasthole camera. B

A Stemming

Boulder

Weak Bed

Stemming

Explosive Stemming

Explosive Strong Bed

Bench Floor Level

Figure 25.16 Fragmenting a boulder in clay mass.

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Such a blasthole can be charged at the bottom up to a level a little lower than the cavity bottom and then a hard plastic plug can be firmly fixed into the blasthole. After this some wooden props are lowered into the blasthole and another hard plastic plug is fixed in the blasthole on the top of the props in the manner shown in Figure 25.16B. After this the blasthole is charged with explosive to the requisite height and the remaining upper portion is filled with stemming material.

25.4

DRILLING AND FIRING PATTERNS

Some information on delay elements introduced in the detonators was given earlier. Sequential delay in the timings of initiation of explosive in the blasthole is an important aspect of every blast. If blastholes in two or more rows are fired at the same instant i.e. without any delay, then the rock mass in the portion lying between the first row and the bench face gets sufficient space beyond the free face to move the fragments horizontally. However, since the fragments formed from the rock mass portion between the first and second rows also try to move in a horizontal direction, they have much less space to move horizontally. The energy released by the blast then tries to move the fragments upward to produce long-traveling flyrocks in huge quantity. Further, when detonation of explosive in all blastholes is simultaneous, the hazard of airblast and ground vibration is also excessively large. Most importantly, simultaneous detonation of explosive in all the blastholes gives very poor fragmentation of the rock mass and requires considerable secondary blasting. This is quite evident from the photograph shown in Figure 27.18, where the left side half of the bench was blasted by appropriately delayed detonation whereas the right hand half was blasted instantaneously. When looked at from the bench surface the blastholes have collaring points at specific locations. This layout is termed the “Drilling Pattern”. B

A Stemming

Explosive in the Blasthole

Hard Plastic Plugs

Explosive filled in the cavity Wooden Props

Bench Floor Level

Figure 25.17 Charging a blasthole that passes through cavity.

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By making intelligent use of delays, blastholes are also fired at different timings. This is irrespective of their layout. If a line connecting all the blastholes that detonate simultaneously is treated as a criterion, there are many shapes in which the blast takes place. The layout of such lines is called a firing pattern. The sequence of detonation timing is called the firing sequence.

25.4.1

Drilling patterns

On the surface of a bench blastholes are drilled in one of the three patterns, viz. square, rectangular, and triangular. These are called drilling patterns and are shown in Figure 25.19. These should be distinguished from firing patterns. When a blasthole is drilled in homogeneous and joint-free rock and is detonated after charging, the zone of cracking around the blasthole is circular as seen in Figure 25.20. If such zones from two adjacent blastholes overlap, then such overlapped portion will have excessive cracking. On the contrary, if there is no gap between the two zones it will generate large pieces of rock mass. Both these situations are undesirable. The upper drawing in Figure 25.21 shows four blastholes in square pattern. The dimensions of the square on which they are positioned are such that there is a zone of

Figure 25.18 Difference between fragmentation achieved by delay based blast and instantaneous blast.

Square Pattern

Rectangular Pattern

Triangular Pattern

Figure 25.19 Blasthole layout patterns.

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Figure 25.20 Cracking zone around a blasthole. Square Pattern Blasthole Zone of Fragmentation Zone of Excessive Fragmentation Zone of No Fragmentation

Triangular Pattern Blasthole Zone of Fragmentation Zone of Excessive Fragmentation Zone of No Fragmentation

Figure 25.21 Layouts of blasthole positions.

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excessive fragmentation as well as of no fragmentation. If the blastholes are brought near to each other the zone of no fragmentation will reduce and will disappear when the length of the square has a certain unique value. But simultaneously the zone of excessive fragmentation will increase. Similar is the case of a triangular pattern shown in the lower drawing in Figure 25.23. It can be geometrically shown that for each of the blastholes in square pattern the area of excessive fragmentation equals 0.570796 * r2 when the zone of no fragmentation disappears. Similarly in a triangular pattern the area of excessive fragmentation equals 0.181172 * r2 when the zone of no fragmentation disappears. This clearly means that blastholes in a triangular pattern will utilize the explosive in a better way. Square and rectangular patterns are still used in practice, but to take advantage of the triangular pattern many times, an offset is given to the blasthole positions in adjacent rows. With this the position of blastholes becomes very similar to that of a triangular pattern as shown in Figure 25.22. Naturally, the advantages of a triangular pattern are achieved by this type of layout to a good extent. With the ease of positioning a blasthole drill precisely by use of GPS systems, the choice of rectangular and square patterns is slowly waning away.

Square Pattern

Staggered Square Pattern

Figure 25.22 Blasthole layout staggered patterns.

200

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B S Firing Time in ms Bench Face

Figure 25.23 V pattern with two directions of throw.

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25.4.2

Firing patterns

In large surface mines five commonly adopted firing patterns are used. They have been summarized below. 1 2 3 4 5

“V” Pattern Echelon Delay Pattern Flat Face Pattern Channel Pattern Sinking Hole Pattern Following are the details of each of these.

25.4.2.1

“V” Pattern

This type of firing arrangement is made on either a square or rectangular drilling pattern. A typical V pattern firing arrangement is shown in Figure 25.23. The blastholes on two rows have the same delay timings. These rows make an angle of 45° with the bench face and 90° with each other. They are in V shape as can be seen from the figure and hence the name. This firing pattern is also called a chevron pattern. V pattern is the most commonly used firing pattern in large surface mines because it has many advantages as listed below. 1 2 3

The fragments of rock are in the form of a tall heap. Material in such a heap can be easily lifted by shovels and loaded into the dumpers. Since rock fragments are thrown towards each other in a 90° angle they collide with each other and further reduce in size. The collision of fragments reduces the hazard of flyrocks. Sometimes, when the number of rows fired in the same blast is more than five, two centrally located blastholes in the first row are given the same delay time. The V shape arrangement of firing then takes a shape as shown in Figure 25.24. This firing pattern Firing Time in ms 350 300 250 200 175 150 150 175 200 250 300 350 300 250 200 175 150 125 125 150 175 200 250 300 250 200 175 150 125 100 100 125 150 175 200 250 200 175 150 125 100

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150 125 100

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25

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Bench Face

Figure 25.24 V pattern with three directions of throw.

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593

increases the probability of collision to give even better fragmentation and throws the muck pile a little farther. 25.4.2.2

Echelon delay pattern

Echelon Delay Pattern is shown in Figure 25.25. It is sometimes referred to as half echelon pattern. It is particularly suitable for mine benches which have two free faces. This firing pattern gives a muck pile that is not flat but well spread on the bench floor. Such muck piles are suitable for loading by wheel loaders. 25.4.2.3

Flat face pattern

The flat face firing pattern is adopted when the number of rows in the blast is five or less. The drilling pattern of the blastholes is usually staggered or triangular because with such a drilling pattern the fragmentation is within acceptable limits. When the blast is meant for getting large stone pieces to be used in a dam or breakwater, a square or rectangular drilling pattern is better suited. A typical flat face firing pattern is shown in Figure 25.26. 25.4.2.4

Channel delay pattern

Sometimes mining activities have to start from the side of a hill. In such case there is no bench and so no bench top or bench floor. A classic excavation of this type of operation is a channel cut for some distance, made before the entry into a large tunnel. The blasted material must be made to accumulate in the blasted area where shovels can approach from one side and the dumpers also approach from the same side. For the purpose of loading muck into the dumper the shovel must slew through a 180° angle rather than the usual 100° to 120°. Under such situations the most suitable firing pattern is the channel delay pattern shown in Figure 25.27.

Firing Time in ms 700 600 500 450 400 350 300 250 200 175 150

500 450 400 350 300 250 200 175 150 125 100 450 400 350 300 250 200 175 150 125 100

75

400 350 300 250 200 175 150 125 100 B S 350 300 250 200 175 150 125 100 75

75

50

50

25

Second Bench Face

600 500 450 400 350 300 250 200 175 150 125

First Bench Face

Figure 25.25 Echelon delay pattern.

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150

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Bench Face

Figure 25.26 Flat face pattern.

Firing Time in ms 200 150 125 100

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Bench Face

Figure 25.27 Channel delay pattern.

On many occasions in such situations the blastholes may have different depths and have to be charged very carefully. Further, the rock mass fragmented by the blast of one hole does not move to a significant distance to create space for blast of the subsequent holes. To remedy this drawback, at least to some extent, the delay interval between the rows of blastholes with the same delay is kept larger. It is very usual to find small diameter blastholes made by small rotary percussion drills in such situations even if the mining project is to be worked with large rotary blasthole drills. Needless to say that almost all large mining projects do have small rotary percussion drills for miscellaneous jobs.

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25.4.2.5

595

Sinking hole pattern

Many mines, particularly coal mines, are located on horizontal planes. To start a mine in such terrain, initially a huge excavation, much like the excavation for a foundation of a skyscraper, is started by blasting vertical holes. The pattern chosen for such blasts is as shown in Figure 25.28. Blast starts with detonation of the explosives in centrally located blastholes. Later blastholes surrounding these holes, but which lie on the periphery of the diamondshaped area, are detonated with the same delay. In this manner the blast spreads outward. In such a blast the fragments do not move in a horizontal direction because there is no space for such movement. Naturally, movement of rock fragments in the upward direction is on a much larger scale. Ground vibrations and flyrock hazard of this type of blast is much higher than in other types of blasts with an equal quantity of explosive. To reduce both these hazards the delay intervals are kept large.

Firing Time in ms 500 450 400 350 300 350 400 450 500 450 400 350 300 250 300 350 400 450 400 350 300 250 175 250 300 350 400 350 300 250 175 125 175 250 300 350 300 250 175 125

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Bench Face Absent

Figure 25.28 Sinking hole pattern.

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Chapter 26

Design of a surface blast

26.1

INTRODUCTION

Earlier chapters dealt with different aspects of explosives and rock fragmentation through a blast. Blasthole drilling is a relatively straightforward subject without much danger, but blasting is rather intricate and highly dangerous. To avoid any fatalities and tragic circumstances, very careful planning is necessary in choosing various parameters related to blastholes, selecting explosives and the other accessories used for engineering the blast. This chapter is devoted to giving the basic knowledge for the design of a very efficient blast in a surface mine without causing any recognizable hazards to the surroundings.

26.2

BLASTHOLES IN A MINE BENCH

A blasthole drill stands on a mine bench top and drills blastholes in it. Blastholes drilled in the rock mass appear as shown in Figure 26.1. In the early days of blasthole drilling, when churn drills were used for the purpose, the blastholes could only be drilled in a vertical direction. This trend continued during the dawn of rotary blasthole drilling as well. Now, inclined blastholes as shown in the right hand drawing in Figure 26.1, are also drilled in many mines. As will be seen, inclined blastholes often prove more appropriate from the viewpoint of reduction in cost. Ideally the angle of inclination of blastholes is 45°, but practical limitations of rotary drills limit the angle of inclination to 30°. On certain occasions, in addition to the main blastholes, some additional vertical or inclined blastholes, like those shown in the left hand drawing in Figure 26.2, are also required for the appropriate fragmentation of rock mass. These blastholes are of half or even quarter depth depending upon the characteristics of the rock mass. Such blastholes are referred to as satellite blastholes. Similarly, situations also arise when additional horizontal blastholes like those shown in the right hand drawing in Figure 26.2 become necessary. They are termed snake holes.

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Burden

First Row Second Row Bench Edge Burden

Bench Floor Subdrilling

Blastholes

Bench Height

Subdrilling

Angle of Inclination of Blastholes

Figure 26.1 Vertical and inclined blastholes in a bench.

Additional Vertical Blastholes termed Satellite holes Additional Horizontal Blastholes termed as snake hole

Figure 26.2 Additional vertical blastholes near bench top or horizontal blastholes at the bench floor.

26.3 TYPES OF BLASTING IN SURFACE MINES No doubt, blasting activities in large mines are carried out for breaking the rock mass, but the associated objectives are different for each one of them. This gives rise to different types of bench blasting techniques, described briefly in the following subsections.

26.3.1

Conventional bench blasting

Conventional bench blasting is also called production blasting. In this type of blasting the purpose of the blast is to break a large portion of rock mass to yield such fragment sizes that can be easily loaded by using rope shovels or hydraulic shovels into the dumpers. These dumpers move the overburden to a waste site and the ore to a crusher site or store yard. In surface coal mines when the coal layer is very thick, even the coal is blasted and loaded into dumpers by using front end loaders. Loading by this equipment is shown in Figure 26.3.

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Figure 26.3 Loading of a dumper by using rope shovel, hydraulic shovel and front end loader.

26.3.2

Secondary blasting

The loading and moving equipment used in mines has limitations. Pieces of rocks beyond a certain size can neither be lifted nor carried. Even the crusher cannot accept rock pieces beyond a certain size for further crushing. Economical reasons always force the blast designer to design the blast in such a way that the largest fragments yielded through the blast are of a size acceptable to the lifting or moving equipment. Due to so many uncontrollable factors in blasting, in some cases large rock pieces are left even by a well designed blast. Such large pieces can be moved only after their further blasting. Such blasting of large rock pieces is called secondary blasting. In early days these large boulders were blasted after drilling one or more blastholes and charging them. Nowadays hydraulic excavators equipped with a hydraulic hammer attachment are more often used for this purpose.

26.3.3

Cast blasting

In many places coal is found in horizontal layers that underlie a large bed of overburden. In such cases the most economical method to reach the coal is by blasting the top overburden layer in the form of a long strip and moving it to one side by using a giant walking dragline as shown in Figure 26.4. In such mines the technique of cast blasting can be adopted with advantage. The advantage of cast blasting can be appreciated by means of the sketch in Figure 26.5. If the bench is blasted in the normal way, the profile of the top of the muck heap reaches points ABCD. Here the blasted overburden in the area ABF has already moved to the area of heap of shifted overburden without use of the dragline. Similarly, if the bench is blasted by using the technique of cast blasting the top of the muck heap reaches points LMNP. In this case all the blasted overburden in the area LMF has already moved to the area of the heap of shifted overburden without use of the dragline. Since area LMF is much larger than area ABF cast blasting proves greatly advantageous by reducing the workload of the dragline. (See also article 26.7.3 below).

26.3.4

Presplitting

In simple words presplitting means creating a fully cracked surface between the area to be blasted and the area to be kept intact. Such a surface is created by

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Figure 26.4 Overburden removal by a walking dragline. Limit Line up to which the Profile After C Profile After Heap of Shifted Overburden Standard Blast Cast Blast can extend D N P Overburden Yet to be Bench Profile Blasted M Heap of L Before Blast Shifted d B Overburden Coal Bed F

A

Figure 26.5 Advantage of cast blasting.

charging closely spaced, small diameter blastholes drilled in a line and blasting them instantaneously. Diameters of blastholes for the presplitting operation is of the order of 100 to 125 mm. The most common method of drilling such blastholes is by rotary percussion drilling through hydraulic drifters because it gives the fastest penetration rates. Usually the presplit surface is at an angle of 15° with the vertical as shown in Figure 26.6. The small diameter of the blastholes and the close spacing ensure that a well-trimmed smooth surface is formed without any damage on the either side of the blasthole plane. The presplit surface does not allow significant transmission of shock waves from the main blast into the rock mass meant to be kept intact. To meet this objective, often a row of buffer blastholes is also drilled in front of the presplit blasthole line. Blastholes in this row are charged to evolve limited energy. Ideally a single fracture connects adjacent blastholes, and half of the hole remains at each presplit hole. If excessive crushing and radial cracking is observed at the plane it indicates excessive use of explosive.

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Row of Large Buffer Blastholes Row of 100 mm Presplit Blastholes

Rows of Large Production Blastholes

Final Face Berm 2m

H 75° 7D 5D

Figure 26.6 Presplitting behind the main bench.

Closer spacing with an appropriate quantity of explosive gives a better surface. However, because of the high cost of drilling, the optimum spacing is the largest at which radial cracks will join and form a continuous undamaged surface. Smooth blasting is also a type of presplitting where the blastholes are of much larger diameter, often drilled with the same drill that is used for drilling production blastholes in the main bench. Rotary blasthole drills like the Bucyrus 39R are very suitable for drilling blastholes for smooth blasting because they can drill blastholes at −15° angle, and therefore can carry out drilling operations by standing on the main bench. This enables leaving much smaller berm as illustrated in Figure 26.7.

26.3.5

Snake hole blasting

Snake holes are the horizontal blastholes drilled at the bottom of the bench as shown in Figure 26.8. When these blastholes are drilled and blasted prior to blasting the main vertical or inclined blastholes, they create a plane of separation in a manner very similar to the presplitting blastholes. The advantage of snake hole blasting techniques are: 1 2 3

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Less fracturing of rock mass below the newly created bench floor. Eventually when this floor becomes the top of bench, the hazard of flyrock is at a reduced level. Improved floor of the bench where the movement of mining equipment is made smoother. Need for less concentration of explosive at the bottom of the main vertical or inclined blastholes in the bench.

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602


Very limited berm required because the mast has been placed very near the bench face. Such placement was possible only because of the negative mast angle

This would have been the centerline of the blasthole if the drill did not have the capability of negative mastangle. In such case it would have become necessary to place the drill between the blasthole and the face of upper bench

15°

Figure 26.7 Usefulness of −15° drilling angle of Bucyrus 39R.

S Snake Holes 0.5+B

S/2

Figure 26.8 Snake hole blasting.

The use of snake hole blasting is often avoided because it needs an additional drill specially equipped to drill horizontal blastholes. Such a drill is usually good for drilling small diameter blastholes and cannot be replaced by the large rotary drills used for production blastholes. In some European countries snake hole blasting is common because in those countries even in very large mines production blastholes are of small diameter and the drills used for production hole blasting can also be used for snake hole blasting. In a few large US mines snake hole blasting has become unavoidable because geological conditions do not allow a smooth bench floor by the normally adopted technique of optimum subdrilling and subsequent blasting with heavy concentration of high explosives at the bottom of the blastholes.

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26.3.6

603

Rip rap blasting

When huge rock-fill dams are to be built on rivers, or long breakwaters are to be constructed for harbors, an enormous quantity of large pieces of rock are required. Depending upon design parameters their weight may reach even 3000 kg. A special blasting technique that yields such large rock pieces is called rip rap blasting (see article 26.7.1 below).

26.4 WHAT IS INVOLVED IN DESIGN OF A BLAST The outcome of a blast is gaged through the following parameters. 1 2 3 4 5 6 7

Fragmentation of the rock mass Displacement of the muck pile Profile of the muck pile Misfires Ground vibrations Flyrock Airblast

In designing a blast a designer has to give due consideration to all the controllable and uncontrollable variables that are likely to have an effect on the outcome parameters, and choose apt magnitudes of controllable variables in such a way that the magnitudes of the outcome parameters is within the desired range. The uncontrollable variables are properties of rock specimen, bedding, dip, strike, faults, joints, discontinuities, ground water and weather. Controllable variables are listed in Table 26.1.

26.4.1

Powder factor

The first step in designing a blast is to chose the appropriate explosive and ascertain the powder factor. In surface mining practice the most commonly used explosive is ANFO. As reasons for this choice have been explained earlier, they are not repeated here.

Table 26.1 Controllable variables involved in blast design.

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Parameter names

Parameter names

Explosive and Powder Factor Number of Free Faces Blasthole Diameter Blasthole Inclination Blasthole Spacing Stemming Height Blasthole Charging Pattern

Blasting Direction Size of the Blast Blasthole Depth Burden Subdrilling Blasthole Drilling Pattern Firing Sequence

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604


As ANFO is mixed on site and poured into the blastholes, it is possible to change the properties of the ANFO that goes into the blasthole at a certain depth. This is done by adding a sensitizer like Al or adding some more powerful explosive to the mixture. Actually by doing so we are changing the basic powder factor of ANFO. The basic powder factor can be chosen by using some formulae presented earlier, but Table 26.2 makes the task easier by presenting appropriate powder factors for ANFO under different conditions of surface mining. In the days of manual calculations, the blasts were designed without considering any variation of powder factor but with modern computerized blast design programs, a blaster gets guidance on changing the explosive density in the same blasthole depending upon the variation in properties of the intact rock and rock mass.

26.4.2

Blasting direction

As has been stated in the previous chapter, when the rock mass is in form of beds and the beds are dipping, the direction of blastholes must be such that shooting is achieved either with the dip of the beds or against the dip of the bed. The worst situation is when shots are along the strike. To remedy the situation the solution is in changing the direction of blast.

26.4.3

Number of free faces

A greater number of free faces available for a blast means more energy in reflected tension waves. In such situations a larger number of blastholes can be fired together. However, in practice no more than two free faces are possible for the same blast. When mucking operations are carried out by shovel and dumper or loader and dumper combination, blasting is carried out with only one free surface. When the mucking operation is to be carried out by dragline usually two free faces, as shown in Figure 26.4, are practiced. The dragline carries out scraping excavation on one surface and then by slewing into about 90° it empties the bucket in front of the other surface. This second surface, which is right angles to the first surface, becomes Table 26.2 Powder factors of ANFO in surface mining. Powder factor Type of mine

Method of excavation

Surface Metal Mining Surface Coal Mining with 60 yd3 (45.87 m3) Dragline 30 yd3 (22.94 m3) Shovel 17 yd3 (13 m3) Front End Loader Surface Coal Mining with Cast Blasting Quarrying Construction Blasting Open Excavation Trenching

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in lb/yd3

in kg/m3

0.6–1.0

0.356–0.593

0.5–0.7 0.6–1.1 0.6–1.6

0.297–0.415 0.356–0.653 0.356–0.950

0.9–1.5 0.6–1.5

0.534–0.890 0.356–0.890

0.25–0.8 2.0–3.0

0.149–0.475 1.187–1.780

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605

necessary to ensure that cast blasting can be carried out on the side of this surface. A careful look at the Figure 26.4 will give a lucid idea of the advantage of two free surfaces in such situation.

26.4.4

Blasthole diameter

Several factors are required to be considered while choosing the appropriate diameter for blastholes. The desired diameter of the blastholes in large surface mining operations is decided mainly by considering the following factors. 1 2 3 4

Desired rate of production Desired fragmentation Properties of rock specimen Properties of rock mass The following throws more light on other factors.

26.4.4.1

Desired rate of production

Since mining is a commercial proposition, the rate of production is fixed on an economical basis even before starting the mining project. It has the highest influence on the desired diameter of the blasthole. If consideration is given to the unconfined compressive strength of rocks, the approximate relation between production rates and blasthole diameter is as shown in Table 26.3. In surface mining practice the rate of drilling and rate of removal of the blasted rock must match. Thus, the diameter of blastholes is loosely related to the capacity of the shovel bucket as matched in Table 26.4. When a dragline is used for overburden removal, the diameters of blastholes can be even larger because the bucket capacity of the dragline chosen for the operation is much larger. Therefore, even if very large pieces of rocks are formed in blasting they can be moved by the dragline. The usual range of diameters of blastholes while using a dragline is 250 mm to 381 mm. Table 26.3 Average production rate in formations of different hardness from blastholes of different diameters. Likely production rate in m3/h for one m blasthole length in rocks with compressive strength as under

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Blasthole diameter

Soft rock UCS < 70 MPa

Medium rock UCS 70 to 180 MPa

Hard rock UCS > 180 MPa

200 250 311

600 1200 2050

150 300 625

50 125 270

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Table 26.4 Blasthole diameters based on shovel bucket capacities. Bucket capacity of the shovel (m3)

Hole diameter range in mm

4.5 7.5 9.17 11.5 15.3 20 35 50

76–127 127–215 171–250 200–270 229–311 250–349 270–381 311–445

Blasthole diameter is also influenced by the bench height. This will be elaborated later in the appropriate section. 26.4.4.2

Desired fragmentation

Closely spaced small diameter blastholes certainly give smaller fragment size but large diameter blastholes do not give proportionally larger fragment size. What the distribution of particle size will be after a blast can be predicted with reasonable accuracy by many mathematical models developed for the purpose. One such model is the Kuz-Ram model. Table 26.5 presents the prediction of fragment size distribution of blasts for 203 and 311 mm diameter blastholes, where all other factors remain the same except those mentioned in the table. The data in the table prove both the points mentioned in the previous paragraph. Actually such mathematical models – even though they indicate a relationship between fragment size distribution and blasthole diameter – are not useful for determining blasthole diameters, but can be used for verification that the fragment size will be good enough for the loading and hauling equipment when a particular diameter is selected. 26.4.4.3

Properties of rock specimen

In the absence of a more rigorous approach as explained in a previous chapter, the powder factors to be chosen for some of the commonly occurring rocks are given in Table 26.6. If blastholes are charged with more than the optimum amount of explosive, the rock mass gets fragmented to small size. With this, the need for crushing decreases but the overall cost increases. 26.4.4.4

Presence of geological structures

Properties of the rock mass also need to be taken into account while fixing the diameter of blastholes. The number and placement of joints in the rock mass may

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Table 26.5 Fragment size distribution as per Kuz-Ram model. Percent passing for blasthole with following details Fragment size (m)

Hole dia. = 203 mm Burden = 5.075 m Spacing = 6.343 m

Hole dia. = 311 mm Burden = 7.775 m Spacing = 9.718 m

0.00 0.05 0.10 0.15 0.20 0.25 0.30 0.35 0.40 0.45 0.50 0.55 0.60 0.65 0.70 0.75 0.80 0.85 0.90 0.95 1.00 1.05 1.10

0.0% 5.2% 17.6% 33.6% 50.2% 65.1% 77.1% 85.9% 91.9% 95.6% 97.7% 98.9% 99.5% 99.8% 99.9% 100.0% 100.0% 100.0% 100.0% 100.0% 100.0% 100.0% 100.0%

0.0% 4.0% 13.7% 26.9% 41.4% 55.5% 67.9% 78.0% 85.7% 91.1% 94.7% 97.0% 98.4% 99.2% 99.6% 99.8% 99.9% 100.0% 100.0% 100.0% 100.0% 100.0% 100.0%

Table 26.6 Powder factors for different rocks.

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Rock category

UCS range MPa

Very Soft Soft

about 50 50–100

Medium

100–200

Hard

200–300

Very Hard and Hard plus Tough

More than 300

Rock name Coal, Potash, Soft Shale, Marl Sandstone, Hard Shale, Limestone, Slate, Conglomerate Schist, Hornfels, Hard Limestone, Marble, Serpentinite, Dolomite Andesite, Dolerite, Hard Iron Ore, Basalt, Granite, Taconite, Skern, Quartzite, Hard Basalt

Powder factor in kg/m3 0.15–0.25 0.25–0.40

0.4–0.60 0.60–0.70 0.70–1.00

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necessitate reduction of blasthole diameter, fixed on the basis of desired production rate. A rock mass in a mining bench is rarely monolithic. It is usually inter-spread with joints. A monolithic block of the rock exists only within the boundaries of such joints. The process of formation of cracks and subsequent fragmentation of rock mass starts from the blasthole and continues undisturbed till it encounters some discontinuity in the form of joint, fault, fold, unconformity etc. Beyond such a discontinuity the formation of cracks and subsequent fragmentation is reduced considerably. Suppose the structure and spacing of joints in a bench is like the one shown in Figure 26.9 A, and if large diameter blastholes with larger spacing as shown in the figure are drilled and blasted, then two of the blocks do not have any blasthole within them. This will result in poor fragmentation within those two blocks. However, if smaller blastholes with smaller spacing, as shown in Figure 26.9B, are chosen then each of the blocks has at least one blasthole within it and will thus result in good fragmentation in all the blocks. Reduction of diameter of blastholes beyond a certain point can also be counter-productive. This happens because as the diameter of the blastholes is reduced, the number of blastholes required to be drilled increases disproportionately. This increases the cost of drilling. Further, the cost of explosives, cost of accessories, and cost of charging the blastholes increase for a larger number of blastholes. In very small diameter blastholes it is not possible to use inexpensive explosives like ANFO.

Bench Face Spacing

Block Without Blasthole

Burden

Block Without Blasthole

A - Rectangular Pattern Comprising of Fewer Large Blastholes Bench Face

Spacing

Burden

B - Rectangular Pattern Comprising of Many Small Blastholes

Figure 26.9 Dependency of blasthole diameter on joint structure.

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26.4.5

609

Blasthole depth

Really, this should be called length of blasthole. However, in surface mines, as most of the blastholes are in a vertical direction, the term is colloquially called depth. The length of the blasthole includes subdrilling. It depends upon height of bench, subdrilling and angle of inclination of the blasthole. For the purpose of material removal after blasting, one of the three alternatives are used. 1 2 3

Combination of Shovel and Dumper Walking Dragline. Combination of Wheel Loader and Dumper

The manner in which these combinations work differs, the most appropriate method of drilling and blasting also differs for each case. 26.4.5.1

Combination of shovel and dumper

Shovel-dumper combination is chosen in most of the metal mines, where the ore body is irregularly placed as shown in Figure 1.3 or on coal mines where the ore body is not in a horizontal or nearly horizontal layer. As shown in Figure 26.10, a shovel starts scraping the blasted rock into its bucket at the level of the ground on which it stands and moves the bucket forward while raising it till the maximum attainable height H is reached, or the bucket is completely filled with blasted rock. It then slews into an angle of about 120° to 135°, with the bucket filled with the broken rock and held at a height of H1, which is higher than the height of the dumper body. When the bucket comes into position above the dumper body, the contents of the bucket are unloaded in dumper body by opening the bottom plate of the bucket. After dumping the broken rock, the shovel turns back again and starts scraping the broken rock face. A complete cycle of excavation and dumping of blasted rock takes average time of about 22 to 26 seconds. The maximum attainable height of scraping, called cutting height, for a shovel depends upon the machine model. For most of the shovels used in large surface mines the cutting height lies between 10 m to 18 m.

H

Scrape Surface H1

Figure 26.10 Working of an electric shovel.

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The height of the bench and so the depth of the blasthole is decided on the basis of maximum cutting height of the shovel, after giving due consideration to the subdrilling needed and the inclination of the blasthole. A formula, as given below, was proposed for fixing bench heights in metal mining that uses shovel-dumper combination. H = 10 + 0.57 * (Cc − 6) where H = Bench height in m Cc = Capacity of shovel bucket in m3 However, this formula should be used with great caution because it holds good only for a few small and medium size shovels. Table 26.7 gives some details of the bucket capacities and heights of cut for some well known rope shovels. 26.4.5.2

Walking dragline

Walking draglines are used in coal mines. They do not need any hauling equipment because with their long booms they can scrape blasted overburden and throw it a long distance away. The working of a walking dragline is shown in Figure 26.11. While removing the overburden lying above a horizontal or near-horizontal coal bed, a dragline stands on the blasted overburden and throws its bucket at a long distance on the top of the exposed coal bed. It then starts pulling the bucket towards itself. In this action the bucket scrapes the overburden and gets filled. The filled bucket is then lifted well above the ground and the dragline turns into an angle of about 90° and dumps the material into a heap on the side. As the draglines are far bigger than shovels, the depth to which they can excavate is much greater. Boom length, bucket capacity, depth of excavation for some well known walking draglines is given in Table 26.8. In some coal mines the coal beds are at a shallow depth and the thickness of the overburden layer is small. In such cases the bench height and depth of blastholes is decided on the basis of the thickness of the layer and inclination of the blastholes rather than the depth of excavation of the dragline.

Table 26.7 Heights of cut for some electric mining shovels.

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Make

Model

Bu. capacity m3

Bu. capacity MT

Height of cut (m)

Bucyrus Intl. Bucyrus Intl. Bucyrus Intl. Bucyrus Intl. P&H Mining P&H Mining P&H Mining

295HD 295HR 395 495 2300XPC 2800XPC 4100XPC

21.3 25.5 35.7 61.2 25.5 35.7 61.2

38 45 63.5 109 45.4 63.5 108.9

18.4 13.5 16.6 16.8

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611

Lifting Rope Slewing Axis of Dragline Scrape Surface Pulling Rope H Coal Bed

Blasted Overburden

Figure 26.11 Working of a walking dragline.

Table 26.8 Depth of excavation for some walking draglines.

Make

Model

Bucyrus Intl.

680 W

Bucyrus Intl.

W2000

Bucyrus Intl.

8050

Bucyrus Intl.

8200

Bucyrus Intl.

8750

P&H Mining

9010C

P&H Mining

9020C

P&H Mining

9030C

26.4.5.3

Boom length (m)

Bucket capacity (m3)

58 (min) 90 (max) 75 (min) 101 (max) 99 (min) 99 (max) 84 (min) 122 (max) 109 (min) 132 (max) 100 (min) 107.7 (max) 88.4 (min) 123.4 (max) 99.1 (min) 129.5 (max)

24 12 34 24 61 45 84 51 116 76 57 38 89 54 117 90

Excavation depth (m)

50.5 74.5

56.4 68.3 54.8 85.7 56.5 76.7

Combination of wheel loader and dumper

In coal mines the layers of coal are soft and distinctly separate from the overburden. Therefore, wheel loaders are sometimes used as equipment for loading coal into dumpers in coal mines.

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A wheel loader works differently than a shovel or a dragline. With its bucket in the lowermost horizontal position, as shown on the right side in Figure 26.12, it moves forward. With this movement the coal or other loose material gets filled into its bucket. When sufficient volume of loose material is filled into the bucket it turns the bucket in upright position and simultaneously starts lifting the bucket in a backward movement on a curve. By the time it goes sufficiently on the back side, the bucket is in its highest position. The wheel loader again starts moving on a curve to approach the dumper placed on its side. When it has moved sufficiently towards the dumper and its bucket is in position just above the body of the dumper, the wheel loader tilts its bucket so as to unload the material in the dumper body. This completes one cycle and the dumper is again ready to start a second cycle. Unlike the shovel, a wheel loader does not scrape material. It has relatively low height and its bucket raises only to sufficiently high level to load a dumper of matching size. A wheel loader finds it a little more difficult to load material from a steeply sloping heap but is at ease while loading material from a low height, well spread heap. It is, therefore, well suited for handling coal. Bench height and, therefore blasthole depth, is limited in the case of coal beds where a wheel loader has to work. The bench heights selected for blastholes of different diameters varies considerably. Usually followed trends in this regard are as expressed through Table 26.9. If blastholes in a bench are vertical with no subdrilling, their blasting gives rise to a large stump as shown in left hand drawing in Figure 26.13. That is the reason why subdrilling to some depth below the intended ground surface created after blasting becomes essential. Subdrilling value i.e. length of subdrilling, is usually kept to 8 to 12 times the blasthole diameter.

26.4.6

Blasthole inclination

Often blastholes are drilled at some angle with the vertical rather than the usual vertically downward direction. Such inclined blastholes offer many advantages but many disadvantages as well. Various parameters associated with vertical and inclined blastholes are shown in Figure 26.13. Advantages of a blast engineered with inclined blastholes are: 1

Length of explosive column in the blasthole increases and at the same time true burden, shown by B1 in the figure, decreases. With this either better

Figure 26.12 Working of a wheel loader.

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Table 26.9 Bench heights chosen for blastholes of different diameters. Bench heights in m (ft) Blasthole dia. mm

ins

152.40 158.75 171.45 187.33 200.03 228.60 250.83 259.88 279.40 311.15 349.25 381.00 406.40

6 6.25 6.75 7.375 7.875 9 9.875 10.625 11 12.25 13.75 15 16

9.14 10.06 (30) (33)

10.97 11.89 13.11 14.02 14.94 15.85 17.07 17.98 18.90 20.12 21.03 21.95 22.86 24.08 24.99 (36) (39) (43) (46) (49) (52) (56) (59) (62) (66) (69) (72) (75) (79) (82)

Preferred

Not Preferred

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B

B

H1 = H * sec α B1 = B * cos α

B H1 H Zone of Poor Fragmentation SD

α

B1

SD1

Figure 26.13 Parameters associated with vertical and inclined blastholes.

2 3

4

5

6

7 8

9

fragmentation results or the burden B can be increased to get the same level of fragmentation. Volume of zone of poor fragmentation near the bench top, shown by square hatching in the figure, decreases. This means better overall fragmentation. Length of subdrilling can be decreased because formation of hump at the newly formed bench floor is decreased. In certain cases subdrilling can be totally eliminated. Yield of blast, i.e. the volume of blasted rock per meter length of the blasthole, is higher in case of inclined blastholes as compared to vertical blastholes. For this reason cost of drilling and blasting is decreased. Possibility of backbreaks is much less in a blast with inclined blastholes. Backbreaks caused by vertical blastholes like those shown in Figure 26.14 are very dangerous. In many cases they make it difficult to place the blasthole drill for drilling holes in the first row. New bench face gets a profile in such a way that the burden for the inclined blastholes in the next round is more or less uniform. The slope of the bench face is also more stable. This is elucidated in Figure 26.15. Probability of misfires is reduced. Large quantum of explosive energy, which always radiates in a plane perpendicular to the blasthole alignment, is directed towards the free surface. In addition to better fragmentation this also results in less energy for creating ground vibrations. Reduced over-crushing of the rock mass, hence less wastage in the case of minerals like coal. Disadvantages linked with inclined blastholes are:

1

2 3

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Aligning the blasthole drill for inclined blastholes is rather difficult and often results in alignment error by incorrectly choosing the spot for starting the drilling operation. Collaring i.e. start of drilling an inclined blasthole, is somewhat more difficult. Chances of hole deviation while drilling inclined blastholes are far higher. Drilling operations have to be done very carefully.

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615

Figure 26.14 Dangerous backbreak. Non Uniform Burden With Vertical Blastholes

Uniform Burden With Inclined Blastholes

Figure 26.15 Uniform burden with inclined blastholes.

4

5 6 7 9

More wear and tear is experienced by all the drill string components. This is particularly true in the case of large diameter rotary drills that have to exert very high feed force and torque on the drill string. Since the blasthole drill has to resist heavy horizontal force, the fatigue life of the drill is also reduced. Throw and flyrock hazards of blasting inclined holes is somewhat higher than that of vertical holes. Lesser penetration rate due to reduced efficiency of flushing. Larger quantum of rock mass is thrown at a greater distance. Due to this the muck pile is of lesser height.

All the above disadvantages become more and more severe as the blasthole inclination increases. In mines equipped with wheel loaders and dumper combination, inclined blastholes are practiced in many cases. The angle of inclination is usually of the order of

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15° to 20° but in some cases it is as high as 30°. Properties of the rock mass, particularly the orientation of joints and other types of discontinuities, play an important role in the choice of angle of inclination. The muck pile obtained from a blast of vertical blastholes is more suitable for shovel and dumper combination. Almost all metal mines practice blasts of vertical blastholes. Long inclined blastholes tend to deviate more than vertical blastholes. Therefore, blastholes in dragline mining are usually vertical.

26.4.7

Burden

If burden is excessively large, then only cracks will be developed in the rock mass. There will be no separation and hence no fragmentation will occur. The energy released by the detonation of explosive will be utilized in causing heavy vibrations in the rock mass. As against this, if the burden is very small, the gases will escape towards the bench face with very high velocity and the fragments of the rock near the face will be thrown in the air violently to cause flyrock problems that can prove disastrous or fatal in many instances. Several formulae have been proposed by mining technologists for the calculation of burden. They all relate burden value in terms of some other parameters. It can be stated that evaluation of burden by using these formulae gives a value lying somewhere between 25 * D to 40 * D, where D is the diameter of the blasthole. The most appropriate value of burden certainly depends upon the hardness of the rock mass. Figure 26.16 shows the variation of burden for different diameters of blastholes and different rocks. Density of rock as well as type of explosive used for blasting have significant influence on the value of burden, because in the case of heavier rocks a higher part of

Soft Rock

9 8

Medium Hard Rock Hard Rock

7

Burden in m

6 5 4 3 2 1

50


300

350

Figure 26.16 Burden as a function of blasthole diameter.

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the explosive energy is spent in lifting and throwing the rock mass. In order to avoid this, burden is reduced while blasting with low energy explosives or when the bench has high density rock mass. Table 26.10 gives guidelines for calculating the value of burden for rocks of different densities in case of ANFO and Slurry Dynamite which are low and high energy explosives respectively. Many researchers have tried to establish a correlation between burden spacing and other blasthole parameters. Details are given in a later subsection.

26.4.8

Blasthole spacing

Spacing is usually calculated from the value of burden. When blastholes are drilled for a mine bench, spacing of value 1.1 * B to 1.5 * B, where B is the value of burden, is chosen. Even within this range the value of 1.1 * B has been found to be more appropriate for large diameter blastholes and 1.5 * B is more suitable for small diameter blastholes. Figure 26.17 gives guidelines for appropriate choice of spacing in terms of burden. Spacing value of less than 1.0 * B should never be used except when using controlled blasting techniques such as smooth blasting or cushion blasting. When blastholes in a row are very closely spaced, a crack from one blasthole progresses to the other blasthole very rapidly. With this, the tendency of forming a huge block of rock mass between the first row and bench face increases. Large craters are also formed at the bottom of the blasthole. This gives rise to stumps at the toe of the previous bench face. If the blastholes have large spacing in the same row, the chances of formation of humps in the newly formed bench face increase. Similarly the likelihood of formation of stumps at the toe of the previous bench increases. Table 26.10 Dependence of burden on rock density and type of explosive. Values of burden in terms of blasthole diameter D for rocks of different densities Type of explosive

Low 2200 kg/m3

Medium 2700 kg/m3

High 3200 kg/m3

ANFO Slurry Dynamite

28 * D 33 * D

25 * D 30 * D

23 * D 27 * D

Spacing value in Terms of Burden

1.5 1.4 1.3 1.2 1.1 150


400

450

Figure 26.17 Spacing as a function of burden.

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Both these phenomena are illustrated in Figure 26.18. Determining the spacing-to-burden ratio is greatly affected by the geological conditions of the rock mass. Conditions when holes in the same row have to be blasted at different instances do arise. Often value of spacing is arrived at after onsite experience.

26.4.9

Subdrilling

Subdrilling means the length of blasthole drilled below the level of surface of the bench floor. This is well illustrated in Figure 26.1. In bench blasting it is always desirable that after the bench blast and subsequent removal of the fragments, the new floor formed will be at the same level as that of the original bench face, and there will be no stumps i.e. protrusions of unfragmented rock out of the newly formed surface. The need for subdrilling is to ensure these objectives are easily achieved. A blasthole is always initiated from its bottom. Therefore, the detonation zone travels from bottom to top. For this reason the compression waves do not travel in a plane perpendicular to the blasthole but at an upward angle. If a blasthole drilled without subdrilling is detonated from the bottom, there is always a zone as shown in the left hand sketch in Figure 26.19 where compression waves do not reach and tension waves are not reflected. Thus, the rock mass in this zone remains unfragmented, giving rise to stumps. Whereas if a blasthole drilled with subdrilling is detonated from the bottom, there is no zone of unfragmented rock as shown in right hand sketch in Figure 26.19. Thus, no stumps remain when the fragmented material is removed Stumps at the Toe of Old Bench Face

Stumps at the Toe Block

Block

Heavily Crushed Zone

Humps at the New Bench Face

Figure 26.18 Effects of spacing as a function of burden. Blasthole Direction of propagation of compression waves

Subdrilling Zone where compression waves do not reach

Figure 26.19 Need for subdrilling.

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619

Subdrilling length has to be optimum. If the subdrilling is too large it will result in: 1 2 3 4

Increase in drilling and blasting costs An increased vibration level Excessive fragmentation of the bench, affecting slope stability Increased risk of cutoffs and overbreak

Optimum length of subdrilling depends upon the strength of the rock. Table 26.11 presents the subdrilling length to be used for blasting in intact rocks of different strength. In cases where the rock is highly fractured, subdrilling length can be reduced by about 25% as shown. In some coal mines there is a horizontal plane of weakness between the overburden layer and coal layer. In such cases, to ensure that only the overburden layer is fragmented and removed without fragmenting the coal layer, no subdrilling is practiced. In fact, in such circumstances it may be worthwhile to have the bottom of the blasthole at some distance above the top of the coal bed. If the blastholes are inclined, the subdrilling length can be reduced as shown in Figure 26.20.

Table 26.11 Recommended subdrilling length in terms of burden. Strength classes of the rock

Subdrilling in intact rock condition

Subdrilling in highly fractured conditions

Soft Rock with Easy Toe Medium Rock with Normal Toe Hard Rock with Difficult Toe

0.1 B to 0.2 B 0.3 B 0.4 B to 0.5 B

0.07 B to 0.15 B 0.25 B 0.3 B to 0.4 B

Inclination Angle in °

40

30

20

10

0 0.2 B

0.3 B 0.4 B Subdrilling Length

Figure 26.20 Reduction of subdrilling with blasthole inclination.

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26.4.10

Stemming height

While charging the blasthole, some part of the blasthole is intentionally filled with inert material like rock chips formed in drilling. This is called stemming. Almost invariably some length of a blasthole at its top is used for stemming. The purpose of stemming is to confine gases generated in detonation of the explosive and utilize the energy generated by the blast in the most effective manner. If stemming length is very small it leads to a premature escape of the gases leading to an airblast and a danger of flyrock, the hurling of rock fragments in a blast. On the other hand, if stemming length is excessively long, the portion around a blasthole near bench top is not subjected to sufficient cracking by the blast and large boulders are formed. It also leads to reduced loosening of the rock and subsequent difficulty in its removal. Apart from stemming length, the other parameters associated with stemming column are the type and size of material to be used for stemming. Studies have shown that coarse angular material, such as crushed rock, is the most effective stemming product. The optimal stemming length varies between 18D to 30D, where D is the diameter of the blast hole. Crushed rock of size between 0.04 D and 0.06D is found to be most suitable as stemming material. It has been found to effectively lower the stemming length by about 41%. Use of stemming plugs in the blasthole gives lower mean fragment size. For some specific objectives, stemming is also used in between two or more zones of explosive columns.

26.4.11

Size of the blast

The shape of the blast refers to the geometrical form of the area on the bench top from which all the blastholes are detonated together with delays in their detonation timings. In surface mines the shape of the area is almost invariably rectangular. Such area has length and width. The length is measured along the bench crest and the width is measured perpendicular to the bench face. As the first, second, third rows, and so on, are blasted with delay timings, the fragmented rock mass from the burden of the first row is thrown on the bench floor at a long distance. The fragmented rock mass from the burden of the second row is also thrown horizontally but a larger portion of this sits on the top of the rock mass of the first burden but near to the bench. For the blast of the third row the same phenomenon is repeated and the height of the heap increases as shown in Figure 26.21. The advantage of a large blast area is that it allows more effective utilization of loading and hauling equipment. Disadvantages of a large blast area are: 1 2 3

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The intensity and duration of vibrations created in the ground by the blast increases. The face of the bench formed after removing the blasted material from bench floor becomes crooked like the one shown in Figure 26.15. Fly rock hazard increases as it becomes difficult to introduce appropriate delays in the blasting of rear rows.

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Fragments from Third Burden

621

Fragments from Fourth Burden

Fragments from Second Burden Fragments from First Burden

Figure 26.21 Formation of a heap of fragments after the blast.

If the equipment used for material removal can handle tall heaps of blasted fragments, then even 7 or 8 rows can be taken for a blast. In the case of coal beds the number of rows reduces to 3.

26.5

CALCULATION OF BURDEN

The most important parameter in blast design is the burden. This is because all the rockmass contained in the burden needs to be fragmented by the energy evolved in the blast. Due to the impossibility of mathematical derivation, many researchers have tried to arrive at empirical relations between many parameters of the blasthole and the environment and the burden. The following elaboration is based upon this. Some of these relations are explained here below.

Fraenkel (1952) B = (Rv * L0.3 * l0.3 * D0.8)/50 where B = Burden in m L = Length of blasthole in m l = Length of charge in m D = Diameter of blasthole in mm Rv = Resistance factor. It varies between 1 and 6. For rocks with high UCS it is 1.5 and for rocks with low UCS it is 5

Pearse (1955) B = Kv * 10–3 * D * (PD/RT)0.5 where B = Maximum burden in m Kv = Rock property constant that varies between 0.7 to 1.0 D = Diameter of blasthole in mm

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PD = Detonation pressure in kg/cm2 RT = Tensile strength of rock in kg/cm2

Hino (1959) This method of calculating the burden, proposed by Hino, takes into consideration many variables but somewhat intricately. The first equation is: B = (D/4) * (PD/RTd)1/n where B = Burden in m D = Diameter of blasthole in cm PD = Detonation pressure kg/cm2 RTd = Dynamic tensile strength in kg/cm2 The value of the characteristic constant n is to be determined from the data obtained in the crater-forming test for the explosive as under. n = log(PD/RTd)/log(2Do/(d/2)) where Do is the optimum depth of the center of gravity of the explosive charge in cm. It is to be determined graphically by using the equation Dp = ΔΣV10.333 where d = Diameter of the charge Dp = Depth of the CG of charge Δ = Relationship ratio of depth Dp/Dc Dc = Critical depth of CG of charge Σ = Volumetric constant of the charge V1 = Volume of charge used in the test CG means center of gravity.

Langefors and Kihlstrom (1963) Equation proposed by Langefors and Kihlstrom for calculating burden is as under. Bmax = (D/33) * ((ρe * PRP)/(Co * f * (S/B)))0.5 where Bmax = Maximum burden for good fragmentation in m D = Diameter at bottom of blasthole in m ρe = Density of explosive in the blasthole in kg/L

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PRP = Relative Weight strength of the explosive f = Degree of fixation of the blasthole. For vertical blastholes f = 1, for blastholes inclined at 3:1 f = 0.9 and for blastholes inclined at 2:1 f = 0.85 S/B = Spacing to burden ratio The value of Co is to be determined as under. Let the quantity of explosive needed to fragment 1 m3 of rock mass be called C. For hard rocks C is usually taken as 0.4. This value is to be modified as For B = 1.4 to 15 m Co = C + 0.75 For B = <1.4 m Co = C + 0.07/B From the Bmax determined by using above Formulae, the actual B to be used is to be calculated as B = Bmax − ec−db * H where ec = Collaring error in m/m db = Blasthole deviation in m H = Bench height in m

Lopez Jemino (1980) Lopez proposed the following equation for calculating burden B = 0.76 * D * F where B = Burden in m D = Diameter of blasthole in inches F = Correction factor The correction factor is to be evaluated by using following equations. F = fr * fe fr = ((2.7 * 3500)/(ρr * VC))0.333 fe = ((ρe * VD)/(1.3 * 36602))0.333 where ρr = Specific gravity of rock in g/cc VC = Velocity of seismic propagation in m/s ρe = Specific gravity of explosive in g/cc VD = Velocity of detonation for explosive in m/s

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These formulae are valid for blastholes of diameters from 165 to 250 mm. When the diameter of blasthole is larger the burden should be reduced by a factor 0.9.

Konya and Walters (1972 and 1983) In the year 1972 Konya and Walters proposed an equation for calculation of burden as under, B = 3.15 * D * (ρe/ρr)0.333 where B = Burden in ft D = Diameter of explosive in inches ρe = Specific gravity of explosive in g/cc ρr = Specific gravity of rock in g/cc On the basis of burden distance arrived at by using the above equation, the spacing distance can also be calculated as per following equations suggested by Konya and Walters. For instantaneous blast of single row blastholes S = (H + 2B)/3 when H < 4B S = 2B when H > 4B For sequenced blast of single row blastholes S = (H + 7B)/8 when H < 4B S = 1.4B when H > 4B In all the above equations B = Burden calculated as above in ft, S = Spacing in ft H = Bench height in ft Stemming distance to be kept as under, For massive rock T = B For Stratified rock T = 0.7 * B In the year 1983, Konya and Walters refined their approach and came up with the following equations. B = ((2ρe/ρr) + 1.5) * D Stemming distance T = 0.7B Subdrilling distance J = 0.3B

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26.6

625

RELATIONSHIP OF BLASTHOLE PARAMETERS

For every blast there is such a relationship between the geometrical parameters of the blasthole, that proves to be the best for the blast from the viewpoint of fragmentation as well as safety. Ash adapted the simplest approach to establish the relationship. The following elaborations are largely based on his paper published in 1963. All the parameters used for establishing the relationship are shown in Figure 26.22.

Spacing to burden relationship As has been seen earlier in this chapter, spacing is considered to be proportional to the burden and can be expressed as, S = Ks * B where Ks is the proportionality constant.

Burden to blasthole diameter relationship Let us presume that for a certain burden and spacing, one unit of blasthole depth gives satisfactory blasting. Therefore, a volume V of rock mass fragmented by the energy output of the explosive in one unit depth of blasthole works out as:

H = Bench Height D = Blasthole Diameter B = Burden S = Spacing J = Subdrilling T = Stemming Height L = Blasthole Length

S

Crest

S T

H B

L

B

Toe D

J

Figure 26.22 Symbolic nomenclature of a blasthole.

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V = B * S * 1 = Ks * B2 We can equate the explosive energy output from one unit length of blasthole as Eo = (π/4) * D2 * Eu * ρu Since in a blast the explosive is same, and the charging density of the explosive is constant, we have Eo ∝ D2 Similarly, Eo ∝ V ∝ D2 i.e. B2 ∝ D2 or B ∝ D This gives: B = Kb * D Naturally, as the diameter of the blasthole increases the burden also increases, in the burden-to-blasthole diameter ratio Kb.

Subdrilling to burden relationship As has been seen earlier, for all the explosives and particularly ANFO, the velocity of detonation at the point of detonation is somewhat low but increases as the detonation progresses. Further, the energy output is also proportional to the velocity of detonation. The purpose of keeping some subdrilling length is to ensure that the velocity of detonation in the explosive reaches its maximum when the detonation reaches the floor level of the bench. In case of ANFO it has been found that the length required for the velocity of detonation to reach its maximum is 6 times the diameter of the blasthole. It is also true that the detonator and primer together are never at the bottom of the blasthole but at a height of about 2D. Therefore, J = 8D and since D ∝ B, we have J ∝ B. In other words, J = Kj * B where Kj is the subdrilling-to-burden ratio.

Stemming to burden relationship In bench blasting practice, the stemming length is also considered to be related to the burden by a stemming-to-burden ratio Kt. In other words, T = Kt * B

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Based upon some tangible logic involved in the blasting process, Langefors and Kihlstrom have concluded that Kt = 0.7

Bench height to burden relationship Bench height is always more than burden. The ratio of bench height to burden is symbolized by Kh and the equation relating bench height and burden is H = Kh * B On the basis of a study of the outcome of many blasts, Ash established that Kh should be more than 1.6.

Length of blasthole When the blastholes are vertical the length of blasthole is the sum of bench height H and subdrilling length J. The equation below can be established in this regard: L=H+J When blastholes are inclined with the vertical at an angle β, the relationship takes the form: L = H/cosβ + (1 − β/100) * J

26.7

DESIGN OF CONVENTIONAL SURFACE BLAST

Various parameters of a conventional bench blast with blastholes are shown in Figure 26.25. A method to affix the values of these parameters is as under. This method is based on the crater theory put forth by Livingstone. One of the basic assumptions for the blasthole is that Length/Diameter is less than 50. 1 Ascertain the diameter of blasthole D, on the basis of hourly production requirements by using Table 26.3. 2 On the basis of the diameter of the blasthole and the type of explosive used for blasting, find the burden B, and spacing distances S, by using Table 26.12. 3 By using Table 26.13 determine the subdrilling length J, in terms of blasthole diameter. 4 Arrive at the value for stemming length T, on the basis of blasthole diameter from Table 26.14. 5 Bench height H, can be determined on the basis of Table 26.15. As shown by Table 26.9, great variation can be allowed in bench height. 6 Length of blasthole L = H + J 7 Specific volume in m3/m Vb = π * (D2/4) * 10−6

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Table 26.12 Burden and spacing for different explosives in different types of rock mass. Type of rock mass and compressive strength in MPa Type of explosive

Parameter

Soft UCS < 70

Medium 70 < UCS < 180

Hard UCS > 180 MPa

ANFO

Burden B Stemming S Burden B Stemming S

28D 33D 38D 45D

23D 27D 32D 37D

21D 24D 30D 34D

Watergels/ Emulsions

Table 26.13 Subdrilling length for blastholes of different diameter. Blasthole diameter Fragment weight in kg

180–250 mm

250–445 mm

Subdrilling-J

7 D to 8D

5 D to 6D

Table 26.14 Stemming length for different types of rock masses. Type of rock mass and compressive strength in MPa Parameter

Soft UCS < 70


Hard UCS > 180 MPa

Stemming-T

40D

32D

25D

Table 26.15 Bench height for different types of rock masses. Type of rock mass and compressive strength in MPa Parameter

Soft UCS < 70


Hard UCS > 180 MPa

Bench Height-H

52D

44D

37D

8 9 10 11

Breakage volume is Vr = B * S * H Yield of breakage is Yb = Vr/L Bottom charge length Lb = 8D to 10D. Column charge length is Lc = L − T − Lb

From the above basic parameters and properties of explosives like density of bottom charge ρb and density of column charge ρc, other parameters related to the blast can be determined as below.

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629

Concentration of column charge Cc = ρc * Vb Concentration of bottom charge Cb = ρb * Vb Total bottom charge Qb = Lb * Cb Total column charge Qc = Lc * Cc Total quantity of explosive Qt = Qb + Qc Powder Factor Vr = Qr/Vr

Example 1 In a large copper mine, blastholes of diameter 311 mm are being drilled to cope with the desired production rate. The formation has UCS of 150 MPa. ANFO with a charge density of 0.85 kg/L is used as the explosive in the column charge, and emulsion with charge density of 1.35 kg/L is used as explosive in the bottom charge. Determine various parameters of the blast.

Solution 1 Since D = 311 mm and formation UCS is 150 MPa, Burden B = 23D = 23 * 0.311 = 7.153 m Spacing S = 27D = 27 * 0.311 = 8.397 m Subdrilling J = 6D = 6 * 0.311 = 1.866 m Stemming length T = 32 * D = 32 * 0.311 = 9.952 m Bench height H = 44 * D = 44 * 0.311 = 13.684 m Length of blasthole L = H + J = 15.55 m Specific volume Vb = π * 3112/4 * 10–6 = 0.07597 m3/m Breakage volume Vr = 7.153 * 8.397 * 13.684 = 821.91 m3 Yield of breakage Yb = 821.91/15.55 = 52.856 m3/m Bottom charge length Lb = 8 * 0.311 = 2.488 m Column charge length Lc = 15.55–9.952–2.488 = 3.11 m Density of bottom charge is 1.35 kg/L = 1350 kg/m3 Density of column charge is 0.85 kg/L = 850 kg/m3

Stemming Length = 15D

B

Variable Length of Column Charge

Intermediate Stemming = 1 m Bottom Charge = 55D

Figure 26.23 Rip rap production blasting.

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Concentration of bottom charge Cb = 0.07597 * 1350 = 102.56 kg/m Concentration of column charge Cc = 0.07597 * 850 = 64.57 kg/m Total bottom charge Qb = 2.488 * 102.56 = 255.17 kg Total column charge Qc = 3.11 * 64.57 = 200.81 kg Total quantity of explosive Qt = 255.17 + 200.81 = 455.98 kg Powder factor = 455.98/821.91 = 0.555 kg/m3

26.8

DESIGN OF OTHER TYPES OF BLASTS

Apart from conventional surface blasts, described in the previous section, there are other types of blasts in large mines. These are designed on somewhat different criteria. This section deals with such blasts.

26.8.1

Rip rap production blasting

As said earlier, the objective of rip rap blasting is to get large pieces of rock. The following should be noted in the design of a blast in large quarries meant for the purpose. Relevant blasthole alignment is shown in Figure 26.23. 1 The rock mass should be as monolithic as possible. A heavily fractured rock mass is almost useless for the purpose as it gives smaller rock pieces. 2 The rock should be as dense as possible. 3 The diameter of blastholes is around 100 mm. 4 Bench height should be between 15 to 20 m. 5 Blastholes should be inclined at an angle 5° or 10° with the vertical. 6 Burden should be about 35 * D to 40 * D. 7 Spacing should be 1.5 * B to 2.0 * B. 8 Subdrilling should be 10 * D 9 Powder factor to be used in case of bottom charge should be: For rocks with UCS > 100 MPa PF > 650 g/m3 For rocks with UCS < 100 MPa PF < 500 g/m3 10 Charge density in the plane of cut should be: For rocks with UCS > 100 MPa CD > 500 g/m2 For rocks with UCS < 100 MPa CD < 250 g/m2 11 The whole row of the blastholes should have instantaneous charge. Fragment distribution obtained in homogeneous rocks by using above blast design parameters has been found to be as given in Table 26.16.

26.8.2

Snake hole blasting

Initially snake hole blasting was thought of as an excellent means of achieving very good fragmentation. However, soon it was noticed that the disadvantages experienced were far too great to be offset by the advantages.

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Table 26.16 Fragment size distribution achieved in rip rap blasting. Percentage in the fragmented rock Fragment weight in kg

Rock with UCS < 100 MPa

Rock with UCS > 100 MPa

>3000 kg 1000–3000 kg 50–1000 kg Finer Than 50 kg

30 20 25 25

50 25 25 10

B

β

H S Snake Holes 0.5+B

S/2

L

Figure 26.24 Snake hole blasting.

The main disadvantages were as under: 1 2 3 4 5 6

Very specialized drills are required for drilling horizontal blastholes. Drilling horizontal blastholes without significant deviation is rather difficult. Penetration rate of drilling horizontal blastholes is very low. Horizontal drilling activity hampers the movement of other pieces of equipment to a large degree. Working at two levels (on bench top and bench floor) is quite cumbersome. Flyrock havoc caused by blast in the horizontal blastholes is excessively high and cannot be easily controlled.

Today, snake hole blasting is practiced in very few instances. Naturally, there has neither been much accumulated experience nor significant research to arrive at design criteria. Snake blastholes are shown in Figure 26.24. So far the practice followed in the field has been: 1 2

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The horizontal blastholes are of diameter 89 to 110 mm, so drilling by small size top hammer or DTH hammer drill is possible. Vertical blastholes are usually at an angle of 5° to 15°.

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3

4 5 6


Vertical blastholes are drilled to such a depth that the distance between the bottom of the vertical blastholes and the horizontal blastholes, which in other words is the burden Bh of the horizontal holes, is (0.5 + B) m, where B is the burden for vertical blastholes. Spacing of the vertical blastholes is as per the normal criteria of S = Ks * B, where Ks equals 1.0 to 1.5. The spacing Sh of horizontal blastholes is 0.5 * S, where S is the spacing of vertical blastholes. Horizontal blastholes have length L (i.e. depth) equal to n * B/cos β, where β is the angle of inclination of vertical blastholes.

For the following aspects, the field practice is not known but author’s recommendations are: a b c

Stemming length for vertical blastholes should be 0.8 * B to 1.0 * B. Stemming length of horizontal blastholes should be 1.0 * Bh. Explosive used in vertical blastholes can be ANFO but in the horizontal blastholes explosive with higher strength- such as ALANFO- should be used.

26.8.3

Cast blasting

In the early 1980s an idea of using explosive energy for throwing the fragmented rockmass to the adjacent waste heap simultaneously through the same blast was born. Over the last decades the techniques of such use have been refined greatly. For a successful cast blast the following parameters of the bench, pit, wall etc. should be practiced. The parameters are shown in Figure 26.25. Bench height should be more than 12 m. Pit width should be 1.25 times the height of the bench face. Presplitting is essential because it enables the burden for the subsequent blast round to be the same for all the blastholes in the first row. Presplitting blastholes should be 251 mm dia., with spacing of 3 m in soft or tough rocks, and 5 to 6 m in hard and brittle rocks. The type of formation matters in choosing the quantity of explosive. Since soft rocks fracture quickly and the gases formed in the explosion escape early, the throw in the soft rock masses is low. To compensate for this the quantity of explosive Near vertical free face formed

Waste Heap

Bench Height Pit Width

Figure 26.25 Parameters in cast blasting.

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required in blasting soft rock is rather high. Hard rock masses require less quantity of explosive. Burden also greatly depends upon geological factors of the rock mass. It is usually determined from a nomograph on the basis of total charge weight in a blasthole and rock blastability. The explosive selected must be able to give sufficient ejection velocity so the broken fragments are thrown in the appropriate direction to the appropriate distance. The ejection velocity can be calculated from the following empirical equation: Ve = 1.14 * (B/(0.07853 * d2 * ρe * PAP))−1.17 where Ve = Ejection velocity in m/s B = Burden in m d = Charge diameter in cm ρe = Density of explosive g/cm2 PAP = Absolute weight strength in cal/g Minimum ejection velocity is about 15 m/s. Spacing should be about 1.3 to 1.6 times burden. The blasthole layout pattern should be rectangular staggered. Stemming should be 20 to 25 times blasthole diameter. Subdrilling should be low i.e. 4 to 6 times blasthole diameter. This is essential so as to ensure that the coal layer is not broken, mixed and thrown with the overburden. Powder factor varies between 0.3 to 0.8 kg/m. All the blastholes in a row should be fired simultaneously without any delay so long as only one face is involved. If two faces are involved then the firing sequence should be as shown in Figure 26.26. In order to ensure that material fragmented and thrown from the first row blast does not infringe with the material fragmented and thrown by the second row blast,

All the blastholes on the same dotted line are to be blasted simultaneously.

Crest

Bench Top

Toe

Minor Face

Bench Floor Blast Initiation

Major Face

Figure 26.26 Firing sequence for cast blasting with two faces.

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Range DP in m

Powder Factor CE in kg/m3 1.250

35

Strain Energy Factor FE 4.4 4.2

1.125 30

4.0 1.000 3.8

25 0.875

3.6

20

0.750 3.4 0.625

15

3.2 0.500

10

3.0 0.375 2.8 0.250 2.6

Figure 26.27 Nomograph 1.

the delay timing between rows should be worked out on the basis of 20 to 25 ms/m of burden. A company, named D’Appolonia Consulting Engineers, has developed nomographs to evaluate some of the variables used in cast blasting. The nomographs are shown in Figure 26.27 to Figure 26.30 and their use is explained through Example 2.

Example 2 Cast blasting method is to be adopted in an existing coal mine. The following are the details of the existing method. Blasthole diameter −250 mm Strain energy factor FE −3.4 Bench height −13.5 m Throw distance DP −16 m Density of explosive −0.85 kg/L Determine other parameters of the cast blast.

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Design of a surface blast Blasthole Diameter D in mm 300 Blasthole Loading Factor qt kg/m Total Charge Weight Per Blasthole Qt in kg 250 105 1575 225 75 1125 200 900 675 175 45 450 360 30 150 Explosive 270 225 Density in g/cc 125 15 12 100

9 6

75

4.5 3

180 135 90 67.5 45 36 27 18 13.5 9

635

Blasthole Loading Length in m

1.40

15 13.5 12 10.5 9

1.30 1.20 1.15 1.10 1.05 1.00 0.95 0.90 0.85

7.5

0.80

4.5

6

3


Solution 2 The stepwise solution by using D’Appolonia Method is as under. Step 1 – In nomograph 1, draw a straight line from FE = 3.4 to DP 16 m. This line intersects the powder factor (CE) line to give a powder factor value of 0.67 kg/m3. Step 2 – Using nomograph 2, draw a straight line from D = 250 to ρc = 0.85 kg/m3. This line intersects the blasthole loading factor qt as 40 kg/m. Step 3 – Calculate C1 and C2 by assuming K1 and K2 to be equal to 1. This actually means that burden and spacing are both equal to each other. C1 = (10.66 * q1)/CE * K2 = 10.66 * 40/(0.67 * 1) = 636.42 C2 = 0.3 * K1 * 636.42/H = 0.3 * 1 * 636.42/13.5 = 14.14 Step 4 – By using the above values of C1 and C2 find out the value of C3 from nomograph 3.

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C3

C2

13000

1600

C'3 12000

80

75

20 11000

1400

1200 70

10000 8000

1000

6000 5000

7000

4000 600

6

40

4.5 50

1.5 60

400

2000

3000 70 200

30 1000 20 10

30

3000

50

40

9

3

800

5000

Burden B in m 10.5

7.5

9000

65 60

C'2 10

C1

1000 80

0


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637

Total Charge Weight per Blasthole Qt in kg 1360 900 675 540 450

Optimum Burden in m

Blastability Factor FV 2.76 2.70

10.5 9.0

2.60

7.5 FE

FV

2.50

2.0

1.9

2.40

2.2

2.04

2.4

2.18

2.6

2.32

2.8

2.46

3.0

2.60

3.2

2.73

31.5

3.4

2.70

22.5 18

3.6

2.57

3.8

2.43

13.5

4.0

2.30

4.2

2.17

4.4

2.03

315 225

6.0

180 135

4.5

90 67.5 45

9

3.0

1.50

2.30 2.20 2.10 2.00 1.90


Step 5 – Let C3 = C'3 and C2 = C'3. Use these values to find out the burden B from nomograph 3. With this we get value of burden B as 5.8 m. Step 6 – The length of charge in the blasthole works out to l = H − K1 * B = 13.5−1 * 5.8 = 7.7 m Step 7 – Using nomograph 2 again, determine Qt from the values of qt and l. In this case since qt = 40 kg/m and l = 7.7 m. Total charge weight Qt works out to 320 kg. Step 8 – Since FE = 3.4, from the table in nomograph 4 the value of FV can be found out as 2.7. Now, using nomograph 4 with values of Qt and FV to be 320 kg and 2.7 respectively the optimum burden Bo can be found out to be 7.2 m. Step 9 – If the values of B and Bo are nearly equal the use of nomographs is over and stemming length T as well as spacing S can be determined by the equations S = K2 * B and T = K1 * B. Step 10 – In this case B and Bo are not nearly equal. For such cases the calculations are required to be repeated from step 3 by using different values of K1 and K2 and recalculating C1 and C2.

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To avoid blind guesswork about K1 and K2, a rule of thumb has been suggested by D’Appolonia. The rule is to use the relation K2 = K13 for estimation. It can be seen that with a value of K1 = 0.86 and K2 = 0.636, C1 and C2 work out to 1043.6 and 19.95 respectively. For these values B and Qt works out to 7.3 m and 301.3 kg. With this value Bo also works out to 7.3 m. Thus, Spacing S = 0.636 * 7.3 = 4.6428 m. Similarly T = 0.86 * 7.3 = 6.28. Since the procedure described above is based on nomographs backed by sound mathematics, the whole procedure can be converted into a computer program. With such a program it is possible to get quick and more precise results.

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Appendix 1

Properties of atmospheric air at high altitudes

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Table A1.1 Properties of atmospheric air at high altitudes. Atmospheric pressure

Altitude

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m

ft

kPa

psi

−400 −200

−1312.34 −656.17

0 200 400 600 800 1000 1200 1400 1600 1800 2000 2200 2400 2600 2800 3000 3200 3400 3600 3800 4000 4200 4400 4600 4800 5000

0.00 656.17 1312.34 1968.50 2624.67 3280.84 3937.01 4593.18 5249.34 5905.51 6561.68 7217.85 7874.02 8530.18 9186.35 9842.52 10498.69 11154.86 11811.02 12467.19 13123.36 13779.53 14435.70 15091.86 15748.03 16404.20

106.223 103.751 101.325 98.945 96.611 94.322 92.076 89.875 87.716 85.599 83.524 81.489 79.495 77.541 75.626 73.749 71.910 70.109 68.344 66.615 64.922 63.264 61.640 60.050 58.494 56.971 55.479 54.020

15.406 15.048 14.696 14.351 14.012 13.680 13.355 13.035 12.722 12.415 12.114 11.819 11.530 11.246 10.969 10.696 10.430 10.168 9.912 9.662 9.416 9.176 8.940 8.710 8.484 8.263 8.047 7.835

Atmospheric temperature

Gravitational acceleration

°C

°F

m/s2

63.68 61.34 59.00 56.66 54.32 51.98 49.64 47.30 44.96 42.62 40.28 37.94 35.60 33.26 30.92 28.58 26.24 23.90 21.56 19.22 16.88 14.54 12.20 9.86 7.52 5.18 2.84 0.50

Decreases Below Surface of Earth 9.80665 32.1740 9.80603 32.1720 9.80542 32.1700 9.80480 32.1680 9.80419 32.1660 9.80357 32.1639 9.80295 32.1619 9.80234 32.1599 9.80172 32.1579 9.80111 32.1559 9.80049 32.1538 9.79988 32.1518 9.79926 32.1498 9.79865 32.1478 9.79803 32.1458 9.79742 32.1438 9.79680 32.1417 9.79619 32.1397 9.79557 32.1377 9.79496 32.1357 9.79434 32.1337 9.79373 32.1316 9.79311 32.1296 9.79250 32.1276 9.79188 32.1256 9.79127 32.1236

17.60 16.30 15.00 13.70 12.40 11.10 9.80 8.50 7.20 5.90 4.60 3.30 2.00 0.70 −0.60 −1.90 −3.20 −4.50 −5.80 −7.10 −8.40 −9.70 −11.00 −12.30 −13.60 −14.90 −16.20 −17.50

Density ft/sec2

kg/m3 1.2727 1.2487 1.2250 1.2017 1.1786 1.1560 1.1336 1.1116 1.0900 1.0686 1.0476 1.0269 1.0065 0.9864 0.9666 0.9472 0.9280 0.9091 0.8905 0.8723 0.8543 0.8366 0.8191 0.8020 0.7851 0.7685 0.7522 0.7361

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lb/ft3

Barometer mm

Water column m

0.0795 0.0780 0.0765 0.0750 0.0736 0.0722 0.0708 0.0694 0.0680 0.0667 0.0654 0.0641 0.0628 0.0616 0.0603 0.0591 0.0579 0.0568 0.0556 0.0545 0.0533 0.0522 0.0511 0.0501 0.0490 0.0480 0.0470 0.0460

796.741 778.195 760.000 742.151 724.643 707.471 690.630 674.115 657.921 642.044 626.478 611.219 596.263 581.604 567.239 553.162 539.370 525.857 512.620 499.654 486.954 474.518 462.340 450.416 438.742 427.315 416.129 405.182

10.831 10.579 10.332 10.089 9.851 9.618 9.389 9.164 8.944 8.728 8.517 8.309 8.106 7.907 7.711 7.520 7.333 7.149 6.969 6.793 6.620 6.451 6.285 6.123 5.965 5.809 5.657 5.508

Dynamic viscosity kg/(m . s)

1.789 * 10−5 1.783 * 10−5 1.777 * 10−5 1.771 * 10−5 1.764 * 10−5 1.758 * 10−5 1.752 * 10−5 1.745 * 10−5 1.739 * 10−5 1.732 * 10−5 1.726 * 10−5 1.720 * 10−5 1.713 * 10−5 1.707 * 10−5 1.700 * 10−5 1.694 * 10−5 1.687 * 10−5 1.681 * 10−5 1.674 * 10−5 1.668 * 10−5 1.661 * 10−5 1.655 * 10−5 1.648 * 10−5 1.642 * 10−5 1.635 * 10−5 1.628 * 10−5

Thermal conductivity W/(m . K)

0.0253 0.0252 0.0252 0.0251 0.0250 0.0249 0.0248 0.0247 0.0245 0.0244 0.0243 0.0242 0.0241 0.0240 0.0239 0.0238 0.0237 0.0236 0.0235 0.0234 0.0233 0.0232 0.0231 0.0230 0.0229 0.0228

Speed of sound m/s

ft/sec

341.826 341.061 340.294 339.526 338.755 337.983 337.210 336.434 335.657 334.878 334.097 333.314 332.529 331.743 330.955 330.164 329.372 328.578 327.782 326.985 326.185 325.383 324.579 323.773 322.966 322.156 321.344 320.530

1121.476 1118.966 1116.450 1113.929 1111.402 1108.869 1106.331 1103.786 1101.236 1098.680 1096.118 1093.550 1090.976 1088.396 1085.809 1083.217 1080.618 1078.013 1075.402 1072.784 1070.160 1067.529 1064.892 1062.249 1059.599 1056.942 1054.278 1051.608

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Appendix 2

Air properties at various temperatures

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Table A2.1 Properties of air at various temperatures.

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Temperature °C

Density kg/m3

−150

2.866

−100 −50 −40 −30 −20 −10 0 5 10 15 20 25 30 35 40 45 50 60 70 80 90 100 120 140 160 180 200 250 300 350 400

2.038 1.582 1.514 1.451 1.394 1.341 1.292 1.269 1.246 1.225 1.204 1.184 1.164 1.145 1.127 1.109 1.092 1.059 1.028 0.9994 0.9718 0.9458 0.8977 0.8542 0.8148 0.7788 0.7459 0.6746 0.6158 0.5664 0.5243

Specific Heat capacity Cp KJ/(kg . K)

Thermal conductivity W/(m . K)

Thermal diffusivity m2/s

983

0.01171

4.158 * 10−6

966 999 1002 1004 1005 1006 1006 1006 1006 1007 1007 1007 1007 1007 1007 1007 1007 1007 1007 1008 1008 1009 1011 1013 1016 1019 1023 1033 1044 1056 1069

0.01582 0.01979 0.02057 0.02134 0.02211 0.02288 0.02364 0.02401 0.02439 0.02476 0.02514 0.02551 0.02588 0.02625 0.02662 0.02699 0.02735 0.02808 0.02881 0.02953 0.03024 0.03095 0.03235 0.03374 0.03511 0.03646 0.03779 0.04104 0.04418 0.04721 0.05015

8.036 * 10−6 1.252 * 10−5 1.356 * 10−5 1.465 * 10−5 1.578 * 10−5 1.696 * 10−5 1.818 * 10−5 1.880 * 10−5 1.944 * 10−5 2.009 * 10−5 2.074 * 10−5 2.141 * 10−5 2.208 * 10−5 2.277 * 10−5 2.346 * 10−5 2.416 * 10−5 2.487 * 10−5 2.632 * 10−5 2.780 * 10−5 2.931 * 10−5 3.086 * 10−5 3.243 * 10−5 3.565 * 10−5 3.898 * 10−5 4.241 * 10−5 4.593 * 10−5 4.954 * 10−5 5.890 * 10−5 6.871 * 10−5 7.892 * 10−5 8.951 * 10−5

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Dynamic viscosity μ kg/(m . s)

Kinematics viscosity ν m2/s

8.636 * 10−6 −5

1.189 * 10 1.474 * 10−5 1.527 * 10−5 1.579 * 10−5 1.630 * 10−5 1.680 * 10−5 1.729 * 10−5 1.754 * 10−5 1.778 * 10−5 1.802 * 10−5 1.825 * 10−5 1.849 * 10−5 1.872 * 10−5 1.895 * 10−5 1.918 * 10−5 1.941 * 10−5 1.963 * 10−5 2.008 * 10−5 2.052 * 10−5 2.096 * 10−5 2.139 * 10−5 2.181 * 10−5 2.264 * 10−5 2.345 * 10−5 2.420 * 10−5 2.504 * 10−5 2.577 * 10−5 2.760 * 10−5 2.934 * 10−5 3.101 * 10−5 3.261 * 10−5

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Prandtl number Pr

Expansion coefficient 1/K × 103

3.013 * 10−6

0.7246

8.21

−6

0.7263 0.7440 0.7436 0.7425 0.7408 0.7387 0.7362 0.7350 0.7336 0.7323 0.7309 0.7296 0.7282 0.7268 0.7255 0.7241 0.7228 0.7202 0.7177 0.7154 0.7132 0.7111 0.7073 0.7041 0.7014 0.6992 0.6974 0.6946 0.6935 0.6937 0.6948

5.82 4.51

5.837 * 10 9.319 * 10−6 1.008 * 10−5 1.087 * 10−5 1.169 * 10−5 1.252 * 10−5 1.338 * 10−5 1.382 * 10−5 1.426 * 10−5 1.470 * 10−5 1.516 * 10−5 1.562 * 10−5 1.608 * 10−5 1.655 * 10−5 1.702 * 10−5 1.750 * 10−5 1.798 * 10−5 1.896 * 10−5 1.995 * 10−5 2.097 * 10−5 2.201 * 10−5 2.306 * 10−5 2.522 * 10−5 2.745 * 10−5 2.975 * 10−5 3.212 * 10−5 3.455 * 10−5 4.091 * 10−5 4.765 * 10−5 5.475 * 10−5 6.219 * 10−5

3.67

3.43

3.20

3.00 2.83 2.68 2.55 2.43 2.32 2.21 2.11 1.91 1.75 1.61 1.49

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Appendix 3

Bailing velocities in blasthole annulus

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Table A3.1 Velocity of air flowing up between the blasthole and drill pipe. Hole diameter D

Drill pipe diameter d

Annular area

Value 0.833D

Value 0.792D

Annulus/ hole area

Volume of compressed air (m3/min) required for attaining the bailing velocities (m/min) given here below

Inch

mm

Inch

Sq. In

sq. m

in mm

in mm

%

1000

6 6 6 6.25 6.25 6.25 6.75 6.75 6.75 6.75 7.375 7.375 7.375 7.875 7.875 7.875 7.875 9 9 9 9 9.875 9.875 9.875

152.40 152.40 152.40 158.75 158.75 158.75 171.45 171.45 171.45 171.45 187.32 187.32 187.32 200.02 200.02 200.02 200.02 228.60 228.60 228.60 228.60 250.82 250.82 250.82

12034 10134 7981 11686 9532 7126 14980 12826 10419 7759 14892 12232 7767 18756 16096 11631 8847 18467 16215 14409 10609 24583 22778 18978

126.95 126.95 126.95 132.24 132.24 132.24 142.82 142.82 142.82 142.82 156.04 156.04 156.04 166.62 166.62 166.62 166.62 190.42 190.42 190.42 190.42 208.94 208.94 208.94

120.70 120.70 120.70 125.73 125.73 125.73 135.79 135.79 135.79 135.79 148.36 148.36 148.36 158.42 158.42 158.42 158.42 181.05 181.05 181.05 181.05 198.65 198.65 198.65

65.97 55.56 43.75 59.04 48.16 36.00 64.88 55.56 45.13 33.61 54.04 44.38 28.18 59.69 51.22 37.01 28.15 44.99 39.51 35.11 25.85 49.75 46.10 38.41

12.03 10.13 7.98 11.69 9.53 7.13 14.98 12.83 10.42 7.76 14.89 12.23 7.77 18.76 16.10 11.63 8.85 18.47 16.21 14.41 10.61 24.58 22.78 18.98

3.5 4 4.5 4 4.5 5 4 4.5 5 5.5 5 5.5 6.25 5 5.5 6.25 6.675 6.675 7 7.25 7.75 7 7.25 7.75

mm 88.90 101.60 114.30 101.60 114.30 127.00 101.60 114.30 127.00 139.70 127.00 139.70 158.75 127.00 139.70 158.75 169.54 169.54 177.80 184.15 196.85 177.80 184.15 196.85

18.653 15.708 12.370 18.113 14.775 11.045 23.218 19.880 16.150 12.026 23.083 18.960 12.039 29.072 24.949 18.027 13.713 28.623 25.133 22.335 16.444 38.104 35.306 29.416

1300 15.64 13.17 10.37 15.19 12.39 9.26 19.47 16.67 13.54 10.09 19.36 15.90 10.10 24.38 20.92 15.12 11.50 24.01 21.08 18.73 13.79 31.96 29.61 24.67

1600 19.25 16.21 12.77 18.70 15.25 11.40 23.97 20.52 16.67 12.41 23.83 19.57 12.43 30.01 25.75 18.61 14.16 29.55 25.94 23.06 16.97 39.33 36.44 30.36

1900 22.87 19.25 15.16 22.20 18.11 13.54 28.46 24.37 19.80 14.74 28.30 23.24 14.76 35.64 30.58 22.10 16.81 35.09 30.81 27.38 20.16 46.71 43.28 36.06

2200 26.48 22.30 17.56 25.71 20.97 15.68 32.95 28.22 22.92 17.07 32.76 26.91 17.09 41.26 35.41 25.59 19.46 40.63 35.67 31.70 23.34 54.08 50.11 41.75

2500 30.09 25.34 19.95 29.21 23.83 17.81 37.45 32.07 26.05 19.40 37.23 30.58 19.42 46.89 40.24 29.08 22.12 46.17 40.54 36.02 26.52 61.46 56.95 47.44

2700 32.49 27.36 21.55 31.55 25.74 19.24 40.44 34.63 28.13 20.95 40.21 33.03 20.97 50.64 43.46 31.40 23.89 49.86 43.78 38.91 28.64 66.37 61.50 51.24

3000 36.10 30.40 23.94 35.06 28.60 21.38 44.94 38.48 31.26 23.28 44.68 36.70 23.30 56.27 48.29 34.89 26.54 55.40 48.64 43.23 31.83 73.75 68.33 56.93

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Book.indb 649

10.625 10.625 10.625 11 11 11 11 12.25 12.25 12.25 12.25 13.75 13.75 13.75 15 15 15 15 16 16 16 16 16 17.5 17.5 17.5 17.5 17.5 17.5

269.88 269.88 269.88 279.40 279.40 279.40 279.40 311.15 311.15 311.15 311.15 349.25 349.25 349.25 381.00 381.00 381.00 381.00 406.40 406.40 406.40 406.40 406.40 444.50 444.50 444.50 444.50 444.50 444.50

7.75 8.625 9 7.75 8.625 9 9.25 9 9.25 10 10.75 10 10.75 12 10.75 12 12.25 13 12 12.25 13.375 13.625 13.75 12.25 13 13.625 13.75 15 16

196.85 219.07 228.60 196.85 219.07 228.60 234.95 228.60 234.95 254.00 273.05 254.00 273.05 304.80 273.05 304.80 311.15 330.20 304.80 311.15 339.72 346.07 349.25 311.15 330.20 346.07 349.25 381.00 406.40

41.491 30.238 25.047 47.860 36.607 31.416 27.833 54.242 50.658 39.319 27.096 69.950 57.727 35.392 85.952 63.617 58.856 43.982 87.965 83.203 60.562 55.260 52.573 122.669 107.796 94.726 92.039 63.814 39.466

26768 19508 16159 30877 23617 20268 17956 34994 32683 25367 17481 45129 37243 22834 55453 41043 37971 28376 56751 53679 39072 35652 33918 79141 69546 61114 59380 41170 25462

224.81 224.81 224.81 232.74 232.74 232.74 232.74 259.19 259.19 259.19 259.19 290.93 290.93 290.93 317.37 317.37 317.37 317.37 338.53 338.53 338.53 338.53 338.53 370.27 370.27 370.27 370.27 370.27 370.27

213.74 213.74 213.74 221.28 221.28 221.28 221.28 246.43 246.43 246.43 246.43 276.61 276.61 276.61 301.75 301.75 301.75 301.75 321.87 321.87 321.87 321.87 321.87 352.04 352.04 352.04 352.04 352.04 352.04

46.80 34.10 28.25 50.36 38.52 33.06 29.29 46.02 42.98 33.36 22.99 47.11 38.88 23.83 48.64 36.00 33.31 24.89 43.75 41.38 30.12 27.48 26.15 51.00 44.82 39.38 38.27 26.53 16.41

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Gray colored rows indicate the most usually used combinations of blasthole and drill pipe diameters.

26.77 19.51 16.16 30.88 23.62 20.27 17.96 34.99 32.68 25.37 17.48 45.13 37.24 22.83 55.45 41.04 37.97 28.38 56.75 53.68 39.07 35.65 33.92 79.14 69.55 61.11 59.38 41.17 25.46

34.80 25.36 21.01 40.14 30.70 26.35 23.34 45.49 42.49 32.98 22.73 58.67 48.42 29.68 72.09 53.36 49.36 36.89 73.78 69.78 50.79 46.35 44.09 102.88 90.41 79.45 77.19 53.52 33.10

42.83 31.21 25.85 49.40 37.79 32.43 28.73 55.99 52.29 40.59 27.97 72.21 59.59 36.53 88.72 65.67 60.75 45.40 90.80 85.89 62.52 57.04 54.27 126.63 111.27 97.78 95.01 65.87 40.74

50.86 37.07 30.70 58.67 44.87 38.51 34.12 66.49 62.10 48.20 33.21 85.74 70.76 43.38 105.36 77.98 72.15 53.91 107.83 101.99 74.24 67.74 64.44 150.37 132.14 116.12 112.82 78.22 48.38

58.89 42.92 35.55 67.93 51.96 44.59 39.50 76.99 71.90 55.81 38.46 99.28 81.93 50.23 122.00 90.30 83.54 62.43 124.85 118.09 85.96 78.43 74.62 174.11 153.00 134.45 130.64 90.57 56.02

66.92 48.77 40.40 77.19 59.04 50.67 44.89 87.49 81.71 63.42 43.70 112.82 93.11 57.08 138.63 102.61 94.93 70.94 141.88 134.20 97.68 89.13 84.79 197.85 173.86 152.78 148.45 102.92 63.66

72.27 52.67 43.63 83.37 63.77 54.72 48.48 94.49 88.24 68.49 47.20 121.85 100.56 61.65 149.72 110.82 102.52 76.61 153.23 144.93 105.49 96.26 91.58 213.68 187.77 165.01 160.33 111.16 68.75

80.31 58.52 48.48 92.63 70.85 60.80 53.87 104.98 98.05 76.10 52.44 135.39 111.73 68.50 166.36 123.13 113.91 85.13 170.25 161.04 117.22 106.95 101.75 237.42 208.64 183.34 178.14 123.51 76.39

Appendix 4

Air pressure loss in steel pipes

All the necessary information for calculating pressure drop in seamless carbon steel pipes, pipe fittings etc. has been given in chapter 8 of this book. This appendix presents Tables A4.1a and A4.1b, that give pressure drop values for dry compressed air flowing through steel pipes in straight condition. The following conditions have been presumed for calculating pressure drop values given in the tables. Compressed Air Gauge Pressure = 1034 kPa Length of the pipe = 100 m Atmospheric Air Pressure (Abs) = 101325 Pa Atmospheric Temperature = 15°C Dynamic Viscosity of Air = 1.802 kg/(m . s) Absolute Roughness of Pipe = 0.00015 m Gas Constant for Air = 287.1 J/(kg . K) The discharge ratings given in first column on the left, are so chosen that they very closely match with the discharge ratings of compressors mounted on the rotary blasthole drills. Pressure drop values given in Table A4.1a pertain to schedule 40 and schedule 80 pipes as per ANSI Standard. Similarly, the pressure drop values given in Table A4.1b pertain to steel pipes made to DIN Standard. As most of the manufacturers of rotary blasthole drills are located in the United States of America, the original values of compressor pressure and discharge ratings as well as the diameters of the steel tubes are in imperial units but the tables show metric units. For ease in finding equivalent values the compressor discharge ratings have been given in cfm units in column no. 2 of Table A4.1a. Similarly, the nominal diameters of the steel tubes have been given in row no. 4 of Table A4.1a. Pressure drop values given in the tables have been calculated by using Darcy Weisbach equation. The first example in chapter 8 gives all the steps used in the calculations, except for the friction factor. In those examples the friction factor has been found by the use of Moody’s Chart, whereas for calculation of the values in Table A4.1a and A4.1b Haaland’s formula has been used. For actual calculations a spreadsheet program has been used. A complete spreadsheet formulation has been presented in Figure A4.1.

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652

Appendix 4

The main reason for use of Haaland’s formula is that it gives pressure drop values slightly higher than actual. The inaccuracy, therefore, will be on the safer side. Where pressure drop value works out to more than 2% of the compressed air gauge pressure, the pipe is considered unsuitable and instead of giving the pressure drop value the word “Unsuit” is shown. It is to be specifically noted that the pressure drop is inversely proportional to the compression ratio. To facilitate calculation of the pressure drop for gauge pressures other than 1034 kPa and/or atmospheric pressure other than 101.325 kPa, values given in Table A4.2 can be used. It may be noted that steel pipes of all the sizes mentioned in Table A4.1a and A4.1b are available in pressure ratings well above 1500 kPa. EXAMPLE

Find the pressure drop likely to occur in a galvanized schedule 80 pipe of nominal size 4" and length 57 m, when the compressed air flow has a volume of 25.485 m3/min and the gauge pressure of 448000 Pa. The altitude of the place is 3000 m. SOLUTION

Referring to the Table A4.1a we find that pressure loss for Nominal 4" pipe of schedule 80 for compressed air flow of 25.485 m3/min is 2.7720 kPa. This pressure loss is for gauge pressure of 1034 kPa and atmospheric pressure of 101.325 kPa that prevails at an altitude of 0 i.e. sea level. This must be adjusted to compressed air gauge pressure of 448000 Pa i.e. 448 kPa and atmospheric pressure of 70.109 kPa. From Table A4.2 we find that the pressure drop factor for 448 kPa and 70.109 kPa is 2.1913. Hence actual pressure drop for compressed air of gauge pressure of 448 kPa at an altitude of 3000 m will be 2.7720 * 2.1913, which equals to 6.0743 kPa.

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Table A4.1a Pressure loss for compressed air pressure of 1034 kPa in schedule 40 and schedule 80 seamless carbon steel pipes.

Flow volume m3/min

cfm

23.361 25.485 26.901 29.733 32.564 33.98 35.396 45.307 53.802 56.634 73.624 75.04 79.287 84.951 103.356 107.604

825 900 950 1050 1150 1200 1250 1600 1900 2000 2600 2650 2800 3000 3650 3800

Pressure loss in kpa for following nominal sizes of schedule 40 seamless carbon steel pipes

Pressure loss in kpa following nominal sizes of schedule 80 seamless carbon steel pipes

3"

3"

3.5"

4"

5"

6"

8"

10"

Actual internal diameter in mm 77.93 7.3817 8.7710 9.7637 11.9085 14.2652 15.5237 16.8354 Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit

3.5"

4"

5"

6"

8"

10"

122.25

146.33

193.68 247.65

0.2814 0.3333 0.3703 0.4502 0.5379 0.5846 0.6333 1.0285 1.4429 1.5966 2.6816 2.7847 3.1054 3.5603 5.2528 5.6901

0.0673 0.0795 0.0882 0.1070 0.1275 0.1385 0.1499 0.2422 0.3387 0.3745 0.6265 0.6504 0.7248 0.8303 1.2224 1.3236

Actual internal diameter in mm

90.12

102.26

128.19

154.05

202.72

254.51 73.66

84.45

97.18

3.45750 4.10624 4.56968 5.57081 6.67055 7.25773 7.86967 12.84569 18.07587 20.01747 Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit

1.7915 2.1266 2.3659 2.8827 3.4502 3.7532 4.0689 6.6351 9.3314 10.3321 17.4054 18.0777 20.1705 Unsuit Unsuit Unsuit

0.5560 0.6592 0.7329 0.8919 1.0663 1.1594 1.2564 2.0440 2.8706 3.1773 5.3436 5.5494 6.1900 7.0987 10.4810 11.3549

0.2162 0.2559 0.2843 0.3455 0.4127 0.4484 0.4857 0.7882 1.1052 1.2228 2.0525 2.1313 2.3765 2.7243 4.0181 4.3523

0.0534 0.0631 0.0700 0.0848 0.1010 0.1097 0.1187 0.1916 0.2679 0.2961 0.4950 0.5139 0.5726 0.6558 0.9650 1.0448

0.0170 0.0200 0.0222 0.0268 0.0319 0.0346 0.0374 0.0600 0.0836 0.0923 0.1536 0.1594 0.1775 0.2030 0.2980 0.3224

4.5629 5.4201 6.0325 7.3556 8.8090 9.5852 10.3940 16.9722 Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit

2.3348 0.7104 2.7720 0.8425 3.0843 0.9367 3.7589 1.1403 4.4998 1.3637 4.8954 1.4829 5.3076 1.6071 8.6587 2.6160 12.1802 3.6752 13.4873 4.0682 Unsuit 6.8446 Unsuit 7.1084 Unsuit 7.9295 Unsuit 9.0944 Unsuit 13.4305 Unsuit 14.5510

9.9116 11.7788 13.1131 15.9963 19.1646 Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit Unsuit

0.0195 0.0230 0.0254 0.0308 0.0366 0.0397 0.0429 0.0689 0.0961 0.1061 0.1767 0.1834 0.2042 0.2336 0.3430 0.3712

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Book.indb 654

Table A4.1b Pressure loss for compressed air pressure of 1034 kPa in DN seamless carbon steel pipes. Flow Volume 3

m /min

Pressure loss in kpa for following diameters of seamless carbon steel pipes

cfm

80 mm

90 mm

100 mm

110 mm

120 mm

130 mm

140 mm 150 mm 160 mm 180 mm 190 mm 200 mm 225 mm 250 mm

23.361

825

6.43670

3.48158

2.01213

1.22690

0.78196

0.51725 0.35315 0.24778 0.17803 0.09760 0.07414 0.05716 0.03153 0.01857

25.485

900

7.64753

4.13487

2.38866

1.45584

0.92744

0.61319 0.41845 0.29345 0.21074 0.11542 0.08764 0.06753 0.03721 0.02189

26.901

950

8.51266

4.60155

2.65760

1.61932

1.03130

0.68167 0.46505 0.32603 0.23407 0.12812 0.09725 0.07492 0.04125 0.02425

29.733

1050 10.38188

5.60970

3.23844

1.97233

1.25552

0.82945 0.56558 0.39631 0.28437 0.15549 0.11797 0.09083 0.04995 0.02933

32.564

1150 12.43563

6.71714

3.87634

2.35992

1.50164

0.99163 0.67587 0.47338 0.33952 0.18547 0.14065 0.10825 0.05946 0.03488

33.98

1200 13.53234

7.30843

4.21688

2.56680

1.63298

1.07816 0.73471 0.51449 0.36892 0.20146 0.15274 0.11753 0.06452 0.03784

35.396

1250 14.67536

7.92466

4.57175

2.78237

1.76982

1.16831 0.79599 0.55730 0.39955 0.21810 0.16532 0.12719 0.06979 0.04091

45.307

1600 Unsuit

12.93557

7.45659

4.53419

2.88152

1.90036 1.29348 0.90469 0.64793 0.35292 0.26724 0.20537 0.11239 0.06571

53.802

1900 Unsuit

18.20245 10.48778

6.37422

4.04873

2.66865 1.81535 1.26892 0.90822 0.49406 0.37387 0.28712 0.15687 0.09155

56.634

2000 Unsuit

20.15767 11.61288

7.05708

4.48182

2.95366 2.00890 1.40398 1.00472 0.54636 0.41337 0.31739 0.17333 0.10111

73.624

2600 Unsuit

Unsuit

19.56562 11.88277

7.54167

4.96681 3.37570 2.35744 1.68572 0.91520 0.69184 0.53075 0.28921 0.16834

75.04

2650 Unsuit

Unsuit

20.32152 12.34138

7.83242

5.15806 3.50553 2.44799 1.75038 0.95020 0.71826 0.55099 0.30019 0.17470

79.287

2800 Unsuit

Unsuit

Unsuit

13.76894

8.73741

5.75334 3.90959 2.72978 1.95159 1.05911 0.80046 0.61395 0.33436 0.19451

84.951

3000 Unsuit

Unsuit

Unsuit

15.79433 10.02130

6.59779 4.48273 3.12946 2.23695 1.21354 0.91700 0.70320 0.38278 0.22256

103.356 3650 Unsuit

Unsuit

Unsuit

Unsuit

14.80050

9.74072 6.61555 4.61651 3.29849 1.78779 1.35029 1.03496 0.56264 0.32671

107.604 3800 Unsuit

Unsuit

Unsuit

Unsuit

16.03547 10.55278 7.16656 5.00064 3.57267 1.93607 1.46215 1.12059 0.60905 0.35356

11/22/2010 2:43:38 PM

Appendix 4

655

Figure A4.1 Screen image of the worksheet (Calculation of compressed air pressure drop when the air flows through a steel pipe). Table A4.2 Pressure drop factor.

Book.indb 655

Altitude in m

Inverse of compression ratio for different gauge pressures of the compressed air Atmospheric pressures pressure in kPa 1034 862 758 689 448 414 379 350

0

101325

1.0000

1.1785

1.3212

1.4365

2.0668

2.2031

2.3637

2.5155

500

95465

1.0052

1.1858

1.3303

1.4473

2.0890

2.2285

2.3929

2.5486

1000

89875

1.0102

1.1927

1.3390

1.4576

2.1108

2.2532

2.4214

2.5810

1500

84552

1.0150

1.1994

1.3475

1.4677

2.1319

2.2772

2.4492

2.6126

2000

79495

1.0196

1.2059

1.3556

1.4773

2.1523

2.3006

2.4762

2.6434

2500

74670

1.0240

1.2121

1.3635

1.4867

2.1722

2.3233

2.5025

2.6734

3000

70109

1.0283

1.2180

1.3710

1.4956

2.1913

2.3452

2.5279

2.7025

3500

65768

1.0323

1.2237

1.3782

1.5042

2.2098

2.3664

2.5526

2.7307

4000

61640

1.0362

1.2292

1.3852

1.5125

2.2277

2.3869

2.5765

2.7581

4500

57722

1.0399

1.2344

1.3918

1.5204

2.2450

2.4068

2.5997

2.7846

5000

54020

1.0435

1.2394

1.3981

1.5280

2.2615

2.4258

2.6219

2.8101

11/22/2010 2:43:39 PM

Appendix 5

Air pressure loss in hose pipes

Table 5.1a and 5.1b in this appendix give pressure loss in compressed air flow in hose pipes in their stretched condition. The method of calculation of pressure loss is the same as that for steel pipes except that Haaland’s equation for friction factor is used. All other conditions, except the absolute roughness of the hose pipe and pipe diameters, are the same. Since modern hose pipes are made from plastic and their inner surface is very smooth, the absolute roughness of these pipes is considered to be 0.000002 m. Hose pipes of all the sizes given in Table 5.1a and 5.1b are available in pressure ratings well above 1500 kPa. Pressure drop factors given in appendix 4 can also be applied to the pressure drop values given in this appendix.

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Book.indb 658

Table A5.1a Pressure loss for compressed air pressure of 1034 kPa in ASTM hose pipes. Flow volume


m3/min

73.03

cfm

76.20

88.90

101.6

114.30

127.00

146.33

152.40

168.28

193.68

202.72

203.20

219.08

254.00

23.361

825

6.04360

4.91769

2.33048

1.22199

0.69202

0.41636 0.21042 0.17304 0.10741 0.05465 0.04390 0.04340 0.03024 0.01487

25.485

900

7.08974

5.76793

2.73184

1.43182

0.81057

0.48755 0.24631 0.20253 0.12569 0.06393 0.05134 0.05076 0.03537 0.01739

26.901

950

7.83005

6.36951

3.01566

1.58014

0.89433

0.53783 0.27165 0.22335 0.13859 0.07048 0.05660 0.05596 0.03898 0.01916

29.733

1050

9.41271

7.65533

3.62191

1.89679

1.07311

0.64512 0.32571 0.26776 0.16610 0.08444 0.06780 0.06704 0.04669 0.02294

32.564

1150

11.12968

9.04993

4.27889

2.23974

1.26663

0.76122 0.38417 0.31579 0.19585 0.09953 0.07991 0.07901 0.05502 0.02702

33.98

1200

12.03868

9.78812

4.62645

2.42108

1.36893

0.82257 0.41506 0.34116 0.21156 0.10749 0.08630 0.08533 0.05941 0.02918

35.396

1250

12.98099 10.55327

4.98656

2.60892

1.47487

0.88610 0.44703 0.36742 0.22782 0.11574 0.09292 0.09187 0.06396 0.03141

45.307

1600

20.49907 16.65524

7.85416

4.10310

2.31685

1.39064 0.70076 0.57580 0.35676 0.18107 0.14533 0.14368 0.09999 0.04906

53.802

1900

Unsuit

Unsuit

10.78784

5.62929

3.17582

1.90488 0.95908 0.78787 0.48791 0.24747 0.19857 0.19632 0.13657 0.06696

56.634

2000

Unsuit

Unsuit

11.86208

6.18765

3.48987

2.09278 1.05340 0.86529 0.53577 0.27168 0.21799 0.21552 0.14991 0.07349

73.624

2600

Unsuit

Unsuit

19.30251 10.04933

5.65933

3.38966 1.70374 1.39897 0.86546 0.43837 0.35162 0.34763 0.24167 0.11836

75.04

2650

Unsuit

Unsuit

19.99900 10.41038

5.86199

3.51071 1.76439 1.44873 0.89618 0.45390 0.36406 0.35993 0.25022 0.12253

79.287

2800

Unsuit

Unsuit

Unsuit

11.52914

6.48973

3.88561 1.95217 1.60278 0.99128 0.50194 0.40257 0.39799 0.27665 0.13545

84.951

3000

Unsuit

Unsuit

Unsuit

13.10449

7.37326

4.41306 2.21625 1.81939 1.12497 0.56946 0.45667 0.45148 0.31378 0.15358

103.356 3650

Unsuit

Unsuit

Unsuit

18.87526 10.60621

6.34136 3.18067 2.61026 1.61278 0.81561 0.65389 0.64644 0.44908 0.21963

107.604 3800

Unsuit

Unsuit

Unsuit

20.34769 11.43034

6.83254 3.42611 2.81148 1.73683 0.87817 0.70400 0.69598 0.48345 0.23640

11/22/2010 2:43:41 PM

Book.indb 659

Table A5.1b Pressure loss for compressed air pressure of 1034 kPa in DN hose pipes. Flow volume


m3/min

80

cfm

90

100

110

120

130

140

150

160

180

190

200

225

250

23.361

825

3.88412

2.19586

1.31939

0.83269 0.54724 0.37205 0.26035 0.18677 0.13691 0.07771 0.05993 0.04684 0.02661 0.01605

25.485

900

4.55480

2.57393

1.54602

0.97544 0.64089 0.43564 0.30479 0.21861 0.16023 0.09092 0.07011 0.05479 0.03111 0.01876

26.901

950

5.02924

2.84126

1.70622

1.07632 0.70707 0.48055 0.33617 0.24110 0.17669 0.10024 0.07729 0.06040 0.03430 0.02068

29.733

1050

6.04310

3.41227

2.04826

1.29164 0.84827 0.57637 0.40312 0.28905 0.21179 0.12012 0.09260 0.07236 0.04107 0.02476

32.564

1150

7.14242

4.03103

2.41873

1.52476 1.00109 0.68005 0.47553 0.34091 0.24975 0.14161 0.10915 0.08528 0.04840 0.02917

33.98

1200

7.72421

4.35835

2.61464

1.64801 1.08187 0.73483 0.51379 0.36831 0.26980 0.15295 0.11789 0.09210 0.05226 0.03149

35.396

1250

8.32716

4.69749

2.81757

1.77565 1.16551 0.79156 0.55340 0.39667 0.29056 0.16470 0.12694 0.09916 0.05626 0.03390

45.307

1600 13.13329

7.39778

4.43198

2.79035 1.83006 1.24203 0.86781 0.62170 0.45518 0.25780 0.19862 0.15511 0.08794 0.05295

53.802

1900 18.05705

10.15987

6.08129

3.82594 2.50772 1.70106 1.18800 0.85075 0.62265 0.35244 0.27147 0.21195 0.12010 0.07229

56.634

2000 19.86140

11.17118

6.68474

4.20463 2.75541 1.86877 1.30495 0.93438 0.68379 0.38698 0.29805 0.23268 0.13183 0.07933

73.624

2600 Unsuit

18.17481

10.85901

6.82164 4.46574 3.02608 2.11148 1.51089 1.10504 0.62476 0.48098 0.37534 0.21248 0.12778

75.04

2650 Unsuit

18.83033

11.24934

7.06617 4.62545 3.13409 2.18672 1.56464 1.14431 0.64691 0.49802 0.38863 0.21998 0.13229

79.287

2800 Unsuit

Unsuit

12.45887

7.82371 5.12012 3.46857 2.41968 1.73108 1.26587 0.71548 0.55075 0.42974 0.24321 0.14623

84.951

3000 Unsuit

Unsuit

14.16215

8.89006 5.81621 3.93912 2.74734 1.96511 1.43676 0.81184 0.62486 0.48751 0.27584 0.16582

103.356 3650 Unsuit

Unsuit

20.40254

12.79336 8.36225 5.65910 3.94433 2.81968 2.06054 1.16333 0.89507 0.69811 0.39472 0.23715

107.604 3800 Unsuit

Unsuit

Unsuit

13.78866 9.01104 6.09714 4.24903 3.03713 2.21921 1.25268 0.96375 0.75162 0.42492 0.25526

11/22/2010 2:43:41 PM

Appendix 6

Air pressure loss in pipe fittings

Pressure loss occurs when a flow enters a flow pipe from a large reservoir. Resistance coefficients depend upon the geometry of the entrance as shown hereunder,

V

When the entrance starts inside the reservoir, as shown in the drawing on the left, the resistance coefficient KL is equal to 0.8 when L ≈ D

D L

Sharp edged entrance. r = 0.

r V

V

D

D

For rounded entrance the resistance coefficient depends upon the roundness as shown in following graph.

Resistance Coefficient

0.5 0.4 0.3 0.2 0.1 0.0 0.0

0.05

0.10

0.15

0.20

0.25

r/D

Figure A6.1 Resistance coefficient at entrance.

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L D

D

D

V

V

V r

Pressure loss also occurs when flow enters a large reservoir from flow pipe. However the resistance coefficients does not depend upon the geometry of the exit. For fully developed laminar flow the resistance coefficient is 2, whereas, for fully developed turbulent flow the resistance coefficient is 1.

Figure A6.2 Resistance coefficient at exit.

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Gradual Contraction

D

θ

V

Gradual Expansion

d

d

V

θ

D

Gradual Expansion 0.8

0.7


0.6 0.5

0.4 30°

0.3

0.2 20° 0.1

0.0 1.0

10° 5° 2.0

3.0

4.0

D/d Gradual Contraction 0.4


150° 0.3 120° 100° 0.2

90 75°

0.1

0.0 1.0

50°–60° 15°–40°

2.0

3.0

D/d

Figure A6.3 Resistance coefficient at gradual contraction and expansion.

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Pressure loss occurs when a the diameter of flow path suddenly increases or decreases.

V

d

D

D

In case of sudden expansion of the flow path, as shown in figure on the left, the resistance coefficient is calculated as KL = (1 – d2/D2)

V

For sudden contraction the pressure loss coefficient can be found from the following graph

d

0.6 0.5 0.4 KL

0.3 0.2 0.1 0.0 0.0 0.1 0.2 0.3 0.4 0.5 0.6 0.7 0.8 0.9 1.0 d2/D2

Figure A6.4 Resistance coefficient for sudden expansion and contraction.

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Appendix 6

665

Tee with straight flow Flanged

: KL = 0.2

Threaded

: KL = 0.9

Tee with side flow Flanged

: KL = 1.0

Threaded

: KL = 2.0

Threaded Coupling : KL = 0.08

90° Mitre Bend (Without Vanes) : KL = 1.1

90° Mitre Bend (With Vanes) : KL = 0.2

Figure A6.5 Loss coefficient at direction control fittings.

For a 90° short radius bend Flanged

: KL = 0.3

Threaded : KL = 0.9

For a 45° sharp bend : KL = 0.4

For a 180° short radius bend Flanged

: KL = 0.2

Threaded : KL = 1.5

Figure A6.6 Resistance coefficient at direction control fittings.

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666

Appendix 6

Globe Valve, Fully Open : KL = 10

Angle Valve, Fully Open : KL = 2

Swing Check Valve : KL = 2

Gate Valve Fully Open : KL = 0.15 1/4 Closed : KL = 0.26 1/2 Closed : KL = 2.1 3/4 Closed : KL = 17

Figure A6.7 Resistance coefficient at flow control valves.

Table A6.1 Resistance coefficients for some pipe fittings.

Book.indb 666

Sr. no.

Fitting type and status (if applicable)

Leq or KL

1 2 3 4 5 6 7 8

Ball Check Valve Butterfly Valve Foot Valve, Poppet Disc Type Foot Valve, Swing Disc Type Butterfly Valve Ball Valve, Fully Open Ball Valve, 1/3 Closed Ball Valve, 2/3 Closed

Leq = 150 Leq = 45 Leq = 420 Leq = 75 Leq = 45 KL = 0.05 KL = 5.5 KL = 200

11/22/2010 2:43:46 PM

Appendix 7

A note on tungsten carbide and other hard metals

INTRODUCTION When it comes to cutting tools, tungsten carbide is the most widely used material. Whether it is metal cutting in manufacturing plants or rock cutting in the mining industry, the tools almost invariably make use of tungsten carbide inserts at the cutting points. Use of tungsten carbide inserted bits has become so widespread that almost all the manufacturers who earlier used to make steel tooth bits have turned towards manufacturers of tungsten carbide inserted bits. This note is intended towards giving some useful details of tungsten carbide inserts that could not find an appropriate place in the main book.

PLACE OF TUNGSTEN CARBIDE IN SUPERHARD MATERIALS There are a few materials that find a place above tungsten carbide in hardness ranking. Some such materials are mentioned in Table A.7.1. However, their use in cutting tools has remained at a very low level for one of many reasons such as: 1 2 3 4

Their hardness is lost as their temperature increases by the heat generated in cutting. Their density is comparable to the materials on which they are used as cutting tools so their separation for reuse is difficult. They are relatively rare to find or difficult to make and hence their use is not viable economically. Tools made by using them cannot be satisfactorily shaped in the powder metallurgical processes.

Some of the ‘harder than tungsten carbide’ materials are used in making grinding tools or some are used as an additive in the tungsten carbide while forming a tungsten carbide tool itself. Materials having hardness comparable to tungsten carbide and used in cutting tools, as stated in previous paragraph, are listed Table A.7.2.

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668

Appendix 7

GENERAL INFORMATION ON TUNGSTEN CARBIDE Tungsten is found in earth’s crust in the form of Wolframite (Fe,Mn)WO4 and Scheelite CaWO4. With almost 80% of global production share, China is the largest producer of tungsten. Annual global production of tungsten is estimated at 28000 tons. Two grades of tungsten carbide viz. W2C and WC were discovered in the laboratory of the School of Pharmacy at the University of Paris, by H. Moissan and P. Williams in the years 1896 and 1898.

Table A7.1 Materials ‘harder than tungsten carbide’. Density g/cm3

Hardness GPa

Bulk modulus GPa

Shear modulus GPa

3.52 3.48

70 45–50 40–45

443 400 496

535 409 332

2.51 4.5 2.66

30–33 32–35

244 Not Known

263 Not Known

AlMgB14 + Add1

2.67

35–40

Not Known

Not Known

AlMgB14 + Add2

2.7

40–46

Not Known

Not Known

Name

Formula

Diamond Boron Nitride Carbon Nitride Boron Carbide Titanium Boride Aluminum Magnesium Boride Aluminum Magnesium Boride Aluminum Magnesium Boride

C BN C3 N4 B4C TiB2 AlMgB14

Table A7.2 Materials of hardness comparable to tungsten carbide.

Book.indb 668

Hardness

Crystal structure

Melting point (°C)

Theoretical density, g/cm3

Modulus of elasticity, GPa

Thermal expansion, mm/m⋅K

Mono Tungsten Carbide

2,200

Hexagonal

2,800

15.63

696

5.2

W2C

Di Tungsten Carbide

3,000

Hexagonal

2,777

17.3

–

–

TiC

Titanium Carbide

3,000

Cubic

3,100

4.94

451

7.7

VC

Vanadium Carbide

2,900

Cubic

2,700

5.71

422

7.2

HfC

Hafnium Carbide

2,600

Cubic

3,900

12.76

352

6.6

ZrC

Zircon Carbide

2,700

Cubic

3,400

6.56

348

6.7

NbC

Niobium Carbide

2,000

Cubic

3,600

7.8

338

6.7

Carbide formula

Carbide name

WC

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Appendix 7

669

Tungsten carbide is shaped by the powder metallurgical process. Cobalt is the main additive. It acts as a binder between two crystals of tungsten carbide. Some other additives are also used in the process to give the desirable properties to the product.

TUNGSTEN CARBIDE AND ROCK DRILLING As far as rock drilling is concerned, almost 95% of the inserts used in bits and other components of the drill string are made from mono tungsten carbide, i.e. WC. The only other viable alternative to tungsten carbide is titanium carbide i.e. TiC. The first instance of use of sintered tungsten carbide inserts for use in rock drilling goes back to 1914 but these experiments were not made public. The experiments must have shown promise but due to the infancy of rock drilling techniques themselves and attention being given towards the first world war situation, it was probably not taken a note of. The invention of cemented carbide tool materials was first disclosed in 1923 in Karl Schroter/s patent application. The two inventions claimed were: 1

2

A unique hard metal alloy composition, namely the combination of the very hard tungsten carbide, WC, with small amounts of iron group metals viz. Fe, Ni and Co. The manufacture of the hard metal alloys by the application of the process of powder metallurgy, namely pressing and sintering of mixed powders of tungsten carbide and binder metal.

Again the patent holder and drilling industry were probably unaware of the promises of the use of tungsten carbide inserts. The first use of carbide inserted bits for rock drilling was made by Demag in Germany in the year 1936. After the end of Second World War these wonderful tools were noticed by other countries in the world. With a gestation period of about a decade, carbide inserted bits became a common tool in rotary percussion drilling in the late 1950’s. Tricone bits with tungsten carbide inserts appeared in the 1960’s.

MANUFACTURE OF TUNGSTEN CARBIDE INSERTS A tungsten carbide insert is formed by a powder metallurgical process wherein sintering is one of the major operation. The complete process is in steps as under, 1 2 3 4 5 6

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Making Fine Metal Powder Mixing Different Powders Pressing the Powder in the Die Presintering Sintering Grinding

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670

Appendix 7

Making fine metal powder The objective of this activity is to make a powder of tungsten carbide and other additives so that they have appropriate particle size, shape, particle size distribution and purity. Methods used for making powders are: reduction of oxides or ores, electrolytic deposition, atomizing, mill grinding or thermal deposition. As a general rule, finer size particles give better strength of the inserts. Special techniques have to be used for obtaining ultra-fine powders of sub micron size.

Blending and mixing different powders For tungsten carbide inserts the main component is tungsten carbide. Cobalt is the most widely used binder. Occasionally other carbides such as TiC, TaC and NbC are also mixed with the tungsten carbide and cobalt to give the desired properties. When tungsten carbide inserts to be formed are larger size special type of wax is also added. Wax is necessary as a lubricant for achieving uniform material density within the insert. This operation is called blending. After blending the powders are required to be mixed so as to get uniform distribution of all the ingredients. Mixing is carried out in special mills having specific properties of the mixing vessel.

Pressing the powder mixture in a die The mixture of the powders is pressed into a die to get the desired shape of the final product with adequate and uniform density. In this operation blended and mixed powder is put into a die either under gravity or by arranging pre-formed tablets of the mixture. The dies are pressed by hydraulic presses. Pressure varies between 150 to 500 MPa depending upon the size and shape of the insert. The final product of the pressing operation is the raw carbide insert of size much larger than the final sintered product.

Pre-sintering Pre sintering is an operation where the insert is heated to low temperatures. This operation is particularly essential in case of tungsten carbide inserts because they contain wax. Pre-sintering removes the wax from the raw insert in the form of gases that are pumped out of the furnace. Pre-sintering does not impart any hardness to raw inserts. Sometimes machining operations are carried out on pre-sintered products as the final sintered product is too hard to be machined.

Sintering The sintering process involves heating the pre-sintered insert to high temperatures so that the binder melts and flows throughout the tungsten carbide matrix. In this operation porosity is removed and a uniform binding is obtained around the tungsten carbide. The volume of the insert considerably reduces in this operation – sometimes by

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Appendix 7

671

as much as 35%. When the sintered insert is subjected to controlled cooling it attains a hardness to the desired level. The micro-structure of a coarse-grained tungsten carbide insert is shown in Figure A7.1. The dark gray particles in the figure are of tungsten carbide and the light gray represents the cobalt binder which firmly holds the tungsten carbide particles together.

Coating Many different types of coatings are applied to the tungsten carbide insert for attaining the desired properties. For example, a polycrystalline diamond coating is given to inserts for extreme wear resistance, or a titanium carbide coating is given to the insert for reducing friction between the insert cutting tool and the metal to be cut. Some coatings are spray-applied on the raw insert after the pre-sintering stage but before final sintering. Certain coatings, particularly polycrystalline diamond coatings (in the raw graphite form), are sprayed after sintering and the insert is again sintered so that the graphite turns into diamond.

Grinding The sintered insert shape is required to be within very precise prescribed limits, particularly in the case of tungsten carbide inserts in button form, because they are shrink-filled into the bit. Precision shaping is carried out by grinding. Grinding wheels of boron nitride or diamond are used in such operations, And use of a coolant is essential.

Tungsten Carbide Grains

Cobalt Binder

Figure A7.1 Micro structure of tungsten carbide.

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672

Appendix 7

PROPERTIES OF TUNGSTEN CARBIDE INSERTS From the viewpoint of rock drilling the important properties of a tungsten carbide insert are hardness, toughness and wear resistance. The tungsten carbide phase in the insert is responsible for hardness and wear resistance. The cobalt phase imparts toughness to the insert. As can be seen from Table A7.3 the hardness of the insert and wear resistance, which is closely related to the hardness, decrease very rapidly with increasing cobalt contents but toughness – represented by the transverse rupture strength – increases. Particle size also plays a very important role in increasing the hardness. Table A7.4 sufficiently indicates the effect of using very fine particles of tungsten carbide and cobalt powder. In each of the two compositions of the mixture used for inserts, the hardness as well as the compressive strength of the insert have increased by using fine particles. Wear resistance has also increased very significantly by using finer particles.

TUNGSTEN CARBIDE INSERTS FOR DRILLING BITS Tungsten carbide inserted bits are used for both rotary and rotary percussion drilling. In rotary drilling the drilling bit and consequently the tungsten carbide inserts are subjected to very heavy but relatively steady feed force. This causes the formation in contact with the insert to fragment. On the other hand, in rotary percussive drilling the insert is subjected to impact loads that lead to formation fracturing. By combining proportions in the mixture of tungsten carbide and cobalt, particle size, additives etc. several grades of tungsten carbide inserts have been made by manufacturers all over the world. To avoid complications, the industry has developed a code for indicating the suitability of inserts for applications involving steady force or impact. From this code the code designations meant for drilling applications are given in Table A7.5. Many reputed manufacturers give an industry code for each type of insert manufactured by them. Table A7.6 gives such information for tungsten carbide inserts manufactured by Kennametal. Generally, tungsten carbide inserts used for drilling have cobalt contents between 6 to 18%. Some manufacturers make inserts by using techniques which enable them to vary the proportion of cobalt from the central portion of the insert to the outer side

Table A7.3 Effect of proportion of tungsten carbide and cobalt on the insert properties.

W

Co

Other

Hardness at room temp. HV

80 90

20 10

0 0

1,050 1,625

Cemented carbide

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Modulus of elasticity GPa

Transverse rupture strength MPa

Coefficient of thermal expansion 10−6/K

Thermal conductivity W/m⋅K

Density g/cm3

490 580

2,850 2,280

6.4 5.5

100 110

13.55 14.5

11/22/2010 2:43:53 PM

Appendix 7

673

of the insert. This gives high hardness on the surface portion and high toughness at the center so that the insert becomes a dual property insert. Coating of Polycrystalline Diamonds and Tantalum Carbide on the tungsten carbide inserts is also used by some manufacturers.

Table A7.4 Effect of grain size of ingredients on the insert properties. Nominal composition

Hardness

Density

Transverse strength,

Compressive Modulus strength, of elasticity

WC

Cobalt Grain size

HRA

(g/cm3)

MPa

MPa

GPa

Relative wear resistance

97 94 94 94 90 90

3 6 6 6 10 10

92.5–93.2 92.5–93.1 91.7–92.2 90.5–91.5 90.7–91.3 87.4–88.2

15.3 15 15 15 14.6 14.5

1,590 1,790 2,000 2,210 3,100 2,760

5,860 5,930 5,450 5,170 5,170 4,000

641 614 648 641 620 552

100 100 58 25 22 7

Medium Fine Medium Coarse Fine Coarse

Table A7.5 Industry classification of sliding wear shocks and perpendicular impact shocks. Code designation

Suitability for wear application

Code designation

Suitability for impact application

C-9 C-10 C-11

No Sliding Wear shock Light Sliding Wear Shock Heavy Sliding Wear Shock

C-12 C-13 C-14

Light Perpendicular Impacts Medium Perpendicular Impacts Heavy Perpendicular Impacts

Table A7.6 Properties and industry codes for Kennametal tungsten carbide inserts. Trans. rupture strength

Density

Comp. Hardness strength

Designation WC

Grain Cobalt size

103xpsi

g/cm3

HRA

K96 K3833 K94 K3109 K92 K91 K3520 K90 KF306

94.5 89 88.5 87.9 84.3 80.5 80 75.2 94

5.5 11 11.5 12.1 15.7 19.5 20 24.8 6

300 430 410 430 475 440 430 430 390

14.85 14.4 14.2 14.2 13.8 13.4 13.45 13.8 14.95

KF310

90

10

500

KF315

85

15

550

Composition by% weight

Book.indb 673

Fine Fine Fine Fine Fine Fine Fine Fine Extra Fine Extra Fine Extra Fine

103xpsi

Relative impact resistance

Industry code

92.1 89.4 89.8 88 88.4 86.8 84.2 84.8 93.3

765 705 720 635 670 605 530 505 850

29 100 57 100 82 89 100 71 –

C-9©-10 C-11©-13 C-10©-12 C-11©-14 C-11©-13 C-11©-14 C-14 C-14 C-9

14.5

92.2

790

–

C-10©-12

13.9

90.4

660

–

C-11©-13

11/22/2010 2:43:54 PM

Appendix 8

Scales of hardness – Measurements and their conversion

Hardness of substances is determined by different types of tests. Minerals and metals are the two main types of substances on which such tests are carried out. Scales used for attributing hardness value to a mineral are: Moh’s Scale, New Moh’s Scale and Knopp’s Scale of Hardness. Of these three scales the Moh’s Scale is the oldest and has been used most widely in the scientific literature in the past. New Moh’s Scale has a few more steps of hardness. Knopp’s Scale is more rational in the relative hardness and therefore its use is ever increasing. Hardness levels of some of the common minerals are as given in the following table. The figures in the parenthesis after the name of some of the minerals indicate the crystallographic character of the crystals of minerals. It may be noted that some reference books give different values of hardness and corresponding minerals particularly in case of New Moh’s Scale of Hardness. Hardness of metals is determined by several tests. The hardness values attributed by these tests are different. Fortunately, due to systematic testing done in the past, the values of metal hardness can be correlated with each other. The table on the next page gives equivalent hardness values for most of the hardness scales. Tables on the next page give equivalent values of hardness by different hardness scales. The meanings of symbols used for explanation are as under. C A D B A1 A2 A3 K1 K2 S1 S2 S3

Book.indb 675

-

Rockwell C Scale with 150 kg load on the diamond brale, Rockwell A Scale with 60 kg load on the diamond brale, Rockwell D Scale with 100 kg load on the diamond brale, Rockwell B Scale with 100 kg load on the 1/16” ball, 15 kg load on 15 N penetrator, 30 kg load on 30 N penetrator, 45 kg load on 45 N penetrator, Diameter of the ball impression in mm, Hardness number, Vicker’s or Firth hardness number, Scleroscope hardness, Tensile Strength in kPa.

11/22/2010 2:43:54 PM

Book.indb 676

Table A8.1 Moh’s and New Moh’s scales of hardness. Moh’s scale of hardness

New Moh’s scale of hardness

Knopp’s scale of hardness

Hardness level

Corresponding mineral

Chemical formula

Hardness Corresponding level mineral

Chemical formula

1

Talc

Mg3Si4O10(OH)2

1

Talc (001)

Mg3Si4O10(OH)2 2.5

Talc (001)

Mg3Si4O10(OH)2

2

Halite (100)

NaCl

25

Halite (100)

NaCl

60

Gypsum

CaSO4⋅2H2O

80

Galena (100)

PbS

125

Calcite

CaCO3

2

Gypsum

CaSO4⋅2H2O

3

Calcite

CaCO3

4

Fluorite

CaF2

3

Galena (100)

PbS

Hardness level

Corresponding mineral

Chemical formula

4

Fluorite (100)

CaF2

175

Fluorite (100)

CaF2

5

Scheelite (111)

CaWO4

370

Scheelite (111)

CaWO4

525

Apatite

Ca5(PO4)3(OH,Cl,F)

600

Magnetite (111)

Fe3O4

730

Orthoclase

KAlSi3O8

5

Apatite

Ca5(PO4)3(OH,Cl,F)

6

Orthoclase

KAlSi3O8

7

Quartz

SiO2

7

Quartz(1011)

SiO2

1060

Quartz (1011)

SiO2

8

Topaz

Al2SiO4(OH,F)2

8

Topaz (001)

Al2SiO4(OH,F)2

1520

Topaz (001)

Al2SiO4(OH,F)2

9

Corundum

Al2O3

9

Corundum (1120)

Al2O3

2120

Corundum (1120)

Al2O3

10

Titanium Carbide

TiC

3000

Titanium Carbide

TiC

11

Boron

Pure Boron i.e. B

4000

Boron

Pure Boron i.e. B

6

10

Diamond

Pure Carbon i.e. C

Magnetite (111)

Fe3O4

11/22/2010 2:43:54 PM

12

Boron Carbide

B4C

4910

Boron Carbide

B4C

13

Boron Carbide

B6.5C

5750

Boron Carbide

B6.5C

14

Diamond Carbando

Pure Carbon i.e. C

8000

Diamond Carbando

Pure Carbon i.e. C

15

Industrial Diamond

Pure Carbon i.e. C

10000

Industrial Diamond

Pure Carbon i.e. C

Table A8.2 Equivalent values of hardness on different scales. Rockwell

Superficial Rockwell

Diamond brale C

A

D

80 79 78 77 76 75 74 73 72 71 70 69 68 67 66 65 64 63 62 61 60 59 58 57 56 55 54 53 52 51 50 49 48 47 46 45 44 43 42 41

92 92 91 91 90 90 89 89 88 87 87 86 86 85 85 84 84 83 83 82 81 81 80 80 79 79 78 77 77 76 76 75 75 74 73 73 73 72 72 71

87 86 85 84 83 83 82 81 80 80 79 78 77 76 76 75 74 73 73 72 71 70 69 69 68 67 66 65 65 64 63 62 61 61 60 59 59 58 57 56

1/16" “N” brale penetrator ball B A1 A2 A3 97 96 96 95 95 94 94 93 93 92 92 91 91 90 90 89 89 88 88 87 87 86 86 86 85 85 84 84 83 83 82 82 81

92 92 91 91 90 89 89 88 87 87 86 85 85 84 83 82 81 80 79 79 78 77 76 75 74 73 72 71 70 69 69 68 67 66 65 64 63 62 61 60

87 87 86 85 84 83 82 81 80 79 78 77 79 75 73 72 74 70 69 68 67 66 65 63 62 61 60 59 57 56 55 54 53 51 50 49 48 47 46 44

Brinell 10 mm ball 3000 kg load K1

2.25 2.30 2.30 2.35 2.35 2.40 2.45 2.55 2.55 2.60 2.60 2.65 2.70 2.75 2.75 2.80 2.85 2.90 2.90 2.95 3.00 3.00 3.05 3.10 3.10

K2

S1

S2

S3

745 710 710 682 682 653 627 578 578 555 555 534 514 495 495 477 461 444 444 432 415 415 401 388 388

1,865 1,787 1,710 1,633 1,556 1,478 1,400 1,323 1,245 1,160 1,076 1,004 942 894 854 820 789 763 746 720 697 674 653 633 613 595 577 560 544 528 513 498 484 471 458 446 434 423 412 402

99 98 97 96 95 93 91 88 87 85 83 82 80 78 77 75 74 72 71 69 68 67 65 64 63 62 61 59 58 56 55

2248 2172 2096 2027 1979 1924 1855 1800 1751 1689 1641 1600 1551 1510 1455 1420 1393 1365 1317 (Continued)

Book.indb 677

11/22/2010 2:43:54 PM

Table A8.2 (Continued). Rockwell Diamond brale

Book.indb 678

C

A

D

40 39 38 37 36 35 34 33 32 31 30 29 28 27 26 25 24 23 22 21 20 18 16* 14* 12* 10* 8* 6* 4* 2* 0*

70 70 69 69 68 68 67 67 66 66 65 65 64 64 63 63 62 62 62 61 61

55 55 54 53 52 52 51 50 49 48 48 47 46 45 45 44 43 42 42 41 40

1/16" ball B

109 109 108 108 107 106 106 105 104 103 103 102 101 100 99 99 98 97 95 94 92 90 89 87 85 84 82 81 79 77 74 72 70 68 65

Superficial Rockwell

Brinell

“N” brale penetrator

10 mm ball 3000 kg load

A1

A2

A3

K1

K2

S1

S2

S3

80 80 79 79 78 78 77 77 76 76 75 75 74 73 73 72 72 71 71 70 69

60 59 58 57 56 55 54 53 52 51 50 50 49 48 47 46 45 44 43 42 42

43 42 41 40 39 37 36 38 34 33 32 30 29 28 27 26 24 23 22 21 20

3.15 3.2 3.25 3.3 3.35 3.35 3.4 3.45 3.5 3.55 3.6 3.65 3.7 3.75 3.8 3.80 3.85 3.90 3.95 4.00 4.05 4.10 4.15 4.25 4.35 4.40 4.50 4.60 4.65 4.80 4.80 4.90 5.00 5.10 5.20 5.30 5.40 5.50 5.60

375 363 352 341 331 331 321 311 302 293 285 277 269 262 255 255 248 241 235 229 223 217 212 203 192 187 179 170 166 156 156 149 143 137 131 126 121 116 112

392 382 372 363 354 345 336 327 318 310 302 294 286 279 272 266 260 254 248 243 238 230 222 213 204 195 187 180 173 166 160 156 150 143 137 132 127 122 117

54 53 51 50 49 48 46 45 44 43 42 41 40 39 38 37 37 36 35 35 34 33 32 31 29 28 27 26 25 25 25 23 22 21

1276 1248 1213 1179 1158 1124 1096 1062 1034 1007 979 951 924 903 867 855 841 814 800 779 765 738 703 676 634 621 600 572 545 531 510 503 483 462 448 427 414 400 386

20 19 18 15

11/22/2010 2:43:55 PM

Appendix 9

Manufacturers of rotary blasthole drilling equipment

Table A9.1 gives a list of items frequently needed in rotary blasthole drilling. Names of manufacturers or suppliers are listed in Table 9.2 where in column 5, the products in their range are listed. Table A9.1 Symbols used for showing the items manufactured by the manufacturers listed in Table A9.2. Item

Book.indb 679

Symbol

Item

Symbol

Item

Symbol

Air Measuring Kit

AMK

Drill Pipe Integrated

DPI

Recovery Tools

RT

Bit Breaker

BB

Drill Stem Wrench

DSR

Shock Absorber External

SAE

Blasthole Camera

BC

Feed Force Measuring Kit

FFK

Shock Absorber Internal

SAI

Blasthole Dewatering Pump

BDP

Drill GPS

GPS

Stabilizers Integral Blade

SIB

Blasthole Inclinometer Drill Mounted

BID

Hoisting Plugs

HP

Stabilizers Roller

SR

Blasthole Inclinometer Handheld

BIH

Lifting Bails

LB

Stabilizers Replaceable Sleeve

SRS

Blasthole Plugs

BP

Laser Measuring Instruments

LMI

Stabilizer Welded Blade

SWB

Blasthole Sampler

BS

Rotary Deck Bushing

RDB

Tricone Bit Carbide Insert

TBC

Crossover Sub

CS

Rotary Blasthole Drill Large (Bit Load 351–551 kN)

RDL

Tricone Bit Steel Tooth

TBS

Drag Bit

DB

Rotary Blasthole Drill Medium (Bit Load 225–351 kN)

RDM

Miscellaneous Items

MI

Drill Computer

DCO

Rotary Blasthole Drill Small (Bit Load 100–225 kN)

RDS

Drill Loggers

DL

Rotary Blasthole Drill Truck/ Rubber Tired

RDT

Drill Pipe Fabricated

DPF

Rotary Blasthole Drill Ultra Large (Bit Load 551kN or More)

RDU

11/22/2010 2:43:55 PM

Table A9.2 Manufacturers of rotary blasthole drilling equipment. Sr. No. Manufacturer name

Book.indb 680

Manufacturer’s address

1

Aquila Drilling System

2

Atlas Copco Drilling Solutions 2100, North First Street, Garland Texas 75040, USA Inc.

100 North East Adams Street, Peoria. Illinois, 61629 USA

3

B. F. Carr & Associates

P.O. Box 220250, St Louis, Missouri63122–250, USA

4

BDS Inc.

P. O. Box 1160, Malta, Montana, 59538 USA

5

Bit Brokers International

P. O. Box 100, 5568 Logan Road, Logan, Illinois, USA

6

Blue Demon

P. O. Box 724, Lebanon, Georgia, 30146 USA

7

Bucyrus International Inc.

1100 Milwaukee Avenue, South Milwaukee, Wisconsin, 53172–0500, USA

8

Colcrete Eurodrill Co. Ltd.

Tower Business Park, Derby Road, Clay Cross, Derbyshire, S45 9 AG United Kingdom.

9

D. K. Thomas Equipment Corporation

P. O. Box 200, 1453 Route 9, Spofford, New Hampshire, 03462 USA

10

DataMetrics Corporation International

1717 Diplomacy Row, Orlando, Florida, 32809 USA

11

Diversified Drilling & Industrial Equipment F.Z.C.

Al Mamoura Building, Third Floor, Block (c), Beside of Ras Al Khaimah Immigration, Ras Al Khaimah, United Arab Emirates

12

Downhole Stabilization Inc.

P. O. Box 2467, 3515 Thomas Way, Bakersfield, California, 93308 USA

13

Driconeq Drilling Construction Equipment AB

P. O. Box 325, Svetsarevagen 4, Sunne 686 26, Sweden

14

Drill Pro International Inc.

108 E, 39th Street, Boise, Idaho, 83714 USA

15

Drill Supplies Inc.

1350 E. Glendale Avenue, Sparks, Nevada, 89431 USA

16

Dryrite Pumps

1854 Stevens Road, Eagle Point, Oregon, 97524 USA

17

DZBO

49100, Dnepropetrovsk, Udarnikov str., ap. 27

18

Eimco Elecon (I) Ltd.

Anand-Sojitra Road,Vallabh Vidyanagar, Gujarat 388120 India

19

Foremost Industries LP

1225 – 64th Avenue NE, Calgary, Alberta, T2E 8P9 Canada

20

Galleon Turbeco Flotek Company

11317 W. County Road. 128 Midland, Texas 79711, USA

21

Geophysical Instrument & Supply Co.

6323 Cambridge Street, Minneapolis, Minnesota, 55416 USA

22

Gill Rock Drill Co. Inc

903–905 Cornwall Road, Lebanon Road, Pennysylvania, 17042 USA

23

Globe Drill

8 Nambi Way, Kalgoorlie, Western Australia, 6430

24

Grant Prideco Drilling Products and Services

1450 Lake Robbins Dr. Suite 600, The Woodlands, Texas, 77380 USA

25

HAUSHERR System-Bohrtechnik GmbH & CO. KG

Heisenbergstr. 11, D-59423 Unna, West Germany

11/22/2010 2:43:56 PM

Tel. / Fax No.

Website

Products manufactured/sold

001 309 675 1000

www.cat.com/minestar

DCO, GPS

001 972 496 7400/

www.cmt.atlascopco.com

RDS, RDM, RDL, RDU, RDT, SR, CS, DSR, TBS, TBC, DPF, RDB, SAE

001 314 497 6075 /001 314 966 6985

www.stemplug.com

BP

001 406 654 1727 /001 406 654 2778

www.bds-usa.com

BDP

001 618 435 5811 /001 618 435 2388

www.bitbrokers.com

DB

001 770 591 2021 /001 770 591 6356

www.bluedemon.com

DB, TBC

001 414 768 4000 /001 414 768 4474

www.bucyrus.com

RDM, RDL, RDU, DPF

044 1246 868700 /044 1246 868701

www.colcrete-eurodrill.co.uk DPF

001 603 363 4706 /001 603 363 4249

www.driller.com

BD, BDG, BDP

001 407 251 4577 /407 251 4588

www.datametrics.com

GPS

00971 7 228 9223 / 00971 7 228 9233

No Website

SAI, SIB, SR, SRS, SWB

001 661 631 1044 /001 881 631 1195

No Website

SIB, SWB

046 565 121 10 /046 565 146 20

www.driconeq.com

CS, HP, LB, DPF, RB, SWB

001 208 336 4970 /001 208 345 1193

www.drillpro.net

RDB TBI, TBS, SIB, DPF

001 775 355 1199 /001 775 355 8775

www. americawestdrillingsupply. com

TBS, TBC, DB, CS, RT

001 541 830 1403 /001 541 830 0208

www.dryritepumps.com

BDP

038 056 7901157, 038 056 7901159

DPF, DPI, RT, CS

091 2692 227812 /091 2692 236506

www.emico.elecon.com

RDM

001 800 661 9190

www.foremost.ca

RDB, SAE

001 432 563 4172 / 001 432 563 1873

SR, SIB, SWB

001 952 929 8000 /001 952 926 5498

www.giscogeo.com

BC

001 717 272 3861 /001 717 272 4140

www.gillrockdrill.com

DPF, RT, CS,BP

061 8 9091 3008 / 061 8 9091 3009

www.globedrill.com.au

RDS

001 281 297 8500 /001 281 297 8525

www.grantprideco.com

DPF, DPI,

049 23 03–98624 /049 23 03–986241

www.hausherr.com

RDS, RDM, RDT, DPF

(Continued)

Book.indb 681

11/22/2010 2:43:56 PM

Table A9.2 Continued.

Book.indb 682

Sr. No. Manufacturer name


26

Heavy Engineering Corporation Ltd.

Dhurwa, Ranchi, Jharkhand, 834004 India

27

Henan Shenlong Petroleum Drilling Tools Co. Ltd

Technical Industrial Park, JiYuan, Henan, China

28

Ideal Supply

P. O. Box 15397, Asheville, North Carolina, 28813 USA

29

JSI Rocktools

11F, Jiulian Building, No. 213, North Baoshan Road, Guiyang 550004, Guizhou, China

30

Kay Rock Bit Company

P. O. Box 684865, Austin, Texas, 78768, USA

31

King Pin International

Hoffsj. Lovensky, 76B NO-0382, Oslo, Norway

32

Kingsland Drill International Inc.

7920 Ward Drive, Mercersburg, Pennysylvania, 17236 USA

33

Legra Pumps

9 Fursden Street, Glenella, Mackay, 4740 Queensland, Australia

34

Leica Geosystems

2700 E Executive Drive, Suite 100, Tucson, Arizona 85706, USA

35

LIM SA

Rue Jean Bartin,Villeurbanne 69100, Rhone, France

36

LMP Precision Engineering Co. Pvt. Ltd.

Plot No. 2P&3, GIDC, Antalia, Bilimora, Gujarat, 396325, India

37

Lone Star Drill Bits

P. O. Box 1219, Stafford, Texas,77497 USA

Marks Products Inc.

1243 Burnsville Road, Williamsville,VA 24487–2147, USA

M&P Flange and Pipe Protection

9426 Katy Fwy Bldg No. 11, Houston, Texas 77055, USA

38

Measurement Devices Ltd.

Acer House, Hackness Road, Northminster Business Park, Upper Poppelton,York,Yo26 6QR, United Kingdom

39

Mills Machine Company Inc.

P. O. Box 1514, Shawnee, Oklahoma, 74802 USA

40

Mining Technology International Inc

4400 North Big Spring, Bldg E, Suite 120, Midland, Texas 79705, USA

41

Modular Mining Systems Inc.

8289 E. Hemisphere Loop, Tucson, Arizona, 85706–5028, USA

42

MTI Group

3 Yeeda Way, Malaga, Western Australia 6090

43

NKMZ

NKMZ, Kramatorsk, Donetsk Region, 84305 Ukraine

44

OMZ

Ermolaevskiy per., 25, str. 1, Moscow, 123379

45

P&H Mining

4400 West National Avenue, Milwaukee, Wisconsin, 53214, U. S. A.

46

Pulsar Measuring Systems Ltd. 3, Forge Close, The Street, Chipperfield, Herts, WD4 9DL United Kingdom

47

R. Smith & Sons Inc.

2605 Chapman Street, Houston, Texas 77009,USA

48

Reamco Inc.

1149 Smede Highway, Broussard, Houston, Texas. 70518 USA

11/22/2010 2:43:56 PM

Tel. / Fax No.

Website


091 651 240 1249 /091 651 2401571

www.hecltd.com

RDM, DPF, CS

086 391 663 5039 /086 391 663 5036

www.petrodrillingtoos.com

DPF, DPI, SIB, CS

001 828 274 0144 /001 828 274 0140

www.idealsupply.com

MI

086 851 6780905 /086 851 6780960

www.jsintl.com.cn

DB, TBS, TBC

001 899 556 8665 /-

www.kayrockbitcompany. com

DB, DPF

047 2252 3638 /047 2252 3640

www.kingpin-international. com

RDB

- /001 717 328 4998

www.kingslanddrill.com

DP

061 7 4942 3922 /061 7 4942 3868

www.legra.com.au

BDP

001 520 529 8729 /-

www.leica-geosystems.com

GPS

033 472 146530

www.limgeo.com

DL, BID

091 2634 284714 /091 2634 281079

www.lmpprecision.com

RDS, RDM, DPF, CS,

- /001 877 817 0978

www.lonestarbits.com

DB, TBS, TBC

001 540–396–4740 /001540–396–4741 www.marksproducts.com

BC

001 800 8882605 /001 713 4636343

www.m-p.com

LB, MI

044 1904 791 139 /044 1904 791 532

www.mdl.co.uk

Ideal Supply

001 405 2734900 /001 405 273 4956

www.millsmachine.com

DB, SWB, CS, RT, LB, HP

001 432 684 4010 /001 432 684 4306

www.mti.com

DPF, DPI, SWB, SR, SAE, CS, LB, RDB

001 520 746 9127 /001 520 889 5790

www.modularmining.com

GPS

061 8 9209 2111 /061 8 209 2200

www.mtigroup.com.au

BP, MI

038 6264 37080 /038 6264 72249

www.nkmz.com

RDM, RDL, DPF

007 095 974 6021/007 095 796 9003

www.omz.ru

RDL,RDM, DPF

001 414 671 4400 /001 414 671 7236

www.phmining.com

RDL, RDU, DPF

044 1923 261412 /044 1923 261299

www.pulsarmeasuring.com

BIH

001713 2238495 /001 713 225–5015

www.r-smith-and-son.com

LB, MI

001 337 364 9244 /001 337 3649603

www.reamcoinc.com

SWB, SIB, SRS, DPF, DPI

(Continued)

Book.indb 683

11/22/2010 2:43:56 PM

Table A9.2 Continued. Sr. No. Manufacturer name

Book.indb 684


49

Reed Hycalog Canada

2700–144–4 th Avenue SW, Calgary, Alberta, T2P 3 N4 Canada

50

Revathi Equipment Limited

Pollachi Road, Coimbtore, Tamil Nadu 641021 India

51

Rotacan

400-B Kirkpatrick Street, North Bay, Ontario, P1B8G5 Canada

52

Rudgormash

53

Sandvik Mining and Construction

SE-811 81 Sandviken, Sweden

54

Schlumberger Ltd

300 Schlumberger Drive, Sugar Land, TX 77478, USA

55

SCHRAMM Inc.

800 E.Virginia Avenue, West Chester, Pennysylvania 19380, USA

56

Shanxi Oil Drilling Industry Central Manufacturing Co.

Jinzhong City, Shanxi Province, Chaoyang Shouyang County, China

57

Smith Drilling

11, Junction Road, Industries North, Germiston, Johannesburg, Gauteng, 2008 South Africa

58

Smith International Inc.

16740 Hardy Street, Houston, Texas, 77032 USA

59

Sphere Drilling Supplies

3112 -80th Avenue SE, Calgary, Albarta, T2C1 J3, Canada

60

Star Iron Works

P. O. Box 155, R. D. 3, Punxsutawney, Pennysylvania, 15767 USA

61

TeeKee

P. O. Box 313, 295 Prince Street, Despatch, Silverton 0127, South Africa.

62

Thread Protectors LLC

806 East Main St. Unit F, Broussard, La. 70518 USA

63

Thunderbird Mining Systems

2609 151st Place N. E., Redmond, WA, 98052–5522, USA

64

Torquato Drilling Accessories

215 Susquehanna Street, Old Forge, Pennsylvania 18519 USA

65

Total Engineering Pvt Ltd.

2 Stockdale Road, O’conner, 6163 Western Australia

66

Trimble Navigation Ltd

935 Stawart Drive, Sunnyvale, California, 94085 USA

67

TSK Bits

TB Tamachi Building, 5&6 Floor, No 11–1, Shiba 4-Chome, Minato-ku, Tokyo 108–0014 Japan

68

UDR Sampling System

332 Treasure Road, Welshpool, WA 6106, Australia

69

UralBMT

Verhnie Sergi, Near Ekaterinburg, Ural Mountain Region Russia

70

Varel International Inc.

P. O. Box 5401571343 Patton, Suite 106, Carrollton, Texas 75007 USA

71

Vulcan Industrial Engineering Co. Ltd.

A-2/440, G.I.D.C.,Vithal Udyognagar, Anand, Gujarat, 388121 India

72

Weatherford International Ltd.

515 Post Oak Blvd.,Suite 600, Houston, Texas 77027, USA

11/22/2010 2:43:56 PM

Book.indb 685

Tel. / Fax No.

Website


001 403 234 9999 /001 403 264 3300

www.reedhycalog.com

TBS, TBC, AMK

091 422 2610851 /91 422 2610427

www.revathi.co.in

RDS, RDM, RDL, DPF, CS

001 705 474 5858 /001 705 474 9595

www.rotacon.com

TBC

046 26 262000 /046 26 260222

www.sandvik.com

RDS, RDM, RDL, RDU, RDT, TBS, TBC, DPF, RDB

www.schlumberger.com

BB

001 610 696 2500 /001 610 696 6950

www.schramminc.com

RDS, DPF

086 354 4602448 /086 354 4602431

www.huanjie.com.cn

DPF, DPI, CS, SIB, SWB,

027 11 873 9830 /027 11 825 6830

www.smithdrilling.co.za

RDS, DPF, TBS, TBC, CS, SWB

001 281 443 3370 /001 281 233 5190

www.smithbits.com

TBS, TBC, AMK

001403 720 9333 /001 403 236 2010

www.sphere-drilling.com

RDB, DPF

001 814 427 2555 /001 800 927 0560

www.starironworks.com

DPF, SWB, CS,

027 12 803 7305 /027 12 803 8970

www.tee-kee.co.za

DPF, RDB, CS, SWB, RT, SAE

001 337 837–8881 /001 337 837–8885 www.threadproducts.com

HP, LB, MI

001 425 869 2727 /001 425 869 2735

www.thirdpac.com

GPS

001570 457 8565 /001570 457 8585

www.torquato.com

TBC

061 8 9314 5333 /061 8 9314 5336

www.totaldrillsupport.com

RDB, DPF

001 408 481 8000 /-

www.trimble.com

GPS

081 3 5730 1913 /081 3 5730 1915

www.tskbits.com

TBS, TBC

061 8 9351 9666 /061 8 9351 9555

www.udr.com.au

BS

001 630 690 5625 /001 630 560 5689

www.uralbmt.com

TBS, TBC

001 972 242 1160 /001 972 242 8770

www.varelintl.com

TBS, TBC, AMK, FFK

091 2692 236032 /091 2692 238723

www.viecl.com

DPF, DPI, CS, SR, SIB, SWB, RDB

001 713 693 4000

www.weatherford.com

SIB, SWB

11/22/2010 2:43:56 PM

Appendix 10

Details of rotary blasthole drills manufactured all over the world

This appendix gives important details of crawler-mounted rotary blasthole drills in Table A10.1 and carrier-mounted rotary blasthole drills in Table 10.2. The following points should be noted in respect of these details. 1

2

3 4

5 6

7

Book.indb 687

Most of the details presented have been checked. However, every manufacturer is in continuous quest for improvement in the product and therefore, details do change with every improvement in the blasthole drill. Details presented here may have changed even before you read this compilation. For most authentic and detailed specifications, the manufacturer of the drill should be contacted. Addresses of all the manufacturers have been given in a separate appendix. Most other specifications of rotary blasthole drills are fairly easy to understand and apprehend. However, since mast height (i.e. length), drill head, number of pipes in the pipe rack, wall thickness of the drill pipes etc. greatly affect the stability of the blasthole drill, one must be very careful in choosing the appropriate combinations of these items. Rather than merely relying on the specifications presented here or in the product brochure, the manufacturer must be consulted to ensure that what is desired from the drill is actually being fulfilled by the drill under consideration. In the following tables wherever some detail of the drill could not be obtained the specific place of the detail is marked with XXX. In the case of type of drill, the first letter indicates source of power and the second letter indicates the mode of power distribution. In this context E means electric, D means diesel and H means hydraulic Abbreviations used in the case of main power source are DE – Diesel Engine, SEL – single electric motor, MEL – many electric motors. In the tables the technical details have been given in International Units but exception has been made in case of weight of the blasthole drill, where more familiar unit viz. kg has been used instead of kN. Rotary drills having bit load capability less than 10000 kg are not included in this appendix because such drills are unsuitable for rotary drilling and are usually used for DTH drilling.

11/22/2010 2:43:57 PM

Table A10.1 Major details of crawler mounted rotary blasthole drills.

Max.

Min.

Max. Machine Weight

Rotary Speed Range

Max. Torque

Max. Bit Max. Feed Load Speed Nos.

Max. Discharge

Pressure

Main power source D = Diesel Engine, SE = Single Electric Motor, ME = Many Electric Motors

mm

mm

kg

RPM

Nm

kN

m/min

Nos.

m3/min

kPa

Type, kW

Blasthole Diameter

Largest compressors provided on the drill

Manufacturer

Model

Type E to E, E to H, D to E, D to H

Atlas Copco

PV 351

D/E to H

406

269

188240

0–87

25760

556

48.15

1

107.6

758

DE, 1230/SEL, 1043

Atlas Copco

DM -M3

D to H

311

251

104327

0–200

13806

400

43.9

1

73.6

758

DE, 709

Atlas Copco

PV275

D/E to H

269

171

83915

0–150

11795

334

38.71

1

73.6

758

DE, 708/SEL, 671

Atlas Copco

PV271

D/E to H

269

171

83915

0–150

11795

334

38.71

1

73.6

758

DE, 708/SEL, 671

Atlas Copco

PV235

D/E to H

251

152

58060

0–160

10575

289

58.8

1

53.8

758

DE, 597/SEL, 522

Atlas Copco

DML

D/E to H

270

149

49895

0–160

9762

267

44.5

1

53.8

758

DE, 596/SEL, 522

Atlas Copco

DML-SP

D to H

251

152

45360

Table 0–100

10169

245

60

1

53.8

758

DE, 596

Atlas Copco

DM50

D to H

254

152

41382

0–107

10575

222

69

1

34

758

DE, 418

Atlas Copco

DM45

D to H

254

149

38556

0–107

10575

200

44.5

1

29.7

758

DE, 380

Atlas Copco

DM30

D to H

171

127

28121

0–100

7321

133

30.48

1

25.5

758

DE, 317

Atlas Copco

DM25-SP

D to H

178

102

28121

Table 0–170

4746

111

21.9

1

25.5

758

DE, 317

Bucyrus International

59R

E to E

445

311

210920

0–120

21647

704

30.4

1

107.6

448

MEL, 896


49R

E to E

406

251

174663

0–125

16812

627

22.8

1

107.6

482

MEL, 896


39HR

E to H

349

229

149668

0–160

18845

543

47.8

1

85

310

SEL, 746


SKL – 20

D/E to H

381

229

131088

0–150

20880

489

37.5

1

102

689

DE, 899/SEL895


SKL – 16

D/E to H

381

229

131088

0–150

20880

489

37.5

1

102

689

DE, 899/SEL895


SKS – 16

D to H

311

229

84700

0–150

15185

411

33.5

1

68

689

DE, 784


SKS – 13

D to H

311

229

84700

0–150

15185

411

33.5

1

68

689

DE, 784


SKS – 10

D to H

311

229

84700

0–150

15185

411

33.5

1

68

689

DE, 784


SKF – 11

D to H

270

152

54583

0–220

12880

276

42.9

1

48.1

689

DE, 597


SKF – 10

D to H

270

152

54583

0–220

12880

276

42.9

1

48.1

689

DE, 597


SKF – 8

D to H

270

152

54583

0–220

12880

276

42.9

1

48.1

689

DE, 597


SKFX – 15

D to H

270

152

65026

0–220

12880

236

38.1

1

48.1

689

DE, 597


SKFX – 12

D to H

270

152

65026

0–220

12880

236

38.1

1

48.1

689

DE, 597

Changsha Keda CMCL

DR400

D to H

405

351

180000

0–150

25000

550

20

1

400

DE, 1500

Changsha Keda CMCL

DR370

D to H

351

269

136000

0–150

17500

490

18

1

84.9

700

DE, 1200

Changsha Keda CMCL

DR310

D to H

311

251

105000

0–150

14000

400

20

1

74.6

760

DE, 950

Changsha Keda CMCL

DR270

D to H

269

193

99000

0–150

6500

311

39

1

53.8

750

DE, 755

Emico Elecon

250 E

E to H

279

229

67000

0–97

14091

329

26.4

1

31.15

689

SEL, 373

Emico Elecon

160 D

D to H

203

127

34000

0–100

5000

142

20.8

1

17

689

DE,220

Globe Drill

KAL500

D to H

229

152

52000

0–125

5000

222

4.5

1

39

700

DE, 676

Hausherr

HBM160

D/E to H

190

130

40000

0–100

8000

160

23

1

30

700

DE, 470/SEL, XXX

Hausherr

HBM120

D/E to H

170

190

30000

0–134

6900

120

35

1

27

700

DE, 373/SEL, XXX

Heavy Eng. Corp.

HE5–250–4

E to E

270

250

90000

0–150

9316

294

1

1

28.31

689

MEL, 310

LMP Precision

SK100

E to H

279

200

65000

0–135

7912

311

3

1

34

689

DE, 373/SEL 373

LMP Precision

RBH-6–100 D/E

D/E to H

165

140

27000

0–135

4971

160

14

1

21.23

689

DE, 76 + DE, 224

NKMZ

SBSHS-250/270–32

E to E

270

243

90000

0–150

6000

400

13

1

40

689

MEL, 648

NKMZ

SBSHS-250H

E to E

250

229

78000

0–120

6500

350

12

1

32

689

MEL, 500

OMZ

SBSh-270-IZ

E to E

311

270

136000

0–120

13000

450

6

1

38

490

MEL, 648

OMZ

SBSh-270–34

E to E

270

229

141000

0–120

8400

350

6

1

38

490

MEL, 648

OMZ

SBSh-G-250

E to E

250

229

90000

0–150

5870

300

6

1

32

490

MEL, 500

P&H Mining

320XPC

E to E

445

311

165000

0–119

25760

667

24.4

1

109

448

MEL, 805

P&H Mining

250XP-ST

D to H

349

243

113400

0–200

16269

466

38.1

1

102

448

DE, 746

P&H Mining

250XP-DL

E to H

349

243

113500

0–200

16269

466

38.1

1

102

448

DE, 746

Revathi Equip’nt

C850E

E to H

349

311

130000

0–100

13050

445

15

1

45.3

689

SEM, 500

Revathi Equip’nt

C750

D/E to H

270

250

67000

0–95

11300

311

19

1

34

689

DE, 402/SEL, 373

Revathi Equip’nt

C650

D/E to H

200

160

40000

0–100

4520

133

12

1

32.5

2413

DE, 402/SEL, 373

Rudgormash

SBSH-250 MNA-32

E to E

270

160

85000

0–120

6500

350

1

1

32

689

Rudgormash

SBSH 160/200 40D

D to H

215

160

55000

0–120

XXX

235

XXX

1

25

780

DE, 485

Sandvik

1190 E

E to H

349

229

140614

0–97

16869

523

17

1

94.4

414

SEL, 746

Sandvik

D90 KS

D to H

349

229

104328

0–97

16869

523

21.6

1

85

552

DE, 839

Sandvik

DR 460

D to H

311

251

93440

0–175

10462

445

XXX

1

56.63

689

DE, 746

Sandvik

D75 KS

D to H

279

229

64864

0–94

14236

409

27

1

56.63

689

DE, 597

Sandvik

D50 KS

D to H

229

152

47627

0–126

9934

267

38

1

34.7

689

DE, 470

Sandvik

D45 KS

D to H

229

152

47627

0–126

9934

245

38

1

25.5

689

DE, 403

Sandvik

D55SP

D to H

254

172

79832

0–130

9934

232

35.4

1

56.63

689

DE, 597

Sandvik

D245S

D to H

203

127

35000

0–114

8282

209

21

1

34.7

689

DE, 470

Sandvik

D25 KS

D to H

171

127

33566

0–96

8282

142

32

1

32.8

689

DE, 470

Schramm

T450BH

D to H

171

114

21000

0–110

8518

110

49

1

29.7

2413

DE, 432

Appendix.indd 688

107

MEL, 500

11/22/2010 4:24:42 PM

For Largest Diameter Blasthole For maximum single pass combination

Nos.

Length each

Corresponding maximum hole depth

Tramming speed

Bailing velocity Indicator for for hard rock largest suitability hole (max. Bit with load/max. largest Hole dia.) comp.

No

m

m

km/h

kN/mm

m/min

1.77

1.370

2783

For maximum depth combination

Pipes in the Rack

Pipes in the Rack Corresponding maximum hole depth

Drill Pipe Diameter

m

m, mm

Drill pipe diameter

Single pass depth Nos.

mm

m

No, m

340

19.81

2

10.67

41.15

Same as for Max. Single Pass Combination

273

12.19

4

12.19

60.95

219

219

11.3

4

12.19

60.06


219

16.76

2

7.62

32


203

9.14

5

9.14

54.84

127

12.19

5

12.19

73.14

194

10.67

4

10.67

53.35

127

9.14

5

10.67

62.49

Kelly 178

18.29

0

0

18.29


178

9.14

2

9.14

27.42

114

9.14

5

9.14

178

9.14

2

9.14

27.42

114

9.14

5

9.14

127

7.9

5

7.6

45.9


Kelly 121

15.24

0

0

15.24

12.2 m 121 mm Kelly for Hole Depths Up to 12.2 m

340

19.81

2

9.14

38.09


1.45

340

21.3

2

9.14

38.09


273

16.5

1

30.2

220

16.5

5

340

20

2

9.14

38.28

273

20

5

340

16.5

1

13.41

29.91

273

16.5

235

16.46

2

7.62

31.7

235

16.5

235

13.4

1

12.19

25.59

235

235

10.3

1

10.67

20.96

178

11

1

10.67

178

10.1

1

178

8.6

178

Length each

Single pass depth

12.19

5

12.19

73.14

1.287

4223

1.77

1.240

3841

1.77

1.240

3841

1.77

1.151

3143

0.989

1943

0.975

2187

54.84

0.876

1319

54.84

0.788

1152

0.780

2476

0.625

1905

1.582

1662

1.45

1.545

2783

82.35

3.05

1.555

2289

9.14

65.7

2.09

1.283

4393

5

13.41

83.55

2.09

1.283

4393

4

7.62

46.98

2.01

1.322

2086

13.4

5

12.19

74.35

2.01

1.322

2086

235

10.3

5

10.67

63.65

2.01

1.322

2086

21.67

178

11

2

10.67

32.34

3.14

1.022

1486

10.67

20.77

178

10.1

4

10.67

52.78

3.14

1.022

1486

1

9.14

17.74

178

8.6

4

9.14

45.16

3.14

1.022

1486

15.85

2

7.62

31.09

178

15.8

2

7.62

31.04

2.09

0.874

178

12.8

1

10.67

23.47

178

12.8

4

10.67

55.48

2.09

0.874

1486

340

19.8

1

19.8

39.6


1.358

2813

273

19.8

1

19.8

39.6


1.396

2221

219

19.8

1

19.8

39.6


1.286

1948

178

19.8

1

19.8

39.6


1.156

1684

219

10.2

3

10.2

40.8


1.178

1327

127

9.14

13.7

13.7

1486

3

9.14

36.56


0.700

863

43.5


0.969

1693

152

11.5

4

8

140

14

1

14

28

140

8

6

8

56

0.842

2315

127

12

1

12

24

127

8

6

8

56

0.706

2692

200

8

3

8

32


1.090

1096

1.116

1449

219

10.67

3

10.67

42.68


127

10.67

2

6.1

22.87

127

6.1

5

6.1

36.6

0.971

2436

200

8

3

8

32


1.481

1548

200

11

2

11

33


1.400

1811

235

19

1

13

32


1.447

1166

235

17

1

17

34


1.296

2737

203

11

2

11

33


1.200

1914

381

21.3

1

21.3

42.6

391

273

19.81

2

19.8

59.41


273

12.2

5

12.2

73.2

235

1.335

2747

273

15.24

3

15.24

60.96


1.275

1220

200

10.67

3

10.67

42.68


1.152

1316

140

11

5

7.62

49.1


0.665

2028

200

8

3

8

32


1.296

1238

140

8.5

4

8.5

42.5


1.093

1196

273

19.81

0

0

19.81

273

12.2

6

12.2

85.4

1.499

2543

273

19.81

0

0

19.81

273

12.2

6

12.2

85.4

1.499

2289

244

16.76

2

7.62

32

244

12.8

3

12.8

51.2

1.430

1939

244

10.67

4

10.67

53.35


1.467

3939

178

9.14

4

9.14

45.7


1.166

2129

140

9.14

6

9.14

63.98


1.068

989

194

16.76

2

7.62

32


0.913

2682

140

8.53

4

9.14

45.09


1.030

2045

140

8.53

2

9.14

26.81


0.832

4332

102

7.62

5

7.62

45.72


0.643

2007

Appendix.indd 689

19.8 12.2

2 6

19.8 12.2

59.4 85.4

1.61

1.499

2625

1.9

1.335

2747

1.9

1.62

11/22/2010 4:04:42 PM

Table A10.2 Major details of truck mounted rotary blasthole drills.

Blasthole diameter

Manufacturer

Book.indb 690

Mode I

Largest compressors

Max.

Min.

Max. machine weight

Rotary speed range

Max. torque

Max. bit load

Max. feed speed

Nos.

Max. discharge

Pressure

Power of drill engine

mm

mm

kg

RPM

Nm

kN

m/min

Nos.

m3/min

kPa

kW 611

Atlas Copco

T4BH

254

191

26309

0-107

10575

133.4

14.63

1

35.4

2413

Reichdrill

T650

172

127

24948

0-108

6463

133.4

45.11

1

30.3

2585

433

Reichdrill

T750

311

222

45359

0-141

12924

311

61

1

48.1

689

522

Sandvik

DR320

172

115

24401

0-89

8204

143

41.4

1

28.3

2413

403

11/22/2010 2:43:59 PM

Largest drill pipe for largest blasthole Pipes in the rack

Book.indb 691

Indicator for hard rock suitability

Bailing velocity for largest hole with largest comp.

Carrier details

Drill pipe diameter

Single pass depth

Nos.

Length each

Maximum hole depth

mm

m

No

m

m

178

7.62

3

7.62

30.48

0.5252

1373

6×4 / 8×4

140

7.62

4

7.62

38.1

0.7755

3864

6×4

219

12.19

2

12.19

36.6

1.0000

1256

127

7.62

6

7.62

51.8

0.8313

2678

Overall drill dimensions (Mast Down)

Front axle capacity

Rear axle capacity

Length

Width

Height

kg

kg

m

m

m

Atlas Copco

9072

20865

10.7

2.4

4.12

Reichdrill

9072

20865

12.1

2.4

4.2

8×4

Reichdrill

18144

31751

18.5

4.3/Operating 5.6 5.6

6×4

Sterling LT9500

9072

20865

App 12.5

2.54

Type

Built by

m/min

4.1

11/22/2010 2:43:59 PM

Appendix 11

Details of drill pipes used for rotary blasthole drilling

Tables A11.1, A11.2 and A11.3 give dimension and weight details of rotary drill pipes commonly used inblasthole drilling. The weights presented are calculated from basic parameters such as dimension anddensity of steel etc. The first column of the tables indicates world-renowned manufacturers who make the drill pipes. Key for the manufacturers is as under 1 2 3 4 5

Book.indb 693

- Drillco International - Mining Technology International - Star Iron Works - Smith Drilling - Atlas Copco

11/22/2010 2:43:59 PM

Table A11.1 Fabricated drill pipes of diameter less than 250 mm.

Book.indb 694

Drill pipe O. diameter

Wall thickness

Cross sectional area

Moment of inertia

Producer

Inch

mm

In

mm

in2

mm2

in4

mm4

3 1, 2, 3 1, 2 1 1, 2, 3 1, 2, 3, 5 1, 2, 3 1, 3, 4, 5 5 3, 5 1, 2 5 1 4, 5 3 3 1, 2, 3 1, 3, 5 4 1, 2, 5 5 3 3, 5 1, 2, 4, 5 1 1 1, 2, 3, 5 1 5 5

3.5 4 4.5 4.5 5 5 5.5 5.5 6 6.25 6.25 6.25 6.5 6.625 6.75 7 7 7 7.5 7.625 7.625 7.625 8.625 8.625 8.625 9.25 9.25 9.25 9.25 9.625

88.9 101.6 114.3 114.3 127 127 139.7 139.7 152.4 158.75 158.75 158.75 165.1 168.28 171.45 177.8 177.8 177.8 190.5 193.68 193.68 193.68 219.08 219.08 219.08 234.95 234.95 234.95 234.95 244.48

0.75 0.5 0.375 0.5 0.5 0.75 0.5 0.75 0.75 0.5 0.75 1 0.75 0.864 0.75 0.5 0.75 1 1 0.75 0.875 1 0.906 1 1.5 0.75 1 1.5 1.5 1.5

19.05 12.7 9.525 12.7 12.7 19.05 12.7 19.05 19.05 12.7 19.05 25.4 19.05 21.946 19.05 12.7 19.05 25.4 25.4 19.05 22.225 25.4 23.012 25.4 38.1 19.05 25.4 38.1 38.1 38.1

6.48 5.498 4.86 6.283 7.069 10.014 7.854 11.192 12.37 9.032 12.959 16.493 13.548 15.637 14.137 10.21 14.726 18.85 20.42 16.199 18.555 20.813 21.97 23.955 33.576 20.028 25.918 36.521 36.521 38.288

4180.3 3546.9 3135.3 4053.7 4560.4 6460.5 5067.1 7220.6 7980.6 5827.1 8360.7 10640.9 8740.7 10088.5 9120.7 6587.2 9500.8 12161 13174.4 10450.8 11971 13427.7 14174.5 15454.6 21661.7 12921 16721.3 23561.9 23561.9 24702

6.6 8.6 10.4 12.8 18.1 23.3 24.8 32.4 43.5 37.6 49.9 58.9 56.9 66.3 64.6 54.2 72.9 87.2 110.4 96.8 107.5 116.8 165.9 177.1 222.5 182.3 223.7 284.5 284.5 326.7

2739124 3575547 4337824 5312241 7539295 9703778 10318006 13465764 18101205 15654509 20775203 24510372 23702043 27609704 26893219 22577023 30360222 36286690 45950882 40309899 44725060 48611153 69047485 73708936 92613522 75871823 93130220 118402950 118402950 135991287

11/22/2010 2:44:00 PM

Book.indb 695

Body weight

Drill pipe weight in kg for different lengths in m as under

lb/ft

kg/m

6.096

7.62

9.144

10.668

12.192

13.716

15.24

16.764

22.05 18.71 16.54 21.38 24.05 34.08 26.73 38.08 42.09 30.73 44.1 56.12 46.1 53.21 48.11 34.74 50.11 64.14 69.49 55.12 63.14 70.82 74.76 81.51 114.25 68.15 88.19 124.27 124.27 130.29

32.84 27.87 24.63 31.85 35.83 50.76 39.81 56.73 62.7 45.78 65.69 83.6 68.67 79.26 71.66 51.75 74.64 95.55 103.51 82.11 94.05 105.5 111.37 121.42 170.19 101.52 131.38 185.12 185.12 194.08

200 170 150 194 218 309 243 346 382 279 400 510 419 483 437 315 455 582 631 501 573 643 679 740 1037 619 801 1128 1128 1183

250 212 188 243 273 387 303 432 478 349 501 637 523 604 546 394 569 728 789 626 717 804 849 925 1297 774 1001 1411 1411 1479

300 255 225 291 328 464 364 519 573 419 601 764 628 725 655 473 683 874 946 751 860 965 1018 1110 1556 928 1201 1693 1693 1775

350 297 263 340 382 542 425 605 669 488 701 892 733 846 764 552 796 1019 1104 876 1003 1125 1188 1295 1816 1083 1402 1975 1975 2070

400 340 300 388 437 619 485 692 764 558 801 1019 837 966 874 631 910 1165 1262 1001 1147 1286 1358 1480 2075 1238 1602 2257 2257 2366

450 382 338 437 491 696 546 778 860 628 901 1147 942 1087 983 710 1024 1311 1420 1126 1290 1447 1528 1665 2334 1392 1802 2539 2539 2662

500 425 375 485 546 774 607 865 956 698 1001 1274 1047 1208 1092 789 1138 1456 1577 1251 1433 1608 1697 1850 2594 1547 2002 2821 2821 2958

551 467 413 534 601 851 667 951 1051 767 1101 1401 1151 1329 1201 868 1251 1602 1735 1376 1577 1769 1867 2035 2853 1702 2202 3103 3103 3254

11/22/2010 2:44:00 PM

696

Appendix 11

Table A11.2 Fabricated drill pipes of diameter more than 250 mm. Drill pipe O. diameter

Wall thickness


Moment of inertia

Producer

Inch

mm

In

mm

in2

mm2

in4

mm4

1, 2, 3, 4, 5 5 1, 2 1, 2 5 5 1, 2, 4 1, 2 4, 5 1, 2, 5 2 2 2 2 4 4

10.75 10.75 10.75 10.75 11.75 12.25 12.75 13.375 13.375 13.375 14 14 14.375 14.375 15 15

273.05 273.05 273.05 273.05 298.45 311.15 323.85 339.73 339.73 339.73 355.6 355.6 365.13 365.13 381 381

1 1.25 1.5 2 1.25 1 1 1 1.25 1.5 1.5 2 1.5 2 1.25 1.5

25.4 31.75 38.1 50.8 31.75 25.4 25.4 25.4 31.75 38.1 38.1 50.8 38.1 50.8 31.75 38.1

30.631 37.306 43.59 54.978 41.233 35.343 36.914 38.877 47.615 55.96 58.905 75.398 60.672 77.754 53.996 63.617

19761.6 24068.6 28122.2 35469.5 26602.1 22801.8 23815.2 25082 30719.1 36102.9 38003 48643.9 39143.1 50164 34836.1 41043.3

367.8 428.1 478.5 553.6 576.3 563.6 641.7 749.1 884.3 1002.1 1167.1 1394.9 1274.2 1527.3 1286.6 1467.2

153092137 178209075 199151563 230443981 239874485 234568629 267080564 311786071 368078982 417120238 485763550 580587047 530374401 635708568 535533882 610682934

Table A11.3 Integral drill pipes.

Book.indb 696

Drill pipe diameter d

Internal diameter


Moment of inertia

Producer

Inch

mm

In

mm

in2

mm2

in4

mm4

1 1 1 1 1 1 1 1 1 1 1 1 1

4 4.5 5 5.5 6.25 6.5 7 7.625 8.625 9.25 10.75 12.75 13.375

101.6 114.3 127 139.7 158.75 165.1 177.8 193.68 219.08 234.95 273.05 323.85 339.73

1.5 1.75 2.063 2.313 2.813 2.813 2.813 2.813 3.25 3.25 4.75 4.75 4.75

38.1 44.45 52.388 58.738 71.438 71.438 71.438 71.438 82.55 82.55 120.65 120.65 120.65

10.799 13.499 16.294 19.558 24.467 26.97 32.272 39.451 50.13 58.905 73.042 109.956 122.78

6967.2 8709 10512.2 12618.2 15785.1 17400.2 20820.5 25452.1 32342.2 38003 47123.8 70939 79212.6

12.3 19.7 29.8 43.5 71.8 84.6 114.8 162.9 266.2 353.9 630.6 1272.2 1545.9

5127078 8186645 12400083 18111984 29897870 35193430 47778077 67787495 110788796 147299752 262457893 529538457 643451068

11/22/2010 2:44:01 PM

Appendix 11

Body weight

Drill pipe weight in kg for different lengths in m as under

lb/ft

kg/m

12.192

13.716

15.24

16.764

18.288

19.812

20.117

21.336

104.23 126.94 148.33 187.08 140.31 120.26 125.61 132.29 162.02 190.42 200.44 256.56 206.45 264.58 183.74 216.47

155.26 189.1 220.95 278.68 209.01 179.15 187.11 197.06 241.35 283.65 298.58 382.18 307.54 394.13 273.7 322.47

1893 2306 2694 3398 2548 2184 2281 2403 2943 3458 3640 4660 3750 4805 3337 3932

2130 2594 3031 3822 2867 2457 2566 2703 3310 3891 4095 5242 4218 5406 3754 4423

2366 2882 3367 4247 3185 2730 2852 3003 3678 4323 4550 5824 4687 6007 4171 4914

2603 3170 3704 4672 3504 3003 3137 3304 4046 4755 5005 6407 5156 6607 4588 5406

2839 3458 4041 5096 3822 3276 3422 3604 4414 5187 5460 6989 5624 7208 5005 5897

3076 3746 4377 5521 4141 3549 3707 3904 4782 5620 5915 7572 6093 7809 5423 6389

3123 3804 4445 5606 4205 3604 3764 3964 4855 5706 6007 7688 6187 7929 5506 6487

3313 4035 4714 5946 4459 3822 3992 4204 5149 6052 6371 8154 6562 8409 5840 6880

Weight

Book.indb 697

697

Drill Pipe Weight in kg for Different Lengths in m as Under

lb/ft

kg/m

7.62

9.144

10.668

12.192

36.75 45.93 55.45 66.55 83.26 91.77 109.81 134.24 170.58 200.44 248.55 374.15 417.79

54.74 68.42 82.59 99.14 124.02 136.71 163.58 199.97 254.1 298.58 370.24 557.35 622.36

417 521 629 755 945 1042 1246 1524 1936 2275 2821 Not Made Not Made

501 626 755 907 1134 1250 1496 1829 2323 2730 3385 Not Made Not Made

584 730 881 1058 1323 1458 1745 2133 2711 3185 Not Made Not Made Not Made

667 834 1007 1209 1512 1667 1994 2438 3098 Not Made Not Made Not Made Not Made

11/22/2010 2:44:02 PM

Appendix 12

Details of stabilizers and other miscellaneous items

The details of various types of stabilizers given in this appendix are for the products made by the specific manufacturers mentioned in each case. There may be significant differences between such dimensions for similar products made by other manufacturers. Even for the same manufacturer the dimensions must be checked from the manufacturers before taking any important decision.

Shoulder to Shoulder Length

Drill Bit

Drill Pipe

Body Diameter

Blasthole Diameter

Figure A12.1 Sketch of Atlas Copco roller stabilizer.

Table A12.1 Details of Atlas Copco roller stabilizer. Blasthole diameter Body diameter

Book.indb 699

Sh. to Sh. length

Inches

mm

Inches

mm

Inches

mm

9.000 9.875 10.625 11.000 12.250 13.750 15 16

228.60 250.82 269.88 279.40 311.15 349.25 381 406.4

7 or 7.625 7.625 or 8.625 8.625 or 9.250 8.625 or 9.250 9.625 or 10.750 10.750 to 11.750 11.750 to 13.375 13.375

177.8 or 193.7 193.7 or 219.1 219.1 or 235 219.1 or 235 244.5 to 273.1 273.1 or 298.5 298.5 or 339.7 339.7

32 or 42 32 or 42 or 50 32 or 42 or 50 32 or 42 or 50 32 or 42 or 50 32 or 42 or 50 42 or 50 42 or 50

812.8 or 1066.8 812.8 or 1066.8 or 1270 812.8 or 1066.8 or 1270 812.8 or 1066.8 or 1270 812.8 or 1066.8 or 1270 812.8 or 1066.8 or 1270 1066.8 or 1270 1066.8 or 1270

11/22/2010 2:44:03 PM

700

Appendix 12


Drill Pipe

Body Diameter

Drill Bit

Blasthole Diameter

Figure A12.2 Sketch of MTI roller stabilizer. Table A12.2 Details of MTI roller stabilizer. Blasthole diameter

Body diameter

Inches

mm

Inches

6.250 6.750 7.875

158.75 171.45 200.025

9.000 9.875 10.625

228.6 250.825 269.875

11.000

279.4

12.250 13.750 15.000

311.15 349.25 381

5.125 5.625 6.250 or 7.000 7.625 8.625 8.625 or 9.25 9.250 or 9.625 10.750 12.250 13.375 or 14.375

Sh. to Sh. length

Weight

mm

Inches

lbs

130.18 142.88 158.75 or 177.8 193.68 219.08 219.08 or 234.95 234.95 or 244.48 273.05 311.15 339.72 or 365.12

26.500 26.500 26.500

673.1 673.1 673.1

26.500 30.750 30.750

673.1 781.05 781.05

30.750

781.05

30.750 41.000 41.000

781.05 1041.4 1041.4

130 165 200 or 220 290 400 400 or 460 460 or 510 650 1000 1300

kg 59 75 91 132 182 182 or 209 209 or 232 295 454 590


Body Diameter

Blasthole Diameter

Figure A12.3 Sketch of MTI welded straight blade stabilizer.

Book.indb 700

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Appendix 12

701

Table A12.3 Details of MTI welded straight blade stabilizer. Blasthole Diameter

Body Diameter

Inches

mm

Inches

6.250 6.750 7.875

158.75 171.45 200.025

9.000 9.875 10.625

228.6 250.825 269.875

11.000

279.4

12.250 13.750 15.000

311.15 349.25 381

5.125 5.625 6.250 or 7.000 7.625 8.625 8.625 or 9.25 9.250 or 9.625 10.750 12.250 13.375 or 14.375

Sh. to Sh. Length

Weight

mm

Inches

lbs

130.18 142.88 158.75 or 177.8 193.68 219.08 219.08 or 234.95 234.95 or 244.48 273.05 311.15 339.72 or 365.12

26.500 26.500 26.500

673.1 673.1 673.1

145 185 250

66 84 113

28.500 30.750 30.750

673.1 781.05 781.05

350 470 520

159 213 236

30.750

781.05

580

263

30.750 41.000 41.000

781.05 1041.4 1041.4

790 1100 1500

358 499 680

kg

A C

D

B E

F H

Figure A12.4 Sketch of IDS welded straight blade stabilizer.

Table A12.4 Details of IDS welded straight blade stabilizer. H

Book.indb 701

A

B

C

D

E

F

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

in

Mm

6.250 6.750 7.875 9.000 9.875 10.625 11.000 12.250 13.750 15.000

158.75 171.45 200.025 228.6 250.825 269.875 279.4 311.15 349.25 381

60 60 61 61 61 64 64 64 66 66

1524 1524 1549.4 1549.4 1549.4 1625.6 1625.6 1625.6 1676.4 1676.4

28 28 28 28 28 28 28 28 28 28

711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2

18 18 18 18 18 18 18 18 18 18

457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2

14 14 15 15 15 18 18 18 20 20

355.6 355.6 381 381 381 457.2 457.2 457.2 508 508

12 12 12 12 12 12 12 12 14 14

304.8 304.8 304.8 304.8 304.8 304.8 304.8 304.8 355.6 355.6

1.5 1.5 2 2 2 2 2 2 2 2

38.1 38.1 50.8 50.8 50.8 50.8 50.8 50.8 50.8 50.8

11/22/2010 2:44:04 PM

702

Appendix 12

A C

D

B E

F

H

Figure A12.5 Sketch of IDS welded spiral blade stabilizer.

Table A12.5 Details of IDS welded spiral blade stabilizer. H

A

B

C

D

E

F

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

6.250 6.750 7.875 9.000 9.875 10.625 11.000 12.250 13.750 15.000

158.75 171.45 200.025 228.6 250.825 269.875 279.4 311.15 349.25 381

60 60 61 61 61 64 64 64 66 66

1524 1524 1549.4 1549.4 1549.4 1625.6 1625.6 1625.6 1676.4 1676.4

28 28 28 28 28 28 28 28 28 28

711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2 711.2

18 18 18 18 18 18 18 18 18 18

457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2 457.2

14 14 15 15 15 18 18 18 20 20

355.6 355.6 381 381 381 457.2 457.2 457.2 508 508

12 12 12 12 12 12 12 12 14 14

304.8 304.8 304.8 304.8 304.8 304.8 304.8 304.8 355.6 355.6

1.5 1.5 2 2 2 2 2 2 2 2

38.1 38.1 50.8 50.8 50.8 50.8 50.8 50.8 50.8 50.8

F D

G

E

C

B

A

H

Figure A12.6 Sketch of integral spiral blade stabilizer.

Book.indb 702

11/22/2010 2:44:06 PM

Book.indb 703

Table A12.6 Details of integral spiral blade stabilizer. A

B

C

D

E

H

F

G

Weight

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

lb

kg

6.25 6.75 7.875 9 9.875 10.625 11 12.25 13.75 15 17.5

158.75 171.45 200.025 228.6 250.825 269.875 279.4 311.15 349.25 381 444.5

4.75 4.75 6.75 8 8 8 8 9 10 10 10

120.65 120.65 171.45 203.2 203.2 203.2 203.2 228.6 254 254 254

2 2.25 2.25 2.25 2.8125 2.25 2.8125 2.8125 2.8125 2.8125 3

50.8 57.15 57.15 57.15 71.4375 57.15 71.4375 71.4375 71.4375 71.4375 76.2

28 28 28 28 28 30 30 29 31 31 31

711.2 711.2 711.2 711.2 711.2 762 762 736.6 787.4 787.4 787.4

18 17 18 18 18 20 20 19 20 20 20

457.2 431.8 457.2 457.2 457.2 508 508 482.6 508 508 508

2 2 2.5 2.5 2.5 3.25 3.25 3.25 4.5 4.5 4.5

50.8 50.8 63.5 63.5 63.5 82.55 82.55 82.55 114.3 114.3 114.3

62 62 64 63 63 69 69 69 73 73 74

1574.8 1574.8 1625.6 1600.2 1600.2 1752.6 1752.6 1752.6 1854.2 1854.2 1879.6

13 13.25 15 15.25 14 14.75 14 16 17.5 16 14.75

330.2 342.9 381 387.35 355.6 374.65 355.6 406.4 444.5 406.4 374.65

280 370 620 840 840 1000 1000 1550 1670 1670 2100

127.0 167.8 281.2 381.0 381.0 453.6 453.6 703.1 757.5 757.5 952.5

Table A12.7 Details of integral straight blade stabilizer. A

B

C

D

E

H

F

G

Weight

11/22/2010 2:44:08 PM

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

in

mm

lb

kg

6.25 6.75 7.875 9 9.875 10.625 11 12.25 13.75 15 17.5

158.75 171.45 200.025 228.6 250.825 269.875 279.4 311.15 349.25 381 444.5

4.75 4.75 6.75 8 8 8 8 9 10 10 10

120.65 120.65 171.45 203.2 203.2 203.2 203.2 228.6 254 254 254

2 2.25 2.25 2.25 2.8125 2.25 2.8125 2.8125 2.8125 2.8125 3

50.8 57.15 57.15 57.15 71.4375 57.15 71.4375 71.4375 71.4375 71.4375 76.2

28 28 28 28 28 30 30 29 31 31 31

711.2 711.2 711.2 711.2 711.2 762 762 736.6 787.4 787.4 787.4

18 17 18 18 18 20 20 19 20 20 20

457.2 431.8 457.2 457.2 457.2 508 508 482.6 508 508 508

2 2 2.5 2.5 2.5 3.25 3.25 3.25 4.5 4.5 4.5

50.8 50.8 63.5 63.5 63.5 82.55 82.55 82.55 114.3 114.3 114.3

62 62 64 63 63 69 69 69 73 73 74

1574.8 1574.8 1625.6 1600.2 1600.2 1752.6 1752.6 1752.6 1854.2 1854.2 1879.6

13 13.5 15 15.25 14 14.75 14 16 17.5 16 14.75

330.2 342.9 381 387.35 355.6 374.65 355.6 406.4 444.5 406.4 374.65

280 370 620 840 840 1000 1000 1550 1670 1670 2100

127.0 167.8 281.2 381.0 381.0 453.6 453.6 703.1 757.5 757.5 952.5

Book.indb 704

Sleeve Body Diameter

Sleeve Bore Blade Width Diameter Sleeve End Diameter Sleeve

Mandrel Upset Diameter

Fishing Neck Diameter

Mandrel

Crown Length Sleeve End Length

Sleeve Length

Fishing Neck Length

Total Overall Length Blasthole Diameter

Bore ID (Near

Bore ID (String)

Figure A12.8 Sketch of replaceable sleeve stabilizer.

11/22/2010 2:44:09 PM

Book.indb 705

Table A12.8 Details of sleeve of the DTI replaceable sleeve stabilizer. Blasthole diameter

Sleeve Body diameter

Sleeve ID

Sleeve size designation

in

mm

in

mm

in

475 625 775 775 963

6.25–6.75 8.375–9.875 8.875–12.25 9.625–17.5 12.25–17.5

158.75–171.45 212.725–250.825 225.425–311.15 244.475–444.5 311.14–444.5

5.75 7.5 9.25 10 11

146.05 190.5 234.95 254 279.4

4.75 6.25 7.75 7.75 9.625

Sleeve length

Blade width

Minimum crown length

Weight of sleeve

mm

in

mm

in

mm

in

mm

lb

kg

120.65 158.75 196.85 196.85 244.475

14 14 18 18 18

355.6 355.6 457.2 457.2 457.2

2 2.5 3 3 3.5

50.8 63.5 76.2 76.2 88.9

12.2 11.1 14.6 11.1 11.9

309.88 281.94 370.84 281.94 302.26

43 84 162 294 253

19.50 38.10 73.48 133.36 114.76

Table A12.8 A Details of mandrel of the DTI relatable sleeve stabilizer. Maximum fishing neck diameter

Mandrel size designation in

mm

Mandrel upset diameter

Sleeve end diameter

Fishing neck length

Total Sleeve end overall length length

Bore ID (near bit)

Bore ID (string)

in

in

mm

in

in

in

mm

in

mm 57.15

mm

mm

mm

in mm

Weight of mandrel lb

kg

47

4.75–5

120.65– 5.125– 127 5.75

130.175– 4.75 146.05

120.65

23 584.2 32 812.8 62 1574.8 1.5

38.1

2.25

302 136.98

62

6.5–7

165.1– 177.8

7.5

190.5

6.25

158.75

23 584.2 32 812.8 62 1574.8 2.25

57.15

2.8125 71.4375

500 226.80

77

8.25

209.55

9.25

234.95

7.75

196.85

27 685.8 37 939.8 71 1803.4 2.8125 71.4375 2.8125 71.4375

893 405.06

96

10

254

11

279.4

9.625 244.475 27 685.8 37 939.8 71 1803.4 2.8125 71.4375 2.8125 71.4375 1400 635.03

11/22/2010 2:44:10 PM

Table A12.9 Weights of thread protectors.

Book.indb 706

Weight of pin

Weight of box

Weight of pair

Connection type

lb

kg

lb

kg

lb

kg

3 1/2 API Regular 4 1/2 API Regular 5 1/2 API Regular 6 5/8 API Regular 7 5/8 API Regular 8 5/8 API Regular

7 17 21 28 35 55

3.18 7.71 9.53 12.70 15.88 24.95

8 14 18 23 31 46

3.63 6.35 8.16 10.43 14.06 20.87

15 31 39 51 66 101

6.80 14.06 17.69 23.13 29.94 45.81

11/22/2010 2:44:11 PM

Appendix 13

Details of shock absorbers

Clearance Diameter Pin Connection to Drive Head

Shoulder to Shoulder Length Rubber Element Box Connection to Drill Pipes

Figure A13.1 Atlas Copco Shock Absorber.

Table A13.1 Details of Atlas Copco shock absorbers.

Book.indb 707

Model designation

Max. Pull down load in kg

Clearance diameter in mm

Shoulder to shoulder length in mm

Assembly weight in kg

28 22 18 15

68038 38555 27215 15875

863.6 685.8 558.8 508

914.4 762 711.2 711.2

1044 578.5 379 295

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708

Appendix 13

Clearance Diameter Pin Connection to Drive Head

Element Assembly Shoulder to Shoulder Length

Box Connection to Drill Pipes

Figure A13.2 MTI Shock Absorber. Table A13.2

Book.indb 708

Details of MTI shock absorbers.

Model designation

Max. pull down load in kg

Clearance diameter in mm

Shoulder to shoulder length in mm

Assembly weight in kg

130 100 75 50 40

45359–68038 36287–45359 31751–36287 22680–31751 15875–22680

660.4 609.6 457.2 457.2 457.2

927.1 863.6 812.8 838.2 762

953 850 644 508 363

11/22/2010 2:44:12 PM

Appendix 14

Properties of some rocks

Properties of rocks differ considerably from place to place. Further, there is no standard about the nomenclature of the rocks. For this reason, a range of values is often given in the following table rather than a specific value. It is also quite likely that the values for properties for a rock with a particular name as presented in different tables differ considerably. This can happen because the properties have been arrived at after testing samples obtained from different places. Therefore, the tables should not be amalgamated. When a very important decision is to be taken several rock samples from the worksite should be properly tested in a well-equipped laboratory and the values found by such testing should be used for taking the decision.

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710

Appendix 14

Table A14.1 Properties of certain rocks. Schmidt Cercher P wave hardness abrasivity velocity index index m/s

Rock type

Name of the rock

Dry density g/cm3

Igneous

Granite

2.53–2.62

1.02–2.87 54–69

4.5–5.3

Diorite Gabbro Rhyolite Andesite Basalt

2.80–3.00 2.72–3.00 2.40–2.60 2.50–2.80 2.21–2.77

0.10–0.50 1.00–3.57 0.40–4.00 0.20–8.00 67 0.22–22.1 61

4.2–5.0 3.7–4.6

Conglomerate Sandstone Shale Mudstone Dolomite Limestone Gneiss Schist

2.47–2.76 1.91–2.58 2.00–2.40 1.82–2.72 2.20–2.70 2.67–2.72 2.61–3.12 2.60–2.85

1.62–26.4 10–37 20.0–50.0 27 0.20–4.00 0.27–4.10 35–51 0.32–1.16 49 10.0–30.0 31

Phyllite Slate Marble Quartzite

2.18–3.30 2.71–2.78 2.51–2.86 2.61–2.67

1.84–3.64 0.65–0.81 0.40–0.65

Sedimentary

Metamorphic

Porosity %

2.7–3.8 2.0–3.5 1.5–3.8 1.5–4.2 0.6–1.8

1.0–2.5 3.5–5.3 2.2–4.5

2.3–4.2

S wave velocity m/s

4500–6500 3500– 3800 4500–6700 4500–7000 4500–6500 5000–7000 3660– 3700 1500–4600 2000–4600 5500 3500–6500 5000–7000 6100–6700 3460– 4000 3500–4500 5000–6000

4.3–5.9

Table A14.2 Properties of certain rocks.

Book.indb 710

Name of the rock

Porosity

Compressive strength MPa

Tensile strength MPa

Punch shear Strength MPa

Cohesion MPa

Angle of internal friction °

Young’s modulus GPa

Elastic

Density

Strength properties

Specific gravity

Physical properties

Quartz Granite Dolerite Sandstone Limestone (Grade I) Limestone (Grade II)

2.66 2.76 2.84 2.06 2.65 2.04

2.58 2.61 2.7 2.45 2.7 2.63

0.2 0.77 3.44 16.87 11.23 15.52

188.89 169.8 81.95 44.96 59.92 47.2

8.69 9 6.93 4.99 6.35 5.2

25.4 20.63 13.29 8.44 12.79 11.55

34.5 32 20 18 14 6

63 56 48 42 46 40

102 92 58 41.6 47.5 35

11/22/2010 2:44:12 PM

Appendix 14

Poisson’s ratio

Strain at failure MPa

Point load index

Fracture Mode I Toughness MPa

30–70

0.17

0.25

5–15

0.11–0.41

7–30 7–30 5–10 5–15 10–30

30–100 40–100 10–50 10–70 40–80

0.10–0.20 0.30 0.20–0.35 0.30 0.20–0.40 0.20 0.10–0.20 0.35

6–15

>0.41 >0.41

10–15 9–15

>0.41

10−11–10−9 10−12–10−11 10−13–10−10 10−14–10−12 10−11–10−8

30–230 20–170 5–100 10–100 20–120 30–250 100–250 70–150

3–10 4–25 2–10 5–30 6–15 6–25 7–20 4–10

10–90 15–50 5–30 5–70 30–70 20–70 30–80 5–60

0.10–0.15 0.14 0.10 0.15 0.15 0.30 0.24 0.15–0.25

10−14–10−12 10−14–10−11 10−14–10−13

5–150 50–180 50–200 150–300

6–20 7–20 7–20 5–20

10–85 20–90 30–70 50–90

0.26 0.20–0.30 0.35 0.15–0.30 0.40 0.17 0.20

Tensile Coefficient of UC strength strength permeability MPa MPa

Elastic modulus MPa

10−14–10−12

100–300

7–25

10−14–10−12 10−14–10−12 10−14–10−12 10−14–10−12 10−14–10−12

100–350 150–250 80–160 100–300 100–350

10−10–10−8 10−10–10−8

Book.indb 711

1–8

0.15 0.17

0.027–0.041 0.027–0.041

0.1–6

0.12

3–7 5–15 5–10

0.027–0.041 0.11–0.41 0.005–0.027

1–9 4–12 5–15

0.027–0.041 0.11–0.41 >0.41

Longitudinal wave velocity m/s

Shear wave velocity m/s

Shore hardness

Vickers hardness

Micro bit drilling hardness mm/min

Micro bit drilling abrasivity Wt. Loss x 10000

Cercher’s index

Quartz percent %

Protodyanokov index

Index

Poisson’s ratio

Dynamic

0.16 0.20

711

0.26 0.33 0.3 0.28 0.24 0.3

5225 4350 3270 2000 3200 3016

4059 2851 2430 850 1430 1280

82 76 46 41 35.27 26.3

710 630 330 285 240 180

115 100 65 85 37 22.5

0 0 0.38 0.85 0.76 1.4

6.8 6.1 5 5.6 4.8 3.4

100 35–40 0 40–50 15–20 10

20 20 10.4 3.4 8.24 6.3

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712

Appendix 14

Table A14.3 Typical densities of certain rocks. Rock name (type)

Density in kg/m3

Rock name (type)

Density in kg/m3

Andesite (I)

2.61

Dolomite (S)

2.30

Basalt (I)

2.99

Limestone (S)

2.11

Diabase (I)

2.91

Sandstone (S)

2.24

Diorite (I)

2.85

Shale (S)

2.10

Gabbro (I)

3.03

Av. Sedimentary Rocks

2.19

Granite (I)

2.64

Amphibolite (M)

2.96

Peridotite (I)

3.15

Gneiss (N)

2.80

Porphyry (I)

2.74

Granulite (M)

2.63

Pyroxenite (I)

3.17

Phyllite (M)

2.74

Av. Igneous Rocks

2.69

Quartzite (M)

2.60

Av. Acidic Igneous Rocks

2.61

Schist (M)

2.64

Av. Basic Igneous Rocks

2.79

Serpentine (M)

2.78

Clay (S)

1.70

Slate (M)

2.79

Coal (S)

1.35

Av. Metamorphic Rocks

2.76

Table A14.4 Compressive and tensile strengths of certain rocks. Compressive strength

Book.indb 712

Tensile strengt h

Rock name

lb/in2

MPa

lb/in2

MPa

Quartzite Quartzite Quartzite Argilite Diabase Basalt Basalt Basalt Gabbro Gabbro Granite Granite Granite Marble Limestone Limestone Dolomite Hornblende Schist

31650 22250 43700 31400 53300 9800 26500 40800 29600 25050 24350 22000 28950 18150 14200 17800 13800 29600

218.22 153.41 301.30 216.50 367.49 67.57 182.71 281.31 204.08 172.71 167.89 151.68 199.60 125.14 97.91 122.73 95.15 204.08

2510 2550 2950 2620 3550 730 1990 4020 2150 1810 1780 1300 1850 1010 820 910 600 1080

17.31 17.58 20.34 18.06 24.48 5.03 13.72 27.72 14.82 12.48 12.27 8.96 12.76 6.96 5.65 6.27 4.14 7.45

11/22/2010 2:44:13 PM

Table A14.5 Densities of certain minerals. Mineral name

Chemical formula

Density in kg/m3

Anhydrite Apatite Arsenopyrite Asbestos Augite Barite Bauxite Beryl Biotite Calcite Cassiterite Celestite Chalcocite Chalcopyrite Chlorite Chromatite Chromite Cinnabar Corundum Cuprite Diopside Dioptase Dolomite Emerald Fayalite Fluorite Forsterite Franklinite Galena Gibbsite Graphite Gypsum Halite Hematite Hornblende Hypersthene Illite Ilmenite Jacobsite Limonite Magnesite Magnetite Montmorillonite Olivine Opal Orthoclase Orthoferrosilite

CaSO4 Ca5(PO4)3(F,Cl,OH ) FeAsS Mg3Si2O5(OH)4 (Ca,Na)(Mg,Fe,Al,Ti)(Si,Al)2O6 BaSO4 AlO(OH) Be3 Al2Si6O18 K(Mg,Fe++)3[AlSi3O10(OH,F)2 CaCO3 SnO2 SrSO4 Cu2S CuFeS2 Na0.5(Al,Mg)6(Si,Al)8O18(OH)12•5(H2O) CaCrO4 Fe++Cr2O4 HgS Al2O3 Cu2O CaMgSi2O6 CuSiO2(OH)2 CaMg(CO3)2 Ni3(CO3)(OH)4•4(H2O) Fe++2SiO4 CaF2 Mg2SiO4 (Zn,Mn++,Fe++)(Fe+++,Mn+++)2O4 PbS Al(OH)3 C CaSO4.2H2O NaCl Fe2O3 Ca2[Mg4(Al,Fe+++)]Si7 AlO22(OH)2 (Mg,Fe++)2Si2O6 (K,H3O)(Al,Mg,Fe)2(Si,Al)4O10[(OH)2,(H2O)] Fe++TiO3 (Mn++,Fe++,Mg)(Fe+++,Mn+++)2O4 Fe+++O(OH) MgCO3 Fe++Fe+++2O4 (Na,Ca)0.3(Al,Mg)2Si4O10(OH)2•n(H2O) (Mg,Fe)2SiO4 SiO2•n(H2O) KAlSi3O8 (Fe++,Mg)2Si2O6

2.96–2.98 3.16–3.22 6.07 2.53 3.2–3.6 4.48 3–3.07 2.63–2.9 2.8–3.4 2.71 6.8–7 3.9–4 5.5–5.8 4.1–4.3 2.42 3.142 4.5–5.09 8.1 4–4.1 6.11 3.25–3.55 3.28–3.35 2.8–2.9 2.6 4.39 3.01–3.25 3.21–3.33 5.07–5.22 7.2–7.6 2.3–2.4 2.09–2.33 2.3 2.17 5.3 3–3.47 3.2–3.9 2.75 4.72 4.75 3.3–4.3 3.21 5.1–5.2 2–2.7 3.27–3.37 1.9–2.3 2.56 3.88–4.02

(Continued)

Book.indb 713

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Table A14.5 (Continued)

Book.indb 714

Mineral name

Chemical formula

Density in kg/m3

Orthopyroxenes Pentlandite Plagioclase Pyrite Pyrrhotite Quartz Rutile Serpentinite Siderite Sphalerite Sylvite Talc Titanite Tourmaline Troilite Wolframite Zircon

(Fe,Mg)2SiO3 (Fe,Ni)9S8 (Na,Ca)(Si,Al)4O8 FeS2 Fe1-XS SiO2 TiO2 Mg3Si2O5(OH)4 Fe++CO3 (Zn,Fe)S KCl Mg3Si4O10(OH)2 CaTiSiO5 NaFe+++3 Al6(BO3)3Si6O21F FeS (Fe,Mn)WO4 ZrSiO4

3.59 4.6–5 2.61–2.76 5–5.02 4.58–4.65 2.6–2.65 4.25 2.53–2.65 3.96 3.9–4.2 1.99 2.7–2.8 3.45–3.6 3.31 4.58–4.65 7.75 4–4.8

11/22/2010 2:44:14 PM

Book.indb 715

Table A14.6 Properties of certain rocks. Compressive strength Rock name

Specific gravity

Amphibolite Andesite Argillite Basalt Chert, dolomite Conglomerate Diabase Diorite Diorite, augite Dolomite Gabbro Granite Granite, aplitic Granite, gneissic Granite, pre-Cambrian Granodiorite Greenstone Hematite ore Hornfels Limestone Limestone, chalky

3.07 2.81 2.81 2.94 2.67 2.67 2.94 3.01 2.74 2.60 3.00 2.66 2.65 2.66 2.80 2.74 3.02 5.07 3.19 2.68 1.89

lb/in2

MPa

61335 56535 19720 44950 29290 23925 46545 39730 48285 18995 44805 37700 51185 30305

423 183 136 310 202 165 321 274 333 131 309 260 353 209

36540 39005 88015 77285 22330 4205

252 269 607 533 154 29

Poisson’s ratio

0.14

0.29 0.25 0.18 0.33 0.2 0.26 0.02 0.27 0.24

0.28 0.02

Engineering classification of intact rock based on

Modulus of rigidity

Young’s modulus

lb/in2

MPa

lb/in2

GPa

UCS

Compressibility

6641000 3944000

45800 27200

4596500 3436500 4698000 5408500 6119000 4886500 2900000 6394500 3422000 4756000 1299200 7583500 4060000 6104500

31700 23700 32400 37300 42200 33700 20000 44100 23600 32800 8960 52300 28000 42100 40900 26500 5380

104.0 64.6 84.1 77.9 56.2 77.9 95.8 107.0 84.1 47.6 119.0 59.2 80.6 18.6 82.1 68.6 105.0 200.0 95.8 68.1 11.1

A,Very High B, High B, High A,Very High B, High B, High A,Very High A,Very High A,Very High B, High A,Very High A,Very High A,Very High B, High

5930500 3842500 780100

15080000 9367000 12194500 11295500 8149000 11295500 13891000 15515000 12194500 6902000 17255000 8584000 11687000 2697000 11904500 9947000 15225000 29,000,000 13891000 9874500 1609500

1 – Low 2 – Medium 1 – Low 2 – Medium 2 – Medium 2 – Medium 1 – Low 1 – Low 1 – Low 2 – Medium 1 – Low 2 – Medium 2 – Medium 3 – High 2 – Medium 2 – Medium 1 – Low 1 – Low 1 – Low 2 – Medium 3 – High

A,Very High A,Very High A,Very High A,Very High B, High D, Low

11/22/2010 2:44:14 PM

Appendix 15

Bibliography

Table A15.1 List of books, presentations, theses and reports covering drilling and blasting. Name of the book/ Sr. article/product no. literature

Publication Author/editor

Pages

Publisher

Year Type

1 Advanced Drilling Techniques

Maurer W. C.

698

Petroleum Publishing

1980 Book

2 Air and Gas Drilling

Lyons W. C.

656

McGraw Hill Companies Inc.

2000 Book

3 Air Drilling Handbook

Van Omer H.

167

International Society for Explosives Engineers

1987 Book

4 Applied Geomechanics in Drilling

Vaitenko V. S.

266

A. A. Balkema

1995 Book

5 Atlas Copco Manual

Editorial Committee

656

Atlas Copco

1982 Book

6 Bench Drilling Techniquesand Equipment Selection Manual

Ed – Drake R.

92

Ingersoll Rand Company

1990 Book

7 Blaster’s Handbook

Hopler R. B.

744

International Society for Explosives Engineers

1998 Book

8 Blasthole Bit Handbook

Anonymous

28

Hughes Tool Company

1975 Book

9 Blasthole Drilling in Open Pit Mining

Ed. – Ulf Linder, Diana Norwood et al.

208

Atlas Copco Drilling Solutions LLC

2009 Book

10 Blasthole Technology

Anonymous

58

Reed Tool Company

1972 Book

11 Blasting in Opencast Mines and Quarries

Anonymous

24

Indian Explosives Ltd.

1972 Book

12 Blasting Principles for Open Pit Mining (Vol. 1 & 2)

Hustrulid W.

1038

A. A. Balkema

1999 Book

13 Bucyrus Erie Answers Your Questions About Blasthole Drills

Nelmark J.

8

Bucyrus Erie Company

1981 Book

(Continued)

Book.indb 717

11/22/2010 2:44:15 PM

Table A15.1 (Continued) Name of the book/ Sr. article/product no. literature

Author/editor

Pages

Publisher

14 Coal Mining in India

Mathur S. P.

474

M. S. Enterprise

1999 Book

15 Compressed Air and Gas Data

Gibbs C. W.

800

Ingersoll Rand Company

1976 Book

16 Compressed Air and Gas Handbook

Rollins J. P.

903

Prentice Hall

1989 Book

17 Construction Planning, Equipment and Methods

Schexngyder C. J., Peurifoy W. B., Ledbetter W. B.

688


2002 Book

18 Damage and Fracture in Heterogeneous Materials

Mishnaevsky L. L. Jr.,

230

A. A. Balkema

1998 Book

19 Design and Application Data for Varel Drill Bits

Anonymous

24

Varel Manufacturing Co.

1978 Book

20 Diamond Drill Handbook

Cumming

547

J. K. Smit & Sons Diamond Products Ltd.

1975 Book

21 Diamond Drilling Handbook

Heinz W. F.

538

A. A. Balkema

1992 Book

22 Dictionary of Drilling and Borehole

Moureau M.

436

French and European Publications, Inc.

1990 Book

23 Drillability – Drilling Rate Index Catlog

Bruland A., Eriksen S., ohannessen O., Sandberg B.

177

Dept. of Geology, Norwegian Inst. of Technology

1990 Book

24 Drilling and Excavation Technologies for the Future

Several Authors

176

National Academy Press

1994 Book

25 Drilling Practices Manual

Moore P. L.

448

Petroleum Publishing Co.

1974 Book

26 Drilling Technology

Vozniak J. P.

357

ASME

1995 Book

27 Engineering Geology Field Manual

Anonymous

491

U.S. Department of Interior, Bureau of Reclamation

2001 Book

28 Excavation Handbook

Church H. K.

1024


1980 Book

29 Explosives and Rock Blasting

Morhard R. C.

662

Atlas Powder Company

1987 Book

30 Face Drilling

Ed – Mike Smith

160

Atlas Copco Rock Drills AS

2004 Book

31 Geology

Smith A. J.

128

Hamlyn

1974 Book

32 Handbook on Surface Drilling and Blasting

Editorial Committee

320

Tamrock

1984 Book

Publication Year Type

(Continued)

Book.indb 718

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Pages

Publisher

Year Type

33 Hard Rock Miner’s Handbook

de la Vergne J.

330

McIntosh Engineering Limited.

2003 Book

34 History of Oilwell Drilling

Brantly J. E.

1525

Gulf Publishing

1971 Book

35 Improving Compressed Air System Performance

Several Authors

128

U.S. Department of Energy

2003 Book

36 Introductory Mining Engineering

Hartman H. L., Mutamansky J. M.

584

John Wiley & Sons

2002 Book

37 Mine Planning and Equipment Selection 1998

ED – Singhal P. K.

830

A. A. Balkema

1998 Book

38 Mining Engineers Handbook

Ed – Hartman H.

2394

Society for Mining Metallurgy & Exploration

2002 Book

39 Modern Techniques of Rock Blasting

Langfors U., Kihlstrom B.

406

J. Wiley and Sons

1973 Book

40 Novel Drilling Techniques

Maurer W. C.

114

Pergamon Press

1969 Book

41 Open Cut Blasthole Drilling

Anonymous

218

The Australian Drilling Industry Training Committee Limited

1987 Book

42 Open Pit Mine Planning and Design Vol. 1&2

Hustrulid W., Kuchta M.

864

A. A. Balkema

1995 Book

43 Petroleum Engineering Drilling and Well Completion

Gatlin C.

341

Prentice Hall

1960 Book

44 Physical Properties of Rocks and Minerals

Touloukian Y. S., Ho C.Y.

548


1981 Book

45 Pneumatic Handbook

Barber A.

659

Elsevier Press

1997 Book

46 Pneumatic Handbook

Warring R. H

433

Guld Publishing Co.

1982 Book

47 Practical Handbook of Physical Properties of Rocks and Minerals

Carmichael R. S.

760

CRC Press

1990 Book

48 Preventive Maintenance for Rock Drills

Anonymous

107

Compressed Air and Gas Inst.

1963 Book

49 Principles of Rock Drilling

Clark G. B.

91

Colorado School of Mines

1979 Book

50 Principles of Rock Drilling

Rao Karanam U. M., 272 Misra B.

A. A. Balkema

1998 Book

(Continued)

Book.indb 719

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Pages

Publisher

Year Type

51 Principles of Rock Fragmentation

Clark G. B.

610

John Wiley & Sons

1987 Book

52 Reciprocating and Rotary Compressors

Chlumsky V.

534

E & FN Spon Ltd.

1965 Book

53 Rock Bit Design, Selection and Evaluation

Bentson H. G

36

Smithtool

1966 Book

54 Rock Blasting Terms and Symbols

Ed – Rustan Agne

204

A. A. Balkema

1998 Book

55 Rock Drilling Data

Anonymous

224

Padley & Venables Ltd.

1974 Book

56 Rock Drilling Manual – Drill Steel Applications

Anonymous

84

Sandvik – Atlas Copco

1980 Book

57 Rock Drilling Manual – Theory and Technique

Anonymous

44

Sandvik – Atlas Copco

1977 Book

58 Rock Excavation

Dessureault, Sean

217

Dept. of Mining and Geological Engineering, Univ. of Arizona

2003 Book

59 Rock Excavation Handbook for Civil Engineering

Ed – Matti Heinio

364

Sandvik Tamrock

1999 Book

60 Rock Fracture Mechanics; Principles, Design, and Applications

Whittaker B. N., Singh R. N., Sun G.

570

Elsevier Press

1992 Book

61 Rotary Drilling Handbook

Brantly J. E.

825

Palmer Publications

1961 Book

62 Sintered Metal Carbides

Romonova N., Chekulaev P., et al.

332

Mir Publishers

1972 Book

63 SME Mining Engineering Handbook Vol. 1&2

Ed – Cummins A. B., Given I. A.

App 1600

SME of AIME

1973 Book

64 Surface Blast Design

Konya C. J., Walters E. J.

303

Prentice Hall

1990 Book

65 Surface Blaster’s Certification Study Guide

Anonymous

94

Commonwealth of Virginia, Dept. of Mines, Minerals and Energy, Division of Mineral Mining

2009 Book

66 Surface Drilling

Anonymous

114

Atlas Copco

1998 Book

67 Surface Mining

Pfleider E. P.

1061

SME of AIME

1968 Book

(Continued)

Book.indb 720

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Pages

Publisher

Year Type

68 Surface Mining

Kennedy B. A

1175

SME of AIME

1990 Book

69 Surface Mining Equipment

Martin J. W., Bennett T. P., Martin T. J.

455

Martin Consultants Inc.

1982 Book

70 Swedish Blasting Techniques

Anonymous

328

Sprang Tekniska Institutet AB

1974 Book

71 Systematic Drilling and Blasting for Surface Excavations

Anonymous

145

Department of the Army, U.S. Army Corps of Engineers

1972 Book

72 The Drilling of Rocks

McGregor K.

306

CR Books Ltd.

1967 Book

73 The Physics of Rocks

Rzhevsky V., Novik G.

320

Mir Publishers

1971 Book

74 Excavation of Hard Ground Using Drill and Blast Techniques

Graeme Jones

13

AUCTA Tunnel Course

2003 Notes

75 Blaster’s Training Modules

Anonymous

402 Slides

Office of Surface Mining Reclamation and Enforcement, U.S. Department of Interior

2008 Present

76 Blasthole Drilling

Anonymous

27 Slides

Training Module – Office of Surface Mining U.S. Department of Interior, Denver

Present

77 Blasting Rock Practical Exercise

Anonymous

57 Slides

Web Presentation – BlastProd1.pt

Present

78 Drilling of Rock and Earth

Assakkaf I.

72

Department of Civil and Environmental Engineering, University of Maryland

2003 Present

79 Drilling of Rock and Earth

Anonymous

67


2007 Present

80 Excavation Methods

Eberhardt E.

52 Slides

UBC Geological Engineering

2007 Present

81 Explosive and Blasting Agents

Annette Slyman

35

Institute of Mining Engineering, University of Aachen

Present

82 Explosives

Anonymous

42 Slides

Training Module – Office of Surface Mining US Department of Interior, Denver

Present

(Continued)

Book.indb 721

11/22/2010 2:44:16 PM

Table A15.1 (Continued)

Book.indb 722

Name of the book/ Sr. article/product no. literature

Author/editor

Pages Publisher

Year

83 LaserAce Burden Finder

Jagdeepak Sharma

12 Ultra EnviroSlides Systems (P) Ltd.

2005 Present

84 Machine Guidance System – Moving More With Less

Annette Slyman

45 CAT Global Mining Slides

2007 Present

85 Presentation on MREL High Speed Video and Borehole Survey Instruments

Anonymous

20 Web Presentation – Slides Presentation High Speed Video.ppt

Present

86 SimQuarry

Anonymous

27 Sandvik Mining and Slides Construction

Present

87 VOD (Velocity of Detonation) Instruments

Jagdeepak Sharma

22 Ultra EnviroSlides Systems (P) Ltd.

2005 Present

88 Design of Surface Blasts – A Computational Approach

Ashutosh Mishra

77

Dept of Mining Engineering – National Institute of Technology Rourkela

2009 Thesis G

89 Blasting Design Using Fracture Toughness and Image Analysis of the Bench Face and Muckpile

Kwangmin K.

124

Virginia Polytechnic Institute

2006 Thesis M

90 Comparison of Rotary and In-Hole Motor Techniques for Drilling Horizontal Boreholes in Coal

Kravits S. J., Sainato A., Finfinger G. L.

36

1985 Report U.S. Bureau of Mines Report of Investigation No. 8933, U.S. Department of Interior

91 Noise,Vibration and Airblast Management Plan (NVAM) Eagle Rock Quarry Project

Anonymous

7

For Polaris Minerals Corporation by Explosives and Rockwork Technologies Ltd.

92 Rotary Drilling Holes in Coalbeds for Degasification

Cervik J., Fields H. H., Aul G. N.

25


92 Rotary Drilling Techniques Used in the Beckley Coalbed

Goodman T. W.

16


94 Vibration Assessment – Eagle Rock Quarry Project

Anonymous

24


Publication Type

2002 Report

2002 Report

11/22/2010 2:44:16 PM

Table A15.2 List of articles covering drilling and blasting.

Sr. no.

Name of the book/ article/product literature


Pages

Publisher

Year

Type

92

Petroleum Transactions AIME Vol. 216

1959 Article

1

A Laboratory Study of Rock Breakage by Rotary Drilling

Somerton W. H.

2

A New Approach to the Design of Drilling Tools

Mishniavsky L. Jr., 6 Schmauder S.

International Journal Rock Mechanics Mineral Science & Geomechanics

1996 Article

3

A Preliminary Theory of Static Penetration by a Rigid Wedge Into Brittle Material

Paul B., Sikerskie D. L.

372

Transactions SME/AIME Vol. 232

1965 Article

4

A Quiet Series of Innovations Helps “KEEP IT TURNING TO THE RIGHT”

Zaburunov S.

11

World Mining Equipment

5

A Simple Criterion of Machinability of Hard Rocks

Singh S. P.

257–266

International Journal, Mining Geology Engg.

1989 Article

6

A Theoretical Description of Rotary Drilling for Idealized Down-Hole Bit/ Rock Conditions

Paul F., Gnirk J. B., Cheatham J. R.

443

Journal Society Petroleum Engineers

1969 Article

7

A Two Component Model of Blast Fragmentation

Djordjevic N.

13-Sep

Proceedings Australian Institute Mining and Metallurgy

1999 Article

8

Advantages and Disadvantages of Down the Hole Hammer Blasthole Dirlling vs Rotary Drilling in Large Scale Open Pit Surface Mining

Raitt G., Lyon R.

365–376

International Society of Explosives Engineers 1998 G

1998 Article

9

An Empirical Approach for Relating Drilling Parameters

Cunningham R. A. 987

Journal Petroleum Technology

1978 Article

10

An Empirical Equation for Drill Performance Prediction

Rabia H., Brook N.

103–111

21st Symposium on Rock Mechanics

1980 Article

11

An Explosion in Mining

Eloranta J.

12

Eloranta & Associates Inc.

2000 Article

Article

(Continued)

Book.indb 723

11/22/2010 2:44:17 PM

Table A15.2 Continued Name of the book/ Sr. article/product no. literature


Pages

Publisher

Year

11 An Introduction to Drill Vibration

Schivley G. P. Jr.,

647–655

International Society of Explosives Engineers 1997G

1999 Article

12 An Introduction to Drilling Vibrations

Schivley G. P. Jr.,

647–655


1997 Article

13 Assessment of Abrasive Wear Resistance Potential in Rock Excavating Machinery

Atkinson T., Cassapi V. B., Singh R. N.

151–163

International Journal of Rock Mechanics and Mineral Science

1986 Article

14 Asymmetric Blasting: A Rock Mass Dependent Blast Design Method

Segui J. B.

6

Proceedings of EXPLO 2001, Hunter Valley, NSW

2001 Article

15 Bench Blast Modeling Using Numerical Simulation and Mine Planning Software

Firth I. R., Taylor D. L.

4

Mackay School of Mines, University of Nevada, Reno

16 Bench Blast Optimization

Budin B., Newton G., Thompson J.

21

Colorado School of Mines

2003 Article

7

Mining Engineering

2000 Article

17 Bit Geometry Effects on Failure Characteristics of Rock

Type

Article

18 Bit Penetration Into Rock – A Finite Element Study

Wang J. K.

16-Nov

International Journal of Rock Mechanics, Mining Science., & Geomechanics

1976 Article

19 Blasthole Bits – Virtues of Simplicity

Anonymous

3


1994 Article

20 Blasthole Drill: Better by Performance

Carter R. A.

4

Engineering & Mining Journal

1993 Article

21 Blasthole Drilling, Rotary Drilling and the Four Kingdoms

Praillet D.

4


1998 Article

22 Blasthole Drills

Anonymous

Mining Journal

2000 Article

23 Blasthole Drills: Better by Design

Carter R. A.

5

Engineering and Mining Journal

1993 Article

24 Blasting Cap

Anonymous

2

NationMaster.com

Article

(Continued)

Book.indb 724

11/22/2010 2:44:17 PM


Publication Year

Type

Author/editor

Pages

Publisher

25 Blasting in Surface Mines

Anonymous

20

Web Article – H_BLASTING IN SURFACE EXCAVATIONS.pdf

26 Blasting Review & Maintenance Program

Konya Calvin J.

3


1996 Article

27 Blasting Vibrations – Mining Engineer’s Toolbox

Anonymous

2

Web Article

2005 Article

28 Characterization of the Pre and Post Blast Environments

Eloranta J.

10

Proceedings Twenty Ninth Conference of Explosives and Blasting Techniques, ISEE

2003 Article

29 Choice of Optimal Sizes of Rock Destructing Elements and Their Arrangements on the Drilling Bits

Eigheles R. M., 136–150 Strekakive R. M., Mustafina N. N.

Rock Fragmentation (Ed. Eigheles R. M.), VNIIBT Moscow

1975 Article

Article

30 Choosing the Proper Dick R. A., Borehole Size for Olson J. J. Bench Blasting

41–45

Mining Enginnering

1972 Article

31 Concept of Blastability – An Update

Kaushik Dey, Phalguni Sen

8

Indian Mining and Engineering Journal

2003 Article

32 Control of Fragmentation by Blasting

Bozic B.

9

Rudarsko-geološko-naftni zbornik

1998 Article

33 Death of a Dirty Raitt G. Driller: The DM Story

3


1994 Article

Canadian Institute of Mining Bulletin Vol. 75

1982 Article

34 Development Leighton J., Brawner C., of a Correlation Between Rotary Drill Stewart D. Performance and Controlled Blasting Powder Factors 35 Differential GPS Applications at the Morenci Copper Mine

Shchields S., Flinn J., Obregon A.

4

GPS World

2000 Article

36 Down to earth: Mike Woof reports on drilling innovations.

Woof M.

4


2002 Article

37 Drill Accuracy

Harris J. E.

35


1999 Article

(Continued)

Book.indb 725

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Table A15.2 Continued Sr. no.



Pages

Publisher

Year

Type

11

World Mining Equipment Conference

1999 Article

38

Drill Monitoring Developments and the Utilization of the Drill Data

Vynne J.

39

Drill Monitoring Investigations in a Western Canadian Surface Mine

Scoble M., Peck J., Kennedy D.

Proceedings 2nd Conference Large Open Pit Mining

1989 Article

40

Drillability and Physical Properties of Rocks

Singh D. P.

Proceedings Rock Mechanics Symposium Institute of Engineers, Australia

1969 Article

41

Drillability in Hard Rock Drill and Blast Tunneling

Thuro K., Spaum G.

103–109

Falsbau 14

1996 Article

42

Drillability Predictions – Geological Influences in Hard Rock Drill and Blast Tunneling

Thuro K.

426–437

Geological Rundsche 86

1997 Article

43

Drillability Predictions in Hard Rock Tunneling.

Blindheim O.T.

284–289

Tunneling – Institute of Mining and Metallurgy

1979 Article

44

Drillability Studies – Diamond Drilling

Paone J., Bruce W. E.

U.S. Bureau of Mines Report of Investigation No. 6324, U.S. Department of Interior

1963 Article

45

Drillability Studies – Laboratory Percussive Drilling.

Paone J., Madson D., Bruce W.E:


1969 Article

46

Drillability Studies – Percussive Drilling in the Field

Schmidt R. L.


1974 Article

47

Drilling – Mining Engineer’s Toolbox

Anonymous

Web Article

2005 Article

48

Drilling and Blasting in Open Pits and Quarries

Crosby W. A.

Australian Mineral Foundation, Presentation No152/81

1981 Article

49

Drilling and Blasting Pattern Selection for Fragmentation Optimization in Raycon Quarry Ore, Ondo State

Saliu M. A., Akande J. M.

Journal of Engineering and Applied Sciences

2007 Article

50

2

6

(Continued)

Book.indb 726

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Table A15.2 Continued Sr. no.



Pages

Publisher

Year

Type

51

Drilling Machines, Surface

Morrell R., Unger H. F.

Sec. 11.4

SME Mining Engineers Handbook, SME/AIME

1973 Article

52

Dust Control in Surface Mines

Brown J. W.

57


1999 Article

53

Dynamic Stability of an Impact System Connected With Rock Drilling

Fu C. C.

743–749

Journal Applied Mechanics

1969 Article

55

Earth

Arnett W.

5

Web Article

2003 Article

56

Effect of Charge Weight of Vibration Levels from Quarry Blasting

Devine J. F. et al.


1966 Article

57

Energy Balance in Rock Drilling

Simon R.

Journal Society of Petroleum Engr.

1963 Article

58

Engineering Classification and Index Properties for Intact Rocks

Deere D. U., Miller R. P.

Report No. AFWL-TR65–116 Air Force Weapons Lab.

1966 Article

59

Estimating Drilling and Blasting Cost and Analysis and Prediction Model

Borquez G.V.

7


1981 Article

60

Experimental and Theoretical Analysis of Roller Cone Bit Vibrations

Naganawa S., Tanaka S., Sugaya N.

6

Drilling Symposium

1995 Article

61

Explosive Advances

Anonymous

6

World Coal

2002 Article

62

Explosive Material

Anonymous

7

NationMaster.com

63

Explosive Materials

Anonymous

Wikipedia Article

2009 Article

64

Explosives

Anonymous

Web Article – Explosives’04.doc

2004 Article

65

Explosives and Blasting

Sinkala T.

Web Article – Module 6 Explosives and Blasting. pdf

Article

66

Factors in Rotary Drill Evaluation

Frank M. C.

20–33

Mining Congress Journal

1962 Article

67

Flame Jet Gains Ground as a Blasthole Drill

Browning J. A.

80–81

Engineering and Mining Journal,Vol. 170

1969 Article

68

Fluid Drag on a Tool

Rao Bharath

8

www.Ctes.com

1999 Article

69

Focus on Rotary Drilling Rigs

Schivley G. P. Jr.,

Compressed Air Magazine

1987 Article

54

298

Article

(Continued)

Book.indb 727

11/22/2010 2:44:18 PM


Author/editor

Pages

Publisher

Year

70

Geologic Data for Blasting

Eloranta J.

11

Proceedings of the twenty-second conference of Explosives and Blasting Technique, ISEE

1996 Article

71

Going Straight – Straight hole systems boost drilling effectiveness

Woof M.

p. 8


2002 Article

72

GPS Based Machine Guidance in the Mining Industry

Seymour

p. 15

12th International Symposium On Mine Planning And Equipment Selection

2003 Article

73

GPS Guidance System and Reduction of Open Pit Mining Costs and Revenue Loss

Jarosz A. P., Finlayson R.

p. 13

[email protected]

2004 Article

74

GPS Positioning and Equipment Monitoring technology for Blasting Operations

Hendricks C.

14

Proceedings of Minnblast 99 Surface Blasting Conference, Minnesota

1999 Article

75

Ground and Air Vibrations from Blasting

Siskind D. E.

Sec. 11.8

SME Mining Engineering Handbook, SME/AIME

1973 Article

76

Hard Iron Ore Yields Knoerr A. W. to Fusion Piercing

66–69

Engineering and Mining Journal Vol. 147

1946 Article

Proceedings Atlas Copco Bench Drilling Days, Symposium Stockholm Mine and Quarry Engineer

1975 Article

77

Hard Rock Blasting Developments and Possibilities 78 Hard Rock Drilling Techniques 79 Important Considerations When Selecting Rotary Blasthole Drills for High Altitude Use 80 Influence of Joint Direction and Position of Explosive Charge on Fragtmentation

Publication

Fairhurst C.

Fairhurst C., 157 Lacabanne W. D. Lyall Workman J. p. 16

Hafsaoui A., Talhi K.i

8

Calder & Workman, Inc.

The Arabian Journal of Science and Engineering

Type

1957 Article Article

2009 Article

(Continued)

Book.indb 728

11/22/2010 2:44:18 PM



Pages

Publisher

Year

Type

81

Introducing ‘Destruction Work’ as a New Rock Property of Toughness Referring to Drillability in Conventional Drill and Blast Tunneling

Thuro K., Spaum G.

707–713

Eurock 96, Prediction and Performance in R. Mech. and R. Engg. Vol. 2

1996 Article

82

Jet Piercing

Calaman J. J., Rolseth H. C.

325–337

Surface Mining (Ed. Pfleider E. P.), SME/AIME

1968 Article

83

Keep on Running: Rotary Maintenance

Woof M.

p. 6


2003 Article

84

Large Diameter Holes in Overburden Blasting Large Hole Rotary Drill Performance

Piper W. C.

33–35

Mining Congress Journal

1963 Article

Lyall Workman J., 334–347 Calder P.

1996 Article

86

Lubrication of Rock Bits

MacPhail J., Gardner H.

87

Made in Milwaukee

Hammel K.

88

Made in Milwaukee

Woof M.

p. 4

89

Maximing Rotary Blast Hole Drills

Brown J. W.

p. 461

Paul B., Fu C. C.

301–307

International Society of Explosives Engineers 1996G International Society of Explosives Engineers 1996G World Mining Equipment World Mining Equipment International Society of Explosives Engineers 1998G Transactions. SME/AIME Vol. 232

Dobrilovic M., Ester Z., Jankovic B.

7

Rudarsko-geološkonaftni zbornik

2005 Article

Gstalder S., Raynal J.

991

Journal Petroleum Technology

1966 Article

85

90

Maximization of Footage Drilled Before Drill Steel Fails in Fatigue 91 Measurement in Blast Hole Stem and Influence of Stemming Material on Blasting Quality 92 Measurement of Some Mechanical Properties of Rock and Their Relationship to Rock Drillability

140–149

1996 Article

1999 Article 1999 Article 2000 Article

1965 Article

(Continued)

Book.indb 729

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Pages

Publisher

Year

Type

Nauka, Novosibirsk

1985 Article

93 Mechanical and Physical Principles of Design of Rock Breaking Mining Tools

Artsimovich G.V.

94 Mechanical Properties and Drillability of Rocks.

Protodyakonov M. M.

103–118

Proceedings. 5th Symposium on Rock Mechanics

1962 Article

95 Modern Drilling and Blasting Technology at the Minntac Mine

Thompson D.

18

Proceedings of Minnblast 99 Surface Blasting Conference, Minnesota

1999 Article

96 Modern Methods of Rock Fragmentation in Hole Drilling

Suluashin S. S.

Nedra, Moscow

1964 Article

97 Monitoring While Drilling Production Blastholes: Applications in Surface Mining

Peck J., Scoble M. J., Hendricks C.


1988 Article

98 New Blasting Technique to Eliminate Subgrade Drilling, Improve Fragmentation, Reduce Explosive Consumption and Lower Ground Vibrations

Chiapetta F. dd

6

Journal of Explosive Engineering

2004 Article

99 Normalization of Specific Energy

Mellar M.

661–663


1965 Article

100 On Drillability of Rocks by Percussive Drilling

Selmar-Olsen R., Blindheim O. T.

Proceedings 2nd Conference. on Rock Mechanics ISRM

1970 Article

101 On the Penetration of Rock by Three Dimensional Indentors

Miller M. H., Sikerskie D. L.

375–398


1968 Article

102 Open Pit Drilling

Bauer A.

175–180

28th Annual Mining Symp.

1967 Article

103 Open Pit Drilling and Blasting

Bauer A.

115–121

Journal South African Institute of Mining and Metallurgy

1971 Article

(Continued)

Book.indb 730

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Pages

Publisher

Year

Type

104 Optimization of Shape of Drilling Tools

Mishnaevsky L.

1223– 1228

Proceedings. International. Congress Rock Mech.Vol. 2

1999 Article

105 Optimizing Blasthole Spacing

Konya K.

p. 3


106 Optimizing Rotary Drill Performance

Schivley G. P. Jr.,

23-May

International Society of Explosives Engineers 1995 v2 G

107 Optimum Bit Selection and Operation for The Rotary Blasthole Drilling Through Horizontal Drilling Rig (HDR) – A Case Study at KBI Murgul Copper Mine

Ergin H., Kuzu C., Balci C., Tunçdemir H. & Bilgin N.

11

Istanbul Technical University, Istanbul

108 Perfect Cleaning Theory of Rotary Drilling

Maurer W. C.

1270

Journal of Petroleum Technology

1960 Article

109 Performance Characteristics of a Small Scale Rotary Boring Machine Instrumented with Large Drag Picks

Howarth D. F., Adamson W. R.

191–202


1988 Article

110 Physical Properties and Drillability of Mine Rocks

Shepherd R.

468–470, Colliery Engineering Dec. 50, Jan. 51, Feb. 51, 28–34, Mar. 51 51–66, 121–126

1950 Article

111 Predicting Rotary Drill Performance

Schivley G. P. Jr.,

181–189


1994 Article

112 Priming of Explosives

Anonymous

12

Web Article – G_PRIMING OF EXPLOSIVES.pdf

113 Principles of Drilling

Hartman H. L.

Sec 6.1


1968 Article

114 Principles of Drilling

Tandanand S.

Sec.11.3

SME Mining Engg. Handbook, SME/AIME

1973 Article

115 ProVision High Precision GPS Drill System

Anonymous

20

Komatsu Technology Company

2002 Article

Article 1995 Article

Article

Article

(Continued)

Book.indb 731

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Table A15.2 Continued Name of the book/ Sr. article/product no. literature 116 Quantitative Assessment of Rock Texture and Correlation with Drillability and Strength Properties


Pages

Publisher

Year

Howarth D. F., Rowlands J. C.

29

Rock Mechanics and Rock Engineering

1987 Article

4

117 Realising Capacity

Meyhew J.

118 Recording Drilling Parameters in Ground Engineering

Pfister P.

119 Relating Horsepower to Drilling Productivity

Givens R., Williams G., Wingfield B.

120 Relationship Between Thermal Stability and Molecular Structure of Polynitro Arenes

Type

World Coal

2004 Article

Geodrilling

1985 Article


1996 Article

Keshavraz N. H., 4 Pouretedal H. R., Semnani A.

Indian Journal of Engineering and Material Science

2009 Article

121 Review of Rock Drillability and Borability Assessment Methods

Howarth D. F.

12

Transactions International Journal Mining and Metallurgy

1986 Article

122 Rock Abrasiveness Testing for Tunneling

West G.

10


1989 Article

123 Rock Cutting and It’s Potentialities As a New Method of Mining

Cook N. G. W., Toughlins N. C., Weibols G. A.

20

Journal African Institute of Mining and Metallurgy

1968 Article

124 Rock Drillability Related to a Roller Cone Bit

Morris R. I.

Society Petroleum Engrinners Paper 2389

Article

125 Rock Fragmentation and Optimization of Drilling Tools (Chapter 6 in Fracture of Rock Ed. Aliabadi M. H.)

Mishnaesky L. Jr., 36

Computational Mechanics Publications

1998 Article

126 Rock Fragmentation by Concentrated Loading

Simon R.

Failure and Breakage of Rock (Ed. Fairhurst, C.) AIME

1967 Article

20

15

(Continued)

Book.indb 732

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Pages

Publisher

Year

Type

127 Rotary Club: A Turn Around the Latest Large Diesel and Electric Rotary Blasthole Rigs.

Watt J.

5


2002 Article

128 Rotary Drilling

Willaimson T. N. 25


1968 Article

129 Ruggedized Computer Controls Mining Equipment

Mintchell G. A.

2

Control Engineering

2001 Article

130 Selection of Powder Factor in Large Daimeter Blast Holes

Eloranta J.

3


1995 Article

131 SSI Kaltex Drill Adviser

Kerr Cameron

7


1999 Article

132 Stemming Selection for Large Diameter Blastholes

Eloranta J.

4


1994 Article

133 Studies on Drillability of Rocks by Rotary Drills

Kinoshito, S

6

Journal Institute of Mining

1956 Article

134 Swedish Approach to Contour Blasting

Holmberg R, Persson R. A.

Proceedings Fourth Conference of Explosives and Blasting Techniques, ISEE

1978 Article

135 Taming the Dust Devil: An Evaluation of Improved Dust Control for Surface Drills Using Rotoclone Collectors

Page S. J., Organiscak J. A.

5

Engineering & Mining Journal

1995 Article

136 Technical Review of Aquila’s Drill Monitoring Control and GPS Guidance System

Hendrinks C.

14


1999 Article

137 Tests for Rock Drillability

Furby J.

6

Mining and Quarry Engineer

1964 Article

(Continued)

Book.indb 733

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Pages

Publisher

Year

138 The Cercher Abrasivity Index and It’s Relation to Rock Mineralogy and Petrology

Sauna M., Peters T.

7

Rock Mechanics Vol. 15

1982 Article

139 The Concept of Specific Energy in Rock Drilling

Teale R.

17

International Journal of Rock Mechanics and Mining Sci. Vol. 2

1965 Article

140 The Drillability of Rocks

Singh D. P.

6

Mining Science and Engineering

1973 Article

141 The Effect of Apparatus Size and Surface Area of Charge on the Impact Strength of Rock

Rabia H., Brook N.

9


1981 Article

142 The Effect of Some Drilling Variables on the Instantaneous Rate of Penetration

Outman H. D.

13

Transactions SME/AIME Vol. 217

1960 Article

143 The Effects of Blasting on Crushing and Grinding Efficiency and Energy Consumption

Workman L., Eloranta J.

10

Web Article

2005 Article

144 The Hole Picture

Carter Russ

6

Coal Age

2002 Article

145 The Mechanics of Rock Cutting

Nishimatsu Y.

10

International Journal of Rock Mechanics and Mining Science

1972 Article

146 The Perfect Cleaning Theory of Rotary Drilling

Maurer W. C.

10

Journal of Petroleum Technology

1962 Article

147 The Potential for Unifying Drilling, Blasting and Downstream Operations by the Application of Technology

Lyall Workman J. 14

Calder & Workman, Inc.

148 The Prediction and Reduction of Abrasive Wear in Mine Excavating Machinery

Atkinson, T., Cassapi V. B.

Journal Mechanical Engineering

7

Type

Article

1984 Article

(Continued)

Book.indb 734

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Pages

Publisher

Year

Type

Mining Engineering

1978 Article

149 The Pro’s and Con’s of Rotary Blasthole Drill Design

Laswell G. W., Laswell B. J.

150 The Process of Thermal Spalling Behavior on Rocks – An Exploratory Study

Thirumlai K.

23

Rock Mechanics – Theory and Practice SME/AIME (Ed. Somerton)

1970 Article

151 The Stamp Test for Rock Drillability Classification

Wijk G.

8


1989 Article

152 Theory and Practice on Inclined Drilling for Surface Mining

Kochanowsky B. J. 18

Quarterly, Colorado School of Mines

1961 Article

153 Transfer of the Stress Wave Energy in the Drill Steel of Percussive Drill to the Rock

Simon R.

15


1964 Article

154 Twist in the Tale (Rotary Rig Developments)

Woof M.

2


1998 Article

155 Updating Air Practice for Better Open Pit Blasthole Drilling

Clarence Zink

19


1999 Article

156 Wall Control Blasting in Open Pits

Crosby W. A.

Mining Engineering

1982 Article

157 Water Separator Shows Potential for Reducing Respirable Dust Generated on Small Diameter Rotary Blasthole Drills

Listak J. M., Reed W. R.

158 What’s Ahead for Blasting in the Next Millennium

13

Web Article

Article

Many Authors

Drill & Blast

2000 Article

159 What’s New in Drilling and Blasting

Freymuller J.

Rock Products

1999 Article

160 Which Blasthole Rig

Lyall Workman J., Szumansky M.


1997 Article

(Continued)

Book.indb 735

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Book.indb 736


Pages

Publisher

Year

Type

161 Comparison of Rotary and In-Hole Motor Techniques for Drilling Horizontal Boreholes in Coal

Kravits S. J., Sainato A., Finfinger G. L.

36


1985 Report

162 Noise,Vibration and Airblast Management Plan (NVAM) Eagle Rock Quarry Project

Anonymous

7


2002 Report

163 Rotary Drilling Holes in Coalbeds for Degasification

Cervik J., Fields H. H., Aul G. N.

25


1975 Report

164 Rotary Drilling Techniques Used in the Beckley Coalbed

Goodman T. W.

16


1989 Report

165 Vibration Assessment – Eagle Rock Quarry Project

Anonymous

24


2002 Report

166 Optimum Bit Selection and Operation for The Rotary Blasthole Drilling Through Horizontal Drilling Rig (HDR) – A Case Study at KBI Murgul Copper Mine

Ergin H., Kuzu C, Balci C., Tunçdemir H & Bilgin N.

9

Web Article

167 Blast Design – Chapter 19 of Field Manual

Anonymous

90

U.S. Department of Interior, Bureau of Reclamation

168 Explosives – Chapter 2

Anonymous

18

Web Article

Book Ch

169 General Rock Anonymous Removal – Chapter 9

32

Web Article

Book Ch

Article

2001 Book Ch

11/22/2010 2:44:20 PM

Index

Advantages of BECO Threads 120–122 Advantages of Electronic Initiating System 518, 519 Advantages of GPS System 353, 369–371, 591 Advantages of Inclined Blastholes 601, 602, 612–616 Advantages of Mechanical Charging System 571–575 Advantages of Screw Compressors 248, 252–254 Advantages of Sliding Vane Compressors 248–250 Air Blast 535, 545–549, 584 Air Blast Tricone Bit 100 Air Circulation 72, 74, 75, 85, 86, 91–95, 113, 301–303, 648 Air Decking 576, 583–585 Air Pollution 11, 221, 512, 535, 550–552 Air Shock Waves see Air Blast Altitude 23, 110, 223, 326–338, 353, 359, 640, 651–655 Annular Space 16, 127, 182, 296, 300, 301, 305, 307, 318, 319 Bailing Velocity 47, 122, 183, 288, 301–308, 310, 313, 314, 317, 319, 320, 323, 326, 647–649, 689 Bailing Velocity - Effect of Annular Space 302, 305, 307 Bailing Velocity - Effect of Blasthole Inclination 302, 305 Bailing Velocity - Effect of Density of Fragments 302, 304, 307 Bailing Velocity - Effect of Particle Roughness 302, 305 Bailing Velocity - Effect of Quantity of Water Injection 302, 306 Bailing Velocity - Effect of Rate of Fragmentation 302, 305 Bailing Velocity - Effect of Roundness of Fragments 302, 304

Index.indd 737

Bailing Velocity - Formulation 301–307 Bedding Plane Inclination 556, 557 Bedding Planes 41, 556, 558, 559 Bench Height 5, 6, 147, 433, 598, 606, 610, 612, 613, 623–625, 627–634 Bench Nomenclature 6 Bit Breaker 137, 679 Bit Design Feature Summery 103–105 Bit Metallurgy 100–103 Bit Wear Test 60 Bits 14, 20, 21, 46, 68, 69, 71, 72, 75, 83–115, 123, 124 + many more entries Bits - Air Circulation 72, 74, 75, 85, 86, 91–95, 113, 296–301 Bits - Blade Type 22, 83–85, 114 Bits - Claw Type 21, 22, 85 Bits - Drag Type 21, 28, 58, 83–87, 114, 115, 157, 263, 270, 272, 274, 286, 290, 679 Bits - Sealed Bearing 87, 94, 95, 107, 177, 178 Bits - Tooth Spacing 98, 99, 104 Blast Design 347, 370, 473, 597–638 Blast Initiator for 520–525 Blast Initiator for Detonation Wave System 523 Blast Initiator for Electric Detonation 521–523 Blast Initiator for Electronic Detonation 525 Blast Initiator for Hercudet System 523 Blast Initiator for Nonal System 523 Blast Initiator for Safety Fuse 521 Blast Initiator for Shock System 523 Blastability Index 355, 358, 359, 564 Blastability Index - Ashby (1977) 566 Blastability Index - Ghose (1988) 568 Blastability Index - Gupta (1990) 568 Blastability Index - Hainen and Dimock (1976) 566 Blastability Index - Han, Weiya and Shouvi (2000) 569 Blastability Index - Hansen (1968) 565

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738

Index

Blastability Index - JKMRC (1996) 569, 570 Blastability Index - Langefors (1978) 566 Blastability Index - Lilly (1986) 566–568 Blastability of Specimen 350, 555, 563, 564, 633 Blaster’s Multimeter 519 Blaster’s Ohmmeter 519, 520 Blaster’s Tagger 520, 521 Blasthole Camera 139–141, 536, 587, 679 Blasthole Charging Patterns 571, 575–588, 603 Blasthole Deviation 83, 124, 173, 315 Blasthole Dewatering 143, 144, 340, 679 Blasthole Dewatering Pump 143 Blasthole Drill - Advantages of Computerization 365–373 Blasthole Drill - Air Conditioners 181, 186, 188, 189, 219, 340, 387 Blasthole Drill - All Electric 74, 75, 163, 165, 173, 189, 213, 218, 220, 222, 255, 396, 401, 403, 407, 408 Blasthole Drill - Angle Hole Attachments 190, 196–198, 199 Blasthole Drill - Automatic Fire Suppression System 212, 213 Blasthole Drill - Automatic Leveling System 213–215, 350, 355, 359 Blasthole Drill - Automatic Lubrication System 78, 178–181, 338, 399, 402, 409, 414, 429 Blasthole Drill - Auxiliary Winch 154, 155, 175, 176, 181, 190, 204, 205, 387, 429, 430 Blasthole Drill - Batteries 175, 176, 218, 340, 432, 536 Blasthole Drill - Bit Changer 196 Blasthole Drill - Bit Lubricating System 177, 178, 402 Blasthole Drill - Cable Reel 165, 173–176, 189, 220, 387, 426 Blasthole Drill - Carrier Mounted 156, 157, 160, 161, 163, 166, 176, 177, 188 Blasthole Drill - Centralizer 190, 194, 195, 205, 207, 311, 372, 373 Blasthole Drill - Churn 66–68, 74, 153, 597 Blasthole Drill - Computerization 343, 345, 346, 348, 355, 359, 362, 363, 365, 423 Blasthole Drill - Computerization Hardware 362–364, 505 Blasthole Drill - Computerization Software 80, 140, 236, 346, 348, 362, 364, 365, 569, 570 Blasthole Drill - Crawler Mounted 19, 21, 67, 71, 74, 75, 82, 155–157, 159–172, 174, 175, 178, 345, 376, 377, 382, 392, 396, 411, 428, 687–689

Index.indd 738

Blasthole Drill - Depth Indicator 209 Blasthole Drill - Details of 155, 163, 173, 176, 181, 190, 200, 206, 311, 362, 687–689 Blasthole Drill - Diesel Engine 163, 164, 173, 175–177, 215, 217– 222, 228, 255, 330, 333, 341, 397, 401, 403, 408, 432, 687–689 Blasthole Drill - Diesel Hydraulic 156, 157, 175, 219, 222, 397, 400, 402, 420 Blasthole Drill - Double Wall Machinery House 181, 218, 219 Blasthole Drill - Driver’s Cab 154, 164, 173, 177, 188, 189 Blasthole Drill - Dry Dust Collector 156, 162, 164, 184–186, 233, 307, 340, 387, 430 Blasthole Drill - Engine Power 154, 173, 175, 176, 220, 221, 228–231, 330, 335, 376, 386, 387, 687–690 Blasthole Drill - Engine Starting Aid 217, 218 Blasthole Drill - Fast Fuel Fill System 207–209 Blasthole Drill - Fast Retract System 209, 210 Blasthole Drill - Feed Mechanism 21, 117, 124, 141, 190, 200–204, 223, 314, 377, 384, 430 Blasthole Drill - Fire Extinguishers 210, 211, 213, 425 Blasthole Drill - Fuel Tanks 164, 173, 189, 207, 209, 220 Blasthole Drill - Glow Plug 218 Blasthole Drill - Ground Load Bearing 385, 386 Blasthole Drill - Heaters 181, 186, 188, 189, 213, 215, 218, 219, 387 Blasthole Drill - Hydraulic System 75, 154, 155, 173, 177, 222, 335, 391, 428 Blasthole Drill - Hydraulic Test Station 215, 216 Blasthole Drill - Inclinometer 137–139, 679 Blasthole Drill - Language Name Plate 206 Blasthole Drill - Layout 153, 155, 163–165 Blasthole Drill - Leveling Jacks 74, 154, 155, 173–176, 214, 222, 339, 344, 345, 373, 377, 385–387, 430 Blasthole Drill - Lighting 166, 175, 176, 189, 206–208, 214, 387 Blasthole Drill - Machinery House 21, 154, 155, 173, 175, 176, 181, 182, 188, 215, 218, 219, 339, 340, 426 Blasthole Drill - Machinery House Heater 188, 215, 219

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Index Blasthole Drill - Main Frame 138, 154, 155, 157–159, 169, 173, 174, 176, 179, 181, 182, 186, 197, 200, 204, 214, 241, 296, 339 Blasthole Drill - Manifold Flame Heater 218 Blasthole Drill - Mast 21, 64, 66, 78, 80, 117, 137, 138, 153–155, 157–160, 162–164, 172–177, 179, 181, 182, 184, 186–190, 193–198, 200–205, 209, 231, 241, 242, 315, 339, 344–346, 357, 358, 368, 372, 376, 381, 385, 387, 392, 395, 407, 410, 411, 426–430, 602, 687–691 Blasthole Drill - Mast Ladder 190, 191 Blasthole Drill - Mast Raising Cylinders and Mast Braces 164, 189, 191, 196–199 Blasthole Drill - Oil Cooler 154, 164, 165, 173, 176, 181, 253, 254, 335 Blasthole Drill - Operator’s Cab 21, 79, 138, 154–156, 164, 165, 173, 174, 180, 182, 186–188, 191, 215, 216, 219, 355, 358, 363, 364, 368, 424, 430 Blasthole Drill - Pipe Changer 13, 78, 117, 123, 155, 158, 177, 186, 190, 191–195, 311, 313, 348, 384, 385, 392, 410, 411 Blasthole Drill - Primary Requirements 375, 376 Blasthole Drill - Prime Mover 154, 155, 173, 175, 176 also see Engine Power Blasthole Drill - Procurement 368, 419–424 Blasthole Drill - Protective Coating on Steel Components 220 Blasthole Drill - Radiators 173, 181, 340, 416 Blasthole Drill - Remote Propel Control 214, 215, 344 Blasthole Drill - Rotary Head 19, 70, 74, 78, 117, 129, 130, 142, 143, 154–156, 158, 176–178, 181, 182, 186, 190, 193, 194, 198–200, 202, 203, 215, 220, 223, 232, 242, 344, 352, 377, 379, 383, 384, 387 Blasthole Drill - Specifications 408, 421, 422, 687–691 Blasthole Drill - Tool Handling Jib 190, 204, 206 Blasthole Drill - Tow Hook 210 Blasthole Drill - Transformer 165, 173, 175, 176, 189, 215, 220, 221, 333–335, 403, 409 Blasthole Drill - Truck Mounted See Carrier Mounted Blasthole Drill - Types 153–163, 220, 420 Blasthole Drill - Video Camera System 215–217 Blasthole Drill - Welding Outlet 213

Index.indd 739

739

Blasthole Drill - Wet Dust Control 182–184, 186 Blasthole Drilling - Choice of Method 27–29, 420 Blasthole Drilling - DTH 19, 20, 26–28, 77, 129, 248, 254, 311, 420, 631, 687 Blasthole Drilling - History 61–82 Blasthole Drilling - Peculiarities 22, 23 Blasthole Drilling - Rotary 13 15, 20, 21, 23, 26, 28, 29, 46, 64, 68, 71, 72, 74, 75, 77, 78, 80, 82, 83, 86, 117, 119, 121, 124, 125, 128, 129, 132, 134, 136, 144, 145, 147, 153, 155–161, 163, 164, 168, 169, 172, 174, 176–178, 180–182, 184, 187, 189, 190, 196, 198, 204, 207, 209, 210, 213, 215, 217, 221, 223, 229, 230– 232, 241, 247, 248, 251, 264, 270, 280, 286, 295, 308, 317, 322, 326, 330, 335, 336, 339, 341, 343, 345, 346, 348–350, 355, 357–359, 362, 375–379, 381, 383, 385–387, 389, 392–394, 396, 401–404, 419–423, 425 Blasthole Drilling - Rotary Percussion See DTH and Top Hammer Blasthole Drilling - Top Hammer 18–20, 24, 26–28, 75, 77, 260, 631 Blasthole Drilling Patterns 13, 22, 566, 589–591, 603, 608, 633 Blasthole Firing Patterns 520, 588, 589, 592–595 Blasthole Inclination 15, 138–140, 147, 148, 160, 197, 198, 213, 302, 305–307, 318, 319, 352, 359, 377, 380, 381, 384, 385, 420, 422, 433, 456, 553, 556, 559, 565, 597, 598, 603, 609, 610, 612–616, 619, 632 Blasthole Inclinometer 137–139, 679 Blasthole Plugs 134, 583, 584, 588, 620, 679 Blasthole Sampler 145–147 Blasthole Spacing 6, 12, 13, 64, 157, 433, 513, 530, 600, 601, 603, 607, 608, 617, 618, 623–625, 627–630, 632, 633, 635, 637, 638 Blasthole Stemming Height 421, 474, 508, 536, 548, 549, 579, 580, 584, 586–588, 603, 620, 624–629, 632, 633, 637 Blasting - Cast 569, 570, 599, 600, 604, 605, 632–634 Blasting - Conventional Bench 580, 598, 618, 626, 627 Blasting - Hazards 493, 507, 533–553, 595, 597, 615 Blasting - Presplitting 599, 600, 601, 632 Blasting - Rip Rap 583, 603, 629–631

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740

Index

Blasting - Secondary 124, 509, 556, 588, 599 Blasting - Snake Hole 549, 597, 598, 601, 602, 630–632 Breakout Torque 384–385 Burden 6, 12, 147–150, 442–444, 475, 513, 525, 530, 548–550, 563, 565, 568, 598, 603, 607, 608, 612, 614–619, 621–637 Burden Calculation of 621–624 Burden Measuring Instrument 147–150, 525 Cerchar Abrasiveness Test 57, 58 Collaring 215, 340, 344, 345, 369, 370, 431, 534, 588, 614, 623 Comparison of Blasthole and Oilwell Drilling 13 Comparison of Blasthole Drilling Methods 24, 25 Comparison of Crawler and Carrier 161, 162 Comparison of Drill Types 123 Comparison of Slewing and Pivoted Undercarriage 168, 169 Comparison of Water Injection with Dry Dust Collection 181–184 Compressed Air - Discharge 110, 229–231, 236, 238, 247–252, 254–257, 301, 308, 314, 317–320, 322, 323, 325, 326, 336, 337, 386, 655, 657–666 Compressed Air - Flow in Hose Pipes 240, 241, 245, 309, 657–659 Compressed Air - Flow in Pipe Fittings 231, 241, 243, 244, 651, 661–666 Compressed Air - Flow in Steel Pipes 233–240, 651–655 Compressed Air - Flow through Nozzles 91, 93, 182, 232, 233, 245–247, 296, 300, 309, 320–326, 336, 338 Compressed Air - Leakage 247, 248 Compressed Air - Measurements 256, 257 Compressed Air - Pressure Loss 117, 118, 177, 235, 236, 238–247, 251, 256, 257, 309, 310, 322, 323, 336, 338, 352, 373, 652, 655, 657–666 Compressor 47, 154–156, 159, 164, 165, 173, 175–178, 181, 182, 186, 188, 223, 225, 227–231, 233, 236, 238, 246–257, 295, 296, 301, 308–310, 314, 317–320, 322, 323, 326, 330, 335–337, 386, 387, 401, 403, 407, 409, 410, 414, 415, 423, 428, 429, 432, 433, 480, 583, 651, 688, 690 Compressors Rotary Screw 176, 229, 230, 248, 250–254, 337 Compressors Sliding Vane 248–250, 254 Concepts of Drilling 16–22

Index.indd 740

Delay Element 512, 513, 515–517, 588 Detonating Cord 456, 472, 494, 508–513, 516, 517, 530, 536, 548, 583 Detonator 15, 62, 449, 451, 454, 472–475, 496, 498, 502, 507–510, 512–523, 525, 530, 535–537, 575–577, 580, 581, 588, 626 Detonator - Electric 62, 498, 513–516, 520, 521 Detonator - Electronic 518–520 Detonator - Non Electric 473, 516, 517 Dozer 9, 11, 210, 348, 350, 370–372, 426 Dragline 11, 74, 153, 222, 317, 348, 350, 355, 371, 393, 420, 422, 570, 586, 599, 600, 604, 605, 609–612, 616 Drill Collars 71, 315 Drill Operation 343–345, 349, 355, 372, 400, 408, 429 Drill Operation - Pipe Handling 78, 344, 345, 386, 387 Drill Operation - Leveling 111, 154, 174, 213, 214, 343, 344, 346, 354, 369, 386, 387, 423, 428, 429 Drill Operation - Tramming 343, 344, 346, 350, 372, 423, 426, 427, 434, 689 Drill Pipe 13, 16–19, 21, 24, 72, 74, 78, 82, 108, 110, 111, 113, 117–124, 128, 130, 134, 136, 137, 141, 143–146, 153–155, 157, 162, 185, 186, 191–196, 201–205, 207, 209, 232, 236, 256, 296, 300, 301, 309–320, 322, 323, 345, 350, 352, 359, 362, 368, 372, 379, 383, 392, 405, 410, 411, 416, 426, 429, 430, 433, 648, 649, 679, 687, 689, 691, 693–697 Drill Pipe - Choice of 122, 123, 310–319 Drill Pipe - Discarding 317, 320 Drill Pipe - Fabricated 13, 117, 118, 315, 679, 694, 696 Drill Pipe - Integral 13, 117, 118, 125–127, 315, 326, 679, 694, 696, 702, 703 Drill Pipe - Sizes 693–697 Drill Pipe - Surface Treatment 122, 310–313 Drill Stem Wrench 136, 137, 186, 345, 679 Drilling Angle Hole 25, 128, 184, 194 Drilling Cost 307, 389, 392, 393, 404, 406–408 Drilling Cost - Operating 199, 398–404, 406, 407, 413, 418 Drilling Cost - Owning 369, 393–399, 406, 412, 413 Drilling Cost - Salvage Value 112, 393, 396, 397, 413 Drilling Diamond 71, 73, 260, 263, 284–286, 290, 348, 358 Drilling Oilwell 12–14, 23, 64, 71, 86, 119, 132, 306, 315, 348, 349

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Index Drilling Methods 12, 17, 21, 23, 24, 26, 29, 259, 260, 393 Drilling Recorders 78, 108, 110, 111, 349–351 Drilling Records 432–434 Effect of Extremely Cold Weather 25, 215, 217–221, 250, 339–341, 407 Effect of Extremely Hot Weather 338–340 Effect of Heavy Rainfall 327, 338, 340, 377 Effect of Heavy Snowfall 327, 338–340 Effect of Scarce Air on Air Compressors 335, 336 Effect of Scarce Air on Blasthole Flushing 336–338 Effect of Scarce Air on Diesel Engines 330–333 Effect of Scarce Air on Drill Lubrication 338 Effect of Scarce Air on Electric Motors 333–335 Effect of Scarce Air on Hydraulic System 335 Effect of Stormy Wind 338, 339 Explosive - Ammonium Nitrate 448, 460, 466, 472, 473, 475, 477, 478, 480, 482, 485, 487, 488, 490, 498, 572 Explosive - ANFO 447–449, 451, 452, 460, 461, 473, 475, 477–486, 488, 490, 491, 525, 534, 565, 566, 568, 571–573, 575–579, 584, 585, 587, 603, 604, 608, 609, 617, 626, 628, 629, 632 Explosive - Blasthole Pressure 450, 452, 555 Explosive - Boosters 474, 475, 485, 507–509, 515, 559, 575, 576, 582 Explosive - Brisance Value 450, 455 Explosive - Charging Density 450, 455, 456, 555, 626 Explosive - Detonation Pressure 440, 450–452, 476, 477, 484, 487, 489, 491, 576, 577, 580, 622 Explosive - Dynamite 63, 447, 449, 450, 472, 473, 475, 477, 479, 485, 486, 488–491, 498, 499, 501, 507, 617 Explosive - Emulsion 452, 473, 475, 477, 479, 485–487, 502, 507, 534, 572–575, 577–580, 584, 587, 628, 629 Explosive - Fuel Oil 448, 473, 475, 478–481, 487, 572 Explosive - Heat of Explosion 449, 450, 453, 459, 464–466, 468, 469, 484, 487, 489 Explosive - Heavy ANFO 452, 482, 484, 485, 572 Explosive - Nitrocellulose 448, 450, 466, 475, 477, 488, 489, 501, 515 Explosive - Nitroglycerine 63, 488, 489

Index.indd 741

741

Explosive - Oxygen Balance 465–467, 484, 487, 489 Explosive - PETN 446–448, 450, 452, 455, 468, 475, 476, 507, 510, 512, 576, 578 Explosive - Primary 446–448, 473, 475, 476, 512, 514–516, 518 Explosive - Properties of 445, 449–461, 471, 493, 503, 628 Explosive - RDX 447, 450, 455, 468–470, 475, 476 Explosive - Slurry 448, 451, 461, 473, 475, 477, 479, 485, 486, 502, 534, 573–575, 577, 579, 617 Explosive - Sodium Nitrate 448, 475, 478, 485, 488–490 Explosive - Specific Gas Volume 450, 453 Explosive - Strength of explosive 452, 453, 465, 469, 470, 496–498 Explosive - Sulfur 448, 457, 472, 475, 479, 488–490, 573 Explosive - Thermochemistry 453, 463–470 Explosive - TNT 446–448, 450, 452, 455, 468–470, 472, 475, 476, 507, 576, 578 Explosive - Volume of Products of Explosion 465, 467, 468 Explosive - Water Resistance 450, 458, 459, 476, 484, 485, 487, 489, 491, 507, 535 Explosive Strength 452, 453, 465, 469, 496–498, 500 Explosive Strength - Absolute and Relative 452, 453 Explosive Strength - Weight and Bulk 452 Feed Force 17, 18, 21, 24, 25, 28, 71, 77, 80, 85, 89, 95, 105, 112, 117, 118, 123, 124, 128, 130, 141, 142, 154–156, 157, 159, 162, 174, 198, 200, 209, 261, 273–280, 282–284, 287, 289, 311, 314, 315, 317, 318, 344, 345, 348, 349, 352, 357, 358, 360, 361, 376–378, 382–385, 405, 615, 672, 679 Feed Force Measuring Kit 679 Feed Force on Drag Bits 21, 85, 86, 274 Feed Force on Tricone Bits 105 Feed Mechanism - Chainless Rack and Pinion 200, 203, 204 Feed Mechanism - Double Drum and Wire Rope 200 Feed Mechanism - Fixed Sprockets and Chain 200–202 Feed Mechanism - Hydraulic Cylinder and Chain 200, 201 Feed Mechanism - Hydraulic Cylinder and Wire Rope 200, 201 Feed Mechanism - Rack and Pinion with Chain 200, 202, 203

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742

Index

Firing Patterns - Channel Delay Pattern 592–595 Firing Patterns - Echelon Delay Pattern 592, 593 Firing Patterns - Flat Face Pattern 592, 593 Firing Patterns - Sinking Hole Pattern 592, 595 Firing Patterns - V Pattern 592, 593 Flushing 13, 47, 59, 71, 72, 84, 107, 113, 123, 124, 145–147, 154, 181, 183, 186, 223, 231, 251, 254, 275, 279, 288, 295–314, 322, 323, 330, 336–338, 348, 380, 402, 419, 429, 615 Formation Fracture Below a Drill Bit 274–280 Formation Logging Units 348, 349 Gas Laws 225 Gas Laws - Amagat’s Law 225, 227 Gas Laws - Avogadro’s Law 225, 227 Gas Laws - Boyle’s Law 225, 227, 229 Gas Laws - Charles’ Law 225, 227, 468 Gas Laws - General Gas Law 227 Gas Laws - Guy Lussac’s Law 225, 226 Gas Laws - Joule’s Law 225, 226 Gas Laws - Poisson’s Law 225, 226, 229 Geological Cycle 34, 35 GPS 80, 110, 346, 347, 350–356, 362, 365, 369–372, 591, 679, 681, 683 Ground Vibrations 452, 527, 534, 537–545, 550, 595, 603, 614 Ground Vibrations - Damage 537, 541–545 Ground Vibrations - Prediction 539–541 Hardness 9, 17, 27, 46, 50–58, 60, 75, 90, 103, 105, 106, 113, 145, 160, 203, 204, 273, 282, 285, 357, 379, 380, 567, 605, 616, 667, 668, 670–673, 675, 676, 710, 711 Hardness Scales 675–678 Hercudet System 517, 521, 523–525 High Speed Camera 525, 529, 549 Hot Hole Meter 525, 526, 534 Humidity 327, 329–333, 337, 340, 480, 482, 503 Influence of Rock Density on Blasting 554, 555 Influence of Rock Porosity on Blasting 555 Influence of Rock Strength on Blasting 554 Kelly 64, 71, 72, 157–159, 429, 489 Laser Measuring Instruments 679 Lifting Bails and Hoisting Plugs 134, 135

Index.indd 742

Loading ANFO and Heavy ANFO by Truck 572, 573 Loading Slurry and Emulsion ANFO by Truck 573, 574 Machine Digging 7, 8, 11 Maintenance 78, 130, 162, 163, 169, 172, 186, 199, 201–204, 219, 221, 241, 254, 369, 390, 399, 400, 408, 414, 421, 423, 424, 428, 429, 431, 432, 434, 536, 571 Measurement While Drilling 349, 355, 359 Metamorphism 34, 37–39, 45 Mineral Occurrence 44–46 Minerals 1, 2, 3, 10, 12, 32, 34, 35, 37–39, 43–46, 52, 54, 55, 61, 62, 263, 264, 311, 389, 440, 460, 551, 561, 586, 587, 614, 675, 713 Miniature Drill Test 57–60 Misfires 534, 535–537, 603, 614 Movement of Drill by Using GPS 353, 354 Operating a Rotary Blasthole Drill 428–431 Penetration of a Wedge 272 Penetration of an Indenter 292 Penetration Rate Estimation of 282–294 Periodical Check Ups 432 Power for Air Compression 228–231 Priming 458, 580–582, 586–588 Priming - Bottom 580–582, 586 Priming - Multi Point 580, 582 Priming - Special Rock Mass Conditions 586–588 Priming - Top 580–582, 586 Pulldown 71, 111, 129, 133, 165, 176, 181, 188, 201, 319, 433, also see Feed Force Pullout Force and Speed 384 Recovery Tools 144, 145, 679 Ripping 5, 7–9, 11, 12, 74, 137, 555 Rock - Abrasiveness 46, 57–60 Rock - Brittleness 46, 54, 57, 341 Rock - Coefficient of Internal Friction 46, 55, 56, 265, 710 Rock - Compressive Strength 2, 4, 8, 46–49, 51, 54, 57, 105, 111, 273, 273, 285–288, 294, 421, 554, 563, 569, 570, 582, 605, 628, 672, 673, 710, 712, 715 Rock - Density 31, 32, 36, 42, 45–47, 282, 293, 297–302, 304, 307, 308, 318, 539, 553–555, 563, 566, 569, 571, 577, 604, 616, 617, 710, 712–714 Rock - Drillability 46, 281, 282, 292, 379 Rock - Failure Theories 265–270

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Index Rock - Igneous 10, 34–36, 39, 45, 304, 313, 314, 318, 710, 712 Rock - Metamorphic 10, 34, 35, 37–39, 304, 313, 314, 318, 710, 712 Rock - Sedimentary 7, 10, 11, 27, 34–38, 45, 48, 304, 313, 314, 318, 444, 556, 557, 581, 586, 710, 712 Rock - Swell Factor 57 Rock - Tensile Strength 46, 53, 50–52, 262, 282, 442, 510, 511, 554, 622, 675, 710–712 Rock - Toughness 17, 46, 54, 101, 273, 282, 421, 672, 673, 711 Rock Breaking 7, 11, 12 Rock Fracture by Crushing 11, 39, 95, 104, 112, 181, 258, 261–264, 279, 281, 440, 469, 583, 599, 600, 606, 614 Rock Fracture by Impacts 262 Rock Fracture by Scratching 262–264 Rock Fracture by Spalling 263, 440, 442, 561 Rock Fracture by Vaporization 263 Rock Mass - Classification 560–569 Rock Mass - Dip 557, 558, 563, 603, 604 Rock Mass - Folds 42, 43, 556, 557, 608 Rock Mass - Formation of 18, 37, 45, 273, 437, 440, 560, 608 Rock Mass - Joint 282, 554, 561 Rock Mass - Joint Spacing 282, 554, 561 Rock Mass - RMR 562–564 Rock Mass - RQD 562, 563 Rock Mass - RTQI 562, 564 Rock Mass - Strike 557, 559, 563, 603, 604 Rock Mass - Unconformities in 42, 43, 556 Rock Mass - Voids in 123, 146, 449, 533, 556 Rotary Deck Bushing 136, 186, 679 Rotary Speed of Drag Bits 85, 86 Rotary Speed of Tricone Bits 105, 110–112, 277, 278, 281, 282, 287, 289, 290, 292, 294, 320, 344–346, 348–352, 357, 359, 360, 366–368, 378–380, 382, 406, 428, 429, 433, 570, 688, 690 Safety Fuse 62, 472, 474, 509, 520, 521, 535 Safety Items - Close Fitting Shirt and Long Pants 150 Safety Items - Ear Plugs or Ear Muffs 150 Safety Items - Face Shield 151 Safety Items - Rain Coat 151 Safety Items - Respirator 151 Safety Items - Safety Glasses 150, 151 Safety Items - Safety Toed Shoes 150 Safety Items - Safety Vest 150

Index.indd 743

743

Safety Precautions 434 Seismograph 472, 473, 525, 527, 538 Sensitivity of Explosive 450, 453–456, 459, 467, 476, 477, 480, 481, 484–487, 489, 500–503, 511 Sensitivity of Explosive to Detonator Strength 454 Sensitivity of Explosive to Friction 454 Sensitivity of Explosive to Gap 455 Sensitivity of Explosive to Heat 454 Sensitivity of Explosive to Shock 454 Setting Up the Rotary Blasthole Drill 428 Shock Absorbers 128–134, 199, 405, 707, 708 Shock Absorbers - External 130–133, 707, 708 Shock Absorbers - In the Hole 132, 134 Shovel 7, 11, 27, 74, 78, 153, 170, 222, 317, 348, 350, 355, 370–372, 393, 406, 420, 422, 570, 586, 592, 593, 598, 599, 604, 605, 609, 610, 612, 616 Size and Weight of the Drill 376–378 Soil Profile 43, 44 Specific Energy 280, 281, 289, 290 Stabilizer 21, 85, 91, 110, 113, 117, 123–128, 137, 183, 309, 316, 317, 380, 384, 405, 417, 426, 679, 699–706 Stabilizer - Integral Blade 125–127, 679, 702, 703 Stabilizer - Replaceable Blade 125, 126, 704, 705 Stabilizer - Roller 125, 127, 128, 679, 699, 700 Stabilizer - Welded Blade 125, 126, 679, 700–702 Structures - Extrusive 36, 40–42, 45 Structures - Hypabyssal 36, 40, 41 Structures - Plutonic 36, 40–42, 555–557, 560 Sub - Crossover 117, 128, 137, 241, 405, 407, 679 Surface Activities 43 Tests on Explosives 493–505 Tests on Explosives - Ballistic Mortar Test 497 Tests on Explosives - Bullet Impact Test 505 Tests on Explosives - Cap Sensitivity Test 502 Tests on Explosives - Chronograph Test for VOD 494, 495 Tests on Explosives - Cratering Test 497 Tests on Explosives - Critical Diameter Test 504

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744

Index

Tests on Explosives - Critical Height Test 505 Tests on Explosives - Cylinder Compression Test 499 Tests on Explosives - D’Autriche Test for VOD 494 Tests on Explosives - Effect of Wetness Test 503 Tests on Explosives - Electrostatic Discharge Sensitivity Test 503 Tests on Explosives - Fiber Optic Sensor Test for VOD 495 Tests on Explosives - Friction Sensitivity Test 502 Tests on Explosives - Gap Sensitivity Test 502 Tests on Explosives - Heat Resistance Test 500, 501 Tests on Explosives - Internal Ignition Test 503 Tests on Explosives - Koenen Test 505 Tests on Explosives - Material Compatibility Test 504 Tests on Explosives - Optical Measurement of VOD 495 Tests on Explosives - Plate Dent Test 497 Tests on Explosives - Shock Sensitivity Test 500 Tests on Explosives - SLIFER Test for VOD 495 Tests on Explosives - Traulz Lead Block Test 498 Tests on Explosives - Underwater Test 497 Tests on Explosives - Vacuum Stability Test 504 Thread Protectors 134, 135 Tool Joints 118–122 Tool Joints - API Regular 91, 100, 119, 120, 122, 128, 706 Tool Joints - BECO 100, 119–123, 128, 385 Toxic Fumes 450, 456–458, 551

Index.indd 744

Transporting the Drill 431 Tricone Bits 14, 21, 68, 72, 75, 83, 86–115, 128, 177, 196, 262, 270, 276, 277, 279, 280, 286, 290, 294, 311, 320, 322, 379, 383, 402, 405, 417, 420, 669 Tricone Bits - Carbide Insert 20, 84, 87, 88, 95, 101, 102, 105–109, 115, 276, 292, 293, 669–673, 679 Tricone Bits - Cones 18, 21, 69, 71, 86–91, 95, 98–102, 108, 111–114, 275, 294, 383, 429 Tricone Bits - IADC Code 105–107, 379 Tricone Bits - Milled Tooth 87, 88, 95, 101, 112 Tricone Bits - Nozzles 92, 93, 107, 320–323, 325 Tricone Bits - Performance 88, 91, 95, 103, 109–112, 282, 286, 379, 404 Tricone Bits - Reasons for Failure 112–114 Tricone Bits - Rejection of 108, 109, 111, 115 Tricone Bits - Shirttail 92, 95, 97, 109, 113, 124 Tricone Bits - Steel Tooth 87, 105–107, 115, 276, 667, 679 Truck for Blasthole Charging 571–575 Tungsten Carbide 10, 20, 21, 46, 59, 60, 68, 75, 83, 84, 87, 88, 94, 99, 101, 102, 112, 115, 125–127, 286, 292–294, 667–673 Undercarriage 154, 156, 162–170, 174, 181, 376, 395 Undercarriage - Pivoted Crawler 169 Undercarriage - Slewing Crawler 166–169 VOD meter 525, 528 Volcanic Activities 35, 40 Water Jetting 7, 8 Why Compressed Air for Flushing 295

11/22/2010 3:57:28 PM

rotary drilling and blasting in large surface mines .pdf

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