Thesis Shokrollah Zare - Prediction Model and Simulation Tool for Time and Cost of drill and Blast Tunnelling.pdf

CICJ CdglZ\^VcJc^kZgh^ind[ HX^ZcXZVcYIZX]cdad\n I]Zh^h[dgi]ZYZ\gZZd[ e]^adhde]^VZYdXidg ;VXjaind[:c\^cZZg^c\HX^ZcXZVcYIZX]cdad\n 9ZeVgibZcid[8^k^aVcYIgVchedgi:c\^cZZg^c\

9dXidgVaI]Zh^h 9dXidgVaI]ZhZhViCICJ!'%%,/&'.&

H]d`gdaaV]OVgZ

EgZY^Xi^dcBdYZaVcYH^bjaVi^dcIdda [dgI^bZVcY8dhid[9g^aaVcY7aVhi IjccZaa^c\

>ccdkVi^dcVcY8gZVi^k^in

Shokrollah Zare

Prediction Model and Simulation Tool for Time and Cost of Drill and Blast Tunnelling

Doctoral thesis for the degree of philosophiae doctor Trondheim, June 2007

Norwegian University of Science and Technology Faculty of Engineering Science and Technology Department of Civil and Transport Engineering

NTNU Norwegian University of Science and Technology Thesis for the degree of philosophiae doctor Faculty of Engineering Science and Technology Department of Civil and Transport Engineering © Shokrollah Zare ISBN 978-82-471-2825-1 (printed version) ISBN 978-82-471-2839-8 (electronic version) ISSN 1503-8181 Doctoral Theses at NTNU, 2007:129 (1) Printed by Skipnes AS

Abstract The drill and blast method is applied in a wide range of underground spaces, such as tunnels, rock caverns and mines. Various applications of the method as well as the growing demand for underground space necessitate updated design and planning models for blast design, time scheduling and cost estimation. A search of the literature reveals no evidence of published comprehensive prediction models for excavation time and costs, apart from the NTNU model. The thesis aims to provide a practical tool for the tunnelling industry to be used through all phases of a project. The only widely used drill and blast prediction model has been updated and further developed on the basis of the most recent technology, equipment and new field data. The thesis consists of four volumes and one software program, that covers updated blast design, advance rate and cost models. In addition, an Excel-based simulation tool (TunSim), a new engineering tool; facilitates the use of the prediction models as follows: • • • •

Fast and exact calculation of time and cost, as an alternative to the paper versions Possibility to vary all the input data Parametric studies and risk analyses on various input data related to excavation time or costs Built-in calculation for blast design and ventilation design

Comparison of the blast design model with a Swedish model shows that the NTNU model for blast design gives higher number of holes and lower of explosives consumption than the Swedish model. By applying the advance rate and cost model for medium rock conditions, 5 m round length and 3 km tunnel length, the weekly advance rate (100 working hours) varies from 100 (10 m2) to 54 (120 m2) m/week, and the unit excavation costs varies from 6500 (10 m2) to 17500 (120 m2) NOK/m, depending on the equipment combination and number of drilling hammers. The advance rate and costs exclude rock support, and the costs are based on June 2005 price level. The impact of rock conditions, i.e. rock blastability and drillability is investigated. In the range from good to poor blastability, the excavation time will increase by 22 % and the excavation costs will increase by 15 %. The corresponding increase for drillability is 10 % for the time and 13 % for the costs.

iii

In line with developments in tunnelling, the productivity and efficiency of the drill and blast method have increased. The investigation shows a substantial increase in the productivity and reduction in the costs. For a 60 m2 road tunnel, the advance rate has increased from 50 to 81 m/week during the past 30 years, indicating a 60 % increase in the production rate. And the excavation cost has decreased from some 16000 to 10200 NOK/m, indicating a 36 % reduction from 1975 to 2005. In comparison with TBM, the drill and blast method is efficient and very cost effective in hard rocks and larger cross sections in various rocks. The method has better flexibility to deal with the geological risks and the amount of investment is substantially lower.

iv

Preface The thesis is the result of my work at the Department of Civil and Transport Engineering, Norwegian University of Science and Technology in Trondheim, that was supported by a scholarship from the Ministry of Science, Research and Technology of Iran. The thesis includes a combination of different tasks; field data acquisition and analysis, collection of various equipment and materials data, updating, revising and further developing the models for blast design, advance rate and costs, and in parallel, developing a new engineering tool to facilitate the use of the models in computers. Many of people and companies have contributed in different ways with information or assistance to this work. First of all, I acknowledge my supervisor Professor Amund Bruland for his supervision during the PhD study. I would also like to thank, Assistant Professor Vegard Olsen for his general comments and Eivind Hegbom, Researcher for his help with the simulation tool. I acknowledge the English editing assistance from Stewart Clark, NTNU. My friends and colleagues at the department are also appreciated, especially PhD candidates Alex Klein Paste and Yangkyun Kim. I am grateful to the suppliers that have provided data and information for equipment and materials; Thorbjørn Ramsvik and Roar Woldseth (Atlas Copco), Thomas Edvardsen (Sandvik), Reidar Frisell (AMV), Egil Amundsen (Boart Longyear), Jan Kristiansen (Dyno Nobel), Trond Solberg (Orica), Svein Haaland (Korfmann), Helge Opheim (Pon) and Per Ek (Gia). Finally, my warmest thanks to my wife and children for their support and understanding.

v

Table of contents

Abstract Preface 1 Introduction

iii

v 1

1.1 General

1

1.2 Objectives of the thesis

2

1.3 Background of the models

3

2 Drill and blast tunnelling

5

2.1 Introduction

5

2.2 Development of the method

6

2.3 Present state of the art

7

2.3.1 Drilling 2.3.2 Charging and explosives 2.3.3 Ventilation 2.3.4 Loading and hauling 2.3.5 Rock support 2.4 Future developments 3 The NTNU model 3.1 Blast design model 3.1.1 Rock blastability 3.1.2 Cut design 3.1.3 Drilling pattern 3.1.4 Charging 3.1.5 Firing pattern 3.2 Advance rate model 3.2.1 Drilling and charging 3.2.2 Ventilation 3.2.3 Loading and hauling vi

7 9 10 11 11 12 15 15 16 17 17 18 18 19 19 20 20

3.2.4 Scaling and rock support 3.2.5 Weekly advance rate 3.3 Cost model 3.3.1 Cost calculations 3.3.2 Cost model summary 3.4 Field data 3.4.1 Blast design 3.4.2 Advance rate 3.5 Model developments 3.5.1 Overview from 1975 to 1995 3.5.2 Latest developments, 2005 3.5.3 Time and cost trends

21 21 24 25 28 31 31 33 37 37 38 40

4 Simulation tool

45

4.1 Introduction

45

4.2 Advance rate model

48

4.3 Cost model

51

4.4 Sensitivity analysis

55

5 Conclusions and recommendations for further work

59

5.1 Conclusions

59

5.2 Recommendations for further work

61

References

63

Published papers

67

Comparison of tunnel blast design models Estimation model for advance rate in drill and blast tunnelling Progress of D&B efficiency with relation to excavation time and costs Assessment of TBM and D&B based on excavation time and costs Appendices Appendix A Field data Appendix B Equipment price and lifetime Appendix C Cost model in simulation tool

69 79 85 91

99 101 105 115

vii

viii

1 Introduction

1.1 General The thesis consists of four volumes and one software program, i.e. the present chapters which give the background and summary of the thesis and the three reports and software described in the following: 2A-05

Drill and Blast Tunnelling - Blast Design

2B-05

Drill and Blast Tunnelling - Advance Rate

2C-05

Drill and Blast Tunnelling - Costs

TunSim

Simulation Tool

The present volume Prediction Model and Simulation Tool for Time and Cost of Drill and Blast Tunnelling gives the background and discussion of the above reports and software and covers all the topics in the thesis. Report 2A-05 Drill and Blast Tunnelling - Blast Design describes rock blastability and cross section design for various tunnels. The blast design method and data based on parallel hole cut for 48 mm and 64 mm drillhole diameter are presented. The necessary explosives consumption is given for cartridged and bulk explosives. Report 2B-05 Drill and Blast Tunnelling - Advance Rate covers the estimation model for advance rate. The time consumption for major operations of drilling, charging, blasting, ventilation, loading, hauling, scaling and rock support is presented for all tunnel cross sections and excavation methods. Finally, the weekly advance rate for the best combinations of equipment for 48 mm drillhole diameter and 3 km tunnel is given. 1

1 Introduction Report 2C-05 Drill and Blast Tunnelling - Costs presents the detailed excavation costs. The costs of the tunnel operations are given in two detailed and summary chapters. An example of total construction costs including rock support costs is presented. TunSim is a simulation tool developed for fast time and cost prediction and as an alternative to the paper version. The software can be used for all tunnel cross sections and excavation methods. The main inputs are tunnel and drillholes geometry, rock properties and equipment type. Users have the possibility to vary the input by their own experienced data. The software may also be used for risk analysis in time or cost. Further details are given in Chapter 4.

1.2 Objectives of the thesis The drill and blast method is applied in a wide range of underground spaces, i.e. tunnels, rock caverns and mines. Various present and future applications of the method in underground excavation as well as the growing demand for underground space and facilities necessitate updated design and planning models for blast design, time scheduling and cost estimation. A study of the literature reveals no evidence of published comprehensive prediction models for the excavation time and costs other than the NTNU model. Despite a lot of investment in the construction of underground spaces, the research on the subject is obviously not enough to meet all the requirements. The thesis aims to address this issue and provide updated practical tools for the tunnelling industry to be used through all phases of a project: • • • • • •

Preliminary and feasibility studies Site investigations Projects design and optimisation Tendering and contract Construction Possible disputes or claims

The objectives are summarised to the following sub-tasks:

2

•

Update and develop the blast design, advance rate and cost models of NTNU

•

Develop a simulation tool (PC software) for fast and exact prediction of the excavation time and costs. This will also enhance the use of the models through the possibility of varying all the input data

1 Introduction •

Study of the effect of the rock mass properties such as drillability and blastability on the excavation time and costs

•

Assessing risk with regard to deviation or variation in the estimated rock parameters or performance of the tunnelling equipment

1.3 Background of the models The models for blast design and advance rate are empirical, i.e. they are developed based on field studies and statistics. The field data are acquired from various tunnels in Norway such as road and railway tunnels, hydropower and rock cavern projects. The data are analysed and then normalised to be published in the paper version. In that manner, the model corresponds to the demand for prediction and estimation with an acceptable level of accuracy which is quite important since there are no other exact solutions like analytical or numerical models when dealing with the excavation of a tunnel. Rock conditions, equipment performance and the tunnel crew skill level are the main parameters that influence tunnel operations; the combination of these parameters is also an important issue which is considered in the models. This is achieved through recording the proper data in the field studies. An example is the penetration rate of different hammers as a function of Drilling Rate Index, DRI. The updating of the new models is performed based on the following procedures: • • •

Collecting of the new field data from the ongoing or finished projects and analysing the data (the data are shown in Appendix A) Check against the existing model, if the new data do not fit to the existing data/curves, the curves are adjusted to get a better fit to the expanded data In case of technology development or new equipment, the new equipment performance or capacity is introduced to the model

The cost model is not directly based on field cost studies, the input data are mainly equipment and material prices and expected lifetime of equipment. These data are acquired through follow up from different suppliers and contractors. Then, based on the analytical calculations, the cost of each operation is calculated. The method of calculation of the cost items is treated in Chapter 3. Also the updated blast design and advance rate models are used to update the cost model.

3

1 Introduction The results of the models are deterministic. It means for a set of specified input for a tunnel; the models give only one value as a result, e.g. for number of holes, explosive consumption, advance per week or costs. In case of an uncertain input or risk evaluation, the estimation may be repeated for a range of input to get corresponding results for all input values. In most cases input vs. result variation is very useful.

4

2 Drill and blast tunnelling

2.1 Introduction The drill and blast cycle is shown in Figure 2.1. The cycle includes drilling, charging, blasting, ventilation, loading, hauling, scaling and rock support. The method requires heavy and efficient equipment and well organised tunnelling. The NTNU model covers all the operations in the cycle; but the standard advance rate and costs presented in the thesis exclude rock support. The method can be applied to a wide range of rock conditions. The base of the NTNU model is mostly hard rock conditions, but a few of the background data are from medium hard rock masses. The limitations of the models presented in the thesis may be as follows: •

Soft rock conditions, where NATM for example, is a more appropriate method

•

Categories 8 and 9 in the Q system or Q < 0.01 (Barton and Grimstad 1994), where the NATM and the so-called NMT (Norwegian Method of Tunnelling) methods may overlap one another and monitoring is required

•

In severe stability conditions, when heavy rock support or partial excavation is required for advancing; except for singular phenomena like faults or shear zones along a tunnel

In the latter cases, the model may be applied but due to heavy rock support and monitoring the advance rate is dramatically reduced.

5

2 D&B tunnelling

Drilling

Charging

Navigation Surveying

Blasting

Rock supporting

Ventilation

Scaling

Loading & hauling

Fig. 2.1 Drill and blast tunnelling cycle (Sandvik Tamrock Corp. 1999)

2.2 Development of the method During the past decades substantial developments have occurred in drill and blast tunnelling, mainly in drilling and explosives. Figure 2.3 from Atlas Copco clearly demonstrates the development of the drilling. It can be seen that from the beginning of the past century until now the drilling performance and capacity has increased from 5 to 450 drill metres per hour per operator. This is significant progress that has improved the economy and the efficiency of the method. According to Figure 2.2 the generations of rock drills from historical point of view are as follows: • • • •

Hand held drilling Pneumatic drilling Hydraulic drilling Computer-aided rock drilling

Fig. 2.2 Rock drilling generations (Tamrock Drills 1986)

6

2 D&B tunnelling Before pneumatic drilling, the drilling capacity was limited to some few metres per hour. By pneumatic rock drills the drilling capacity is increased up to 65 m/hr (Figure 2.3) using two-boom drilling jumbos. Introducing hydraulic rock drills in the 1970s was a milestone in drilling speed, economy and improvement of the working conditions. The drilling rate for the same two-boom jumbo reaches to 110 m/hr. Since 1980, the focus is set to improve the hydraulic rock drills towards more powerful ones, computerisation and automation together with increasing number of drilling hammers on the drilling jumbo. With a modern hydraulic rock drill and a four-boom drilling jumbo, production reaches 450 m/hr. On the explosives side, before invention of dynamite in 1866 by Alfred Nobel, black powder was used since the 17th century. Dynamite is a nitro-glycerine based explosive. Together with later developments of dynamite in the early 20th century, the electric initiation (with 1 second delay) was introduced in 1922. In late 1940s, short delay detonators 10 – 100 milliseconds, and in the late 1970s, the non-electrical initiation system was developed (Olofsson 2002). The millisecond detonators improved the control of the ground vibration and increased fragmentation. In 1955, ANFO (Ammonium Nitrate and Fuel Oil) was introduced. It was an inexpensive explosive but had low water resistance. In the 1960s, water gel and slurries were introduced, and in 1983 and the following years emulsion explosives were developed and gradually replaced slurry. The slurries partially solved the problem of water resistance and were less powerful than emulsions (Atlas Powder Company 1987). In 1997 and later, emulsion explosives and charging equipment were developed for underground and tunnel construction (Dyno Noble 2005). Emulsion explosives are water resistance, pumpable and produce less toxic gases.

2.3 Present state of the art 2.3.1 Drilling Currently, computerised hydraulic drilling jumbos with different level of automation are widely used in drill and blast tunnelling (Atlas Copco 2004, Sandvik Tamrock Corp. 1999, Korsman et al. 2006). The current generation of drilling jumbos is designed for high productivity, quality drilling, and comfortable working conditions for the operators.

7

2 D&B tunnelling The drill plan is stored in the drilling jumbo computer and therefore the face does not need to be marked up; the navigation is done by laser alignment and is very precise in deciding the position of the drilling jumbo. The excavated tunnel profiles can be scanned by electronic instruments, whereby overbreak and underbreak can be recorded to be used for the optimisation of the drilling process in the coming rounds (Nord 2003). Rock drills Obviously, invention of the hydraulic rock drill in the 1970s was a great progress in drilling speed, economy and improvement of the operator’s working conditions. Figure 2.3 supports the fact and shows the general development of the rock drills and jumbos.

Fig. 2.3 Drilling development 1905 – 2005, drilled metres per hour and operator (Atlas Copco 2006)

As an example, one may look at the Cop 3038. The latest rock drill from Atlas Copco is designed to increase the penetration rate in hard rock tunnelling. According to Atlas Copco; the Cop 3038 is 50 % faster than its predecessor Cop 1838 ME, the time saved with Cop 3038 compared to Cop1838 ME on 3-boom jumbos in a 90 m2 tunnel could be about 40 minutes per round. The high penetration rate may be utilised in limited space to replace 3-boom to 2-boom jumbos (Mining & Construction 2004).

8

2 D&B tunnelling The Cop 3038 delivers the same energy per percussive blow as the Cop 1838 ME, but the frequency has been increased from 60 to 102 Hz. Table 2.1 presents the development of Atlas Copco hydraulic rock drills for face drilling. Table 2.1 Development of Atlas Copco hydraulic rock drills for face drilling (Mining & Construction 2004) Type COP 1038HD COP 1238ME COP 1440 COP 1838ME COP 1838HF COP 3038

Year 1973 1983 1986 1990 2000 2004

Impact power 12 kW 12 kW 21 kW 18 kW 22 kW 30 kW

Impact rate 42 – 60 Hz 40 – 60 Hz 70 Hz 60 Hz 73 Hz 102 Hz

Drill steel Drill bits, drill rods and shank adapters may be referred to as the drill steel, which has concurrently been developed with the introduction of more powerful hydraulic rock drills. The new conical thread system, Magnum SR35, is devised to cope with the weakness of the old threads. The advantages of the new thread system are intended to solve rod breakage at the bit end, less tendency for deviation when collaring, straighter holes, and longer service life (Atlas Copco 2004). The rod length has also increased, even though the 18 feet rods are common in civil tunnelling and gives around 5 m round length, the tendency towards a longer round is increased (Nieminen 2003) and some tunnels have been excavated with 6 m round length (Tunnelling and Trenchless Construction 2005). In the NTNU experimental programme (Rønn and Johannessen 1995) using two rods the drilled length up to 30 feet (8.5 m round length) is recorded in the tunnel. But due to extra time consumption for rod adding the longer rounds are not common in current tunnelling. In underground excavations or mines where the multi-face excavation is required longer rounds may be useful. Such long blast rounds have also been experienced at the underground research laboratory in Canada (Kuczyk 2002). 2.3.2 Charging and explosives Emulsion explosives combined with non-electrical initiation system like NONEL have increased the safety of the charging and blasting operations and become more efficient.

9

2 D&B tunnelling The pumpable emulsion explosives, SSE (Site Sensitised Explosives) are not explosives until pumped in the hole, which means safety in transportation and handling (Johansen and Mathiesen 2000, Atlas Copco 2006). The modern emulsion explosives are oxygen-balanced, producing a minimum of noxious fumes and far less smoke (Olofsson 2002) and provide better working conditions and less ventilation requirements. The emulsion explosives are water resistant and can be used when water is a problem and ANFO can not be used efficiently. In addition, the pumpable emulsion explosives can be pumped into the blastholes by a computer controlled system and the amount of explosives in each hole and total explosives consumption can be measured, which can be used to control the amount of explosives in different holes in the face, especially in the contour and row nearest the contour where less explosives are needed for smooth blasting (NTNU 2007a, Zare and Bruland 2006a). The measuring system may also help to control the uncharged length of blastholes for optimum blasting results. Recently, after a feasibility study, a prototype machine for automatic charging of emulsion explosives was completed and tested in Norway (Hermann and Elvoy 2004). The project is meant to reduce the time consumption for drilling and charging by 20 % with improved work safety. The charging system is mounted on the drilling jumbo as a separate boom and charging is accomplished automatically during drilling without human interference at the face, in compliance with the regulations that prohibits simultaneous manual charging and drilling at the face. Due to advantages of emulsion explosives; a granular emulsion explosive has been developed in Japan. The granular emulsion explosive is made in the form of small cylinder of 4 mm diameter and 4 mm length and can be placed in the blast holes with a charging machine using compressed air (Taguchi et al. 2005). 2.3.3 Ventilation The harmful gases and dust particles above the certain level of acceptance must be removed by the ventilation system. For the cost efficiency and improvement of the work environment, an intelligent ventilation system is developed and tested in two tunnels (Lima and Blindheim 2004, Blindheim 2005). The project includes the development of new duct materials and duct

10

2 D&B tunnelling support system as well as continuous recording of air quality and automatic control of the ventilation fans. 2.3.4 Loading and hauling The loading and hauling capacity is dependent on the tunnel cross section size and excavation method. Using adequate size equipment as the cross section increases is an efficient way to increase the loading capacity. Field studies by NTNU in large cross sections (NTNU 2007b) reveal that a capacity of up to 500 lm3/h is attainable when using wheel loader and dump truck combination and the loader is fully utilised, i.e. the loader is not waiting for a hauling unit. Crusher, conveyor belt and mucking trains are also employed in drill and blast tunnels. The average capacity is 300 m3/h (understood as loose cubic metres). 500 m3/h should be achievable by the method in the future (Girmscheid and Schexnayder 2002). The crusher and conveyor belt are used as part of “high performance drill and blast excavation concept” which uses a suspended platform system for higher performance. Even though, the crusher capacity is the limiting factor, due to the other benefits of the method, the total cycle performance has been increased by 30 % (Girmscheid and Schexnayder 2002). 2.3.5 Rock support Rock bolts and sprayed concrete (shotcrete) are two common types of rock support, both are developed towards more durable, high strength, and quick installation by means of mechanised and automated equipment. The CT-bolt, the newest bolt type on the market can first be used as an ordinary endanchored bolt for temporary support and later on be grouted to a permanent bolt (Norwegian Tunnelling Society 2004). During the past two decades, sprayed concrete has had a significant improvement; today fibre-reinforced sprayed concrete using the wet mix method is widely applied. The sprayed concrete robot is one of the modern equipment for applying sprayed concrete. The capacity by manual spraying is less than 5 to 8 m3/h, while the robot reaches up to 20 m3/h (Girmscheid and Moser 2001). The robot has better safety and quality control.

11

2 D&B tunnelling Using the NTNU models, it is possible to estimate excavation time and cost including rock support, but since the rock support depends a lot on the rock mass conditions, the time and costs which are presented in the thesis do not include the rock support.

2.4 Future developments The overview of development in the previous sections, states the possibility of further development for the method, some areas of possible future developments are treated below. Drilling Drilling is one of the most likely areas of development. Using the current computerised drilling jumbo, the quality and quantity of drilling is substantially improved, providing better economy with increasing productivity and saving time. Further development may focus on: •

Full automation of the drilling jumbo so that the machine is operated by a remote device; without operator interference on the machine (European Construction Technology Platform 2005). This will improve working conditions, saving time and labour costs

•

Further development of rock drills or using new drilling methods like laser drilling to increase the drilling penetration rate and improve working conditions

•

Possibility of using longer rods if the cross section area and rock stability conditions allows. With current technology, the maximum single rod length is 6.4 m (21 feet) which gives approximately 6 m round length

•

Improving quality of the drill steel to make them more durable against abrasion, especially drill bits since they are in contact with the rock and are subject to abrade, therefore the bit changing time will be shortened or eliminated and the drilling rate will be improved

Charging and explosives •

Automation of the charging machine and implementation of charging while drilling as started in Norway (Hermann and Elvoy 2004)

•

Invention of less harmful or pollution-free explosives to improve working conditions and reduce ventilation costs

12

2 D&B tunnelling •

Increasing timing numbers and delay for detonators to get sufficient interval between holes for favourable confinement and control of vibrations that may be undesirable for nearby structures and cause damage to tunnel contours. This will improve the blasting performance and reduce the tunnel overbreak

Loading and transport Mucking out the excavated material by more efficient and parallel ways like crushing or using special substances to deform them in semi-liquid or liquid form and pumping them out; could be investigated in detail. In some projects in Europe (Girmscheid and Schexnayder 2002), the crusher has successfully been used in the tunnel. Even if the crushing capacity is still a limiting factor the total performance of the drill and blast method is increased by the use of the crusher and conveyor. Safety and environment Safety and minimum impact on the environment should be considered not only for each operation, but also as a global goal during tunnel construction. For safety, the number of tunnel crew can be reduced by automation. With the installation of some intelligent instruments in the tunnel; the accidents during construction may be reduced to some extent. Improving pre-investigation tools is another important issue regarding safety to make the ground as transparent as possible. Finding proper reuse for the excavated material and less influence on the natural environment above the underground work, related to flora and fauna will minimise the impact on the environment. Automated drill and blast system concept Explosives are more efficient for rock excavation than mechanical excavation due to lower energy consumption, but the non-continuous cyclic operation in the drill and blast method decreases the competitiveness of the method in small cross sections. Some attempts have been made to utilise the advantages of continuous operations. Automated and continuous drill and blast concept is developed for such purposes (Martin et al. 1989, Clark 1987). The concept is based on short holes, single hole blasting and continuous loading and blasting. The feasibility study was successful and a field test carried out, further work and research is required to implement the concept.

13

2 D&B tunnelling Tunnel factory concept It is a future possibility to develop the tunnel factory concept by using the drill and blast method. The crusher, conveyor belt and suspended platform are the parts of the excavation method. The rock will be transported by the conveyor belt or another continuous method to make the space behind the face available for other equipment. It is more or less a continuous method for excavation, rock support and technical installation etc. As in a TBM all the operations from the excavation until the final utilisation could be done using long backup equipment. The concept is subject to change if any technological development or new method is implemented in the future.

14

3 The NTNU model

3.1 Blast design model The NTNU blast design model is based on parallel hole cut. The tunnel face is divided into the cut, stoping (easers), lifters (invert), row nearest contour and contour. Smooth blasting with double contour blasting is recommended, i.e. the charging density in the contour and the row near the contour is reduced. The design for each part depends on the following rock and geometry parameters, which should be evaluated and determined in advance: • • • •

Rock blastability Drill hole diameter Drill hole length Skill level

Blast design for 48 mm and 64 mm drillhole diameters is given. For diameter in between or lower, the interpolation or extrapolation can be used. In current tunnelling in Norway; 48 mm drillhole diameter is common, in a few cases 51 mm and 45 mm drillholes are also used. The normal drillhole length in the model is 5 m. The 5 m drillhole is common in the current Norwegian tunnelling. A correction is given when the drillhole length is varied from 3 to 9 m. The correction factor for drillhole length depends on the drillhole length and skill level. The skill level refers to the tunnel crew and drilling equipment.

15

3 The NTNU model 3.1.1 Rock blastability The rock blastability is given by the rock blastability index, SPR, which is “the amount of explosives in one inch holes (kg/m3) needed to break the rock to a certain degree of fragmentation, where 50 % of the blasted rock size is under 250 mm (d50 = 250 mm)”. The SPR is developed based on blasting experience in surface rock blasting operations with bulk and cartridged explosives. The SPR is a regression based equation from field data. The rock blastability index SPR is determined as follows (NTNU 2007a):

SPR =

cn cp Ia c

0.736 ⋅ I a ⎛ c ⎞ ⎜ ⎟ ⎝ 1000 ⎠

0.4

0.6

⋅ LT 0.7

⎛ w⎞ ⋅⎜ ⎟ ⎝c⎠

0.25

⋅ ρ 0.2

= dry sonic velocity normal to foliation (m/s) = dry sonic velocity parallel to foliation (m/s) = cp/cn= anisotropy = (cp+cn)/2 = dry sonic velocity (m/s)

w

= detonation velocity of explosives (m/s) ρ = density of rock (g/cm3) LT = charging density of explosives (amount of explosives per volume unit of drillhole, g/cm3 )

The index relates the rock and explosives properties and is meant to aid the evaluation of blastability, and assumes access to laboratory data from a representative sample of the particular rock. Although the SPR is developed in the surface blasting, the relative SPR is valid for blastability evaluation in a tunnel. In the model, a classification is used to distinguish between different rock types, i.e. good, medium and poor blastability; referring to SPR = 0.38, 0.47 and 0.56 respectively. Another issue regarding blastability, is the effect of the degree of fracturing and fracturing orientation. The index does not take into consideration the variation of the rock mass fracturing and orientation of the fractures. “Local blastability scheme” suggested by Scott (1996) shows lack of a universal rock mass blastability index, due to variation in rock mass fracturing. In a tunnel, the degree and orientation of fracturing affect the performance of the blasting operation. The effect of joint parameters and explosives regarding overbreak and damage

16

3 The NTNU model control is investigated by Singh (2005) and Singh and Xavier (2005). Among the other things they found the orientation of 45 degrees is the most unfavourable with regard to contour quality. In general, when the degree of fracturing is very high and/or the fractures are open, poorer blastability is expected. Also when the orientation of fracturing is parallel or close to parallel to the tunnel axis, the rock mass blastability is reduced. 3.1.2 Cut design The cut is the most important part in tunnel blast design for creating an opening, to act as a second free surface. The cut performance is critical to achieve a satisfactory blasting and advance per round (pull). In the NTNU model, the large parallel hole cut is used. The necessary large hole area, distance between a large hole and first charged hole, and burden for the rest of the charged holes in the cut are given by curves and tables (NTNU 2007a). The empty large hole cut presupposes that the rock, which is blasted at each detonation interval, must have space for expansion (at least 80 %) to secure full throw out. This requires precise drilling and the correct firing sequence. For each detonation interval, one has to check that the existing opening gives space for the necessary expansion of the rock that will be blasted. This must be done by detailed calculations for the first two or three detonation intervals; visual evaluation is usually enough for the remaining holes. The cut is designed for an average pull of 90 % of the drillhole length for 48 mm drillhole diameter and 96 % for 64 mm drillhole diameter (field experience). 3.1.3 Drilling pattern After the cut has been designed, the design of the drilling pattern should follow the sequence:

• • • • •

Contour Row nearest the contour Invert Placement of the cut Stoping

Guiding burden and spacing for each part of the face are given for 48 mm and 64 mm drillhole diameters (NTNU 2007a). The burden and spacing are given at the bottom of 17

3 The NTNU model the round, at the face; eccentricity at the bottom of the holes (e.g. look-out angle) must be corrected for. The number of holes can be checked by the necessary number of holes curves. The curves assume contour product quality requirements. If protruding rock is allowed within the theoretical cross section (e.g. water tunnels), the number of holes may be reduced by up to 5 %. When contour blasting method is specified, the number of holes may be increased by up to 8 %. 3.1.4 Charging Tunnel rounds are usually charged with emulsion or ANFO, cartridged explosives may also be used; when water is a problem, ANFO can not be used efficiently. The contour is normally charged with special contour charges, e.g. tube charges or detonating cord. ANFO or emulsion may also be used in the contour and the row nearest contour when mechanised charging systems are used. The charging density in the contour with double contour blasting is 20-25 % and in the row nearest the contour it is 40-60 % of normal charging density. The “double contour” refers to the contour and the row nearest the contour. Necessary consumption of explosives for cartridged and bulk explosives is given in guidance curves as a function of tunnel cross section and rock mass blastability (NTNU 2007a). In NTNU’s experience, ANFO and emulsion explosives have approximately the same charging density. Uncharged length UL is defined as a function of the drill hole length L, uncharged length for the cut and lifter holes is UL = 0.1L and for other holes UL = 0.3L. With regard to the consumption of explosives, it is important to comply with the recommended uncharged length. Reduced uncharged length will result in increased explosives consumption, increased amount of undetonated explosives and poorer working conditions (fumes and particles). Increased uncharged length will result in poorer blasting result, fragmentation and loadability. 3.1.5 Firing pattern The firing pattern must be planned so that each hole or group of holes, gets as favourable confinement and throw conditions as possible. That is ensured by trying to establish a smaller version of the final cross section shape around the cut, and then enlarging this shape. It is also essential to check that the rock blasted at every interval number has enough space for expansion.

18

3 The NTNU model The general sequence is cut, stoping, row nearest the contour, contour, lifter and finally corner holes of the lifters or invert row.

3.2 Advance rate model The model is based on the round cycle time consumption. The round cycle is divided into four major operations: I II III IV

Drilling and charging Ventilation Loading and hauling Scaling and rock support

Operations I and III are divided into three different categories of time: A. Fixed lost time (rig time) All «unproductive» operations regularly repeated from round to round are collected here. The time consumption is also fixed in the sense that it is almost independent with regard to variations in round length, the number of crew members and the number of drilling hammers. Example: Driving the drilling jumbo to and from the face. B. Proportional operational time Proportional operational time is productive time, such as drilling and loading time. The time used is almost proportional to specific drilled metres and amount of broken rock, and inversely proportional to the number and performance of the drilling hammers or size and capacity of the loader. C. Incidental lost time The incidental lost time covers the technically dependent lost time occurring at random during tunnelling operations, for example machine breakdown. Personnel time and delays connected to change of shifts are also included here. A lost time of 6 minutes per hour is regarded as normal for well organised tunnelling. This constitutes 11.1 % of A + B, i.e. 10 % of total time consumption. 3.2.1 Drilling and charging The overall drilling time per round includes:

•

Drilling of charged and empty large holes 19

3 The NTNU model

• • • •

Moving between holes Changing of bits Lack of simultaneousness Rod adding (for drilled length longer than approximately 6 m )

A set of equations are presented (NTNU 2007b) to calculate the time consumption for the above items. Number of holes, drillhole diameter, drillhole length, type and number of rock drills and the rock type Drilling Rate Index, DRI (NTNU 1998) are the main parameters that influence drilling time. For a 63 m2 tunnel with 92 charged holes of 48 mm diameter and 5 m length, and 3 large holes of 102 mm, using a computerised jumbo with 3 drilling hammers of Cop 1838, the total drilling time is estimated 150 minutes per round. The charging time depends on:

• • •

Number of holes Drillhole diameter and length Explosives type and number of charging lines

For a 63 m2 tunnel with 92 charged holes of 48 mm diameter and 5 m length, when the holes are charged by two charging lines, the charging time varies from 60 to 85 minutes per round depending on type of explosives. 3.2.2 Ventilation A ventilation break is a necessary break in the round cycle for diluting and removing the blasting fumes at the face. This is the time from blasting of the round until the concentration of mainly nitrous gases (NOX) at the tunnel face is under the Threshold Limit Value, TLV = 2 ppm. A ventilation break varies from 5 to 30 minutes depending on tunnel cross section and type of explosives. 3.2.3 Loading and hauling In general, loading and hauling time is dependent on the volume of blasted rock per round and loading capacity of the loading and hauling equipment combination. The loading time is calculated by dividing the volume of blasted rock by the loading capacity.

20

3 The NTNU model The tunnel cross section, excavation method and the combination of loading and hauling equipment are the main parameters that influence the loading capacity. The normalised loading capacity for different equipment combinations is given based on field studies and normalisation (NTNU 2007b). For a 63 m2 tunnel and 5 m round length, the loading and hauling time may vary from 100 to 150 minutes per round. 3.2.4 Scaling and rock support The scaling time covers time for scaling the round and checking the rock face to allow further work. The time for rock support is not included in the scaling time. The scaling time depends on the tunnel cross section, scaling method and rock blastability. For medium blastability and 5 m round length, the scaling time varies from 10 to 90 min depending on the tunnel cross section. This may be influenced by especial careful contour blast design. When using continuous rock support with bolts and/or shotcrete, it is common to include rock support time in the round cycle. By utilizing the time when there is no excavation, e.g. during the night, the shotcrete can be sprayed without being a time determinant. Time consumption for polyester or point/end anchored bolts as a function of number of bolts per round is given in NTNU 2007b both for bolt drilling time and bolt mounting time. 3.2.5 Weekly advance rate For the weekly advance rate there is differentiation between the net, standard and gross advance rates, derived from the corresponding round cycle time consumption. The net advance rate is understood as the advance rate achieved for well organised tunnelling excluding time for blasting of niches, correction for job-training and tunnel length, rock support etc. The standard advance rate is determined based on net round cycle time with additional time consumption for blasting of necessary niches and correction for tunnel length and job-training effect. The gross advance rate is determined on the basis of the standard round cycle with additional time consumption for rock support and unforeseen, depending on site conditions. The standard weekly advance rate as a function of the tunnel cross section for 3 km tunnel length is shown in Figure 3.1. The following assumptions are considered:

21

3 The NTNU model 150 1 2 3 4 5 6 7 8

140 130 m/week 120

Haggloader - Shuttlecar, 2 drilling hammers Cat 972G - Load&haul, 2 drilling hammers Cat 980G - Load&haul, 3 drilling hammers Cat 973C - Truck, 3 drilling hammers Cat 966G/972G - Truck, 3 drilling hammers Cat 980G - Truck, 3 drilling hammers Volvo L330E - 35t dump truck, 3 drilling hammers Volvo L330E - 35t dump truck, 4 drilling hammers

110 1

100

3 5

2

90

4

6

7

80 8

70 60 50 40 0

10

20

30

40

50

60

70

80

90

100

110

120

Cross section, m

130 2

Fig. 3.1 Standard weekly advance rate as a function of tunnel cross-section

Rock conditions Rock blastability, rock drillability and rock wear quality influence the advance rate. The numbers of charged and empty holes are dependent to rock blastability. The drilling time is dependent to rock drillability and rock wear quality. Standard weekly advance rate is calculated for medium blastability, SPR = 0.47, medium drillability, DRI = 49, and medium rock wear quality, VHNR = 550. Blast design parameters The NTNU blast design model (NTNU 2007a, Zare and Bruland 2006a) is used to calculate the number of charged and empty holes for different tunnel cross sections. Parallel hole cut is assumed, the diameter of the empty hole(s) are 102 mm and the diameter of charged holes are 48 mm. The round length is 5m and the advance per round is assumed to be 91 % of drilled length.

22

3 The NTNU model Drilling and charging Nowadays, computerised drilling jumbos with different levels of automation are widely used in drill and blast tunnelling (Atlas Copco 2004). The drilling time is dependent on the type and number of drilling hammers. 2 drilling hammers for cross section less than 20 m2, 3 drilling hammers for cross sections 20 - 80 m2 and 4 drilling hammers for cross sections larger 80 m2 are assumed. The penetration rate is calculated based on the field performance data for COP 1838 rock drills. Charging is assumed to take place after drilling: 2 charging lines for cross sections less than 80 m2 and 3 charging lines for cross sections greater than 80 m2, ANFO explosives for track tunneling and emulsion explosives for trackless tunnelling, are assumed. Loading and hauling It is assumed that the numbers of hauling units are sufficient to fully utilise the loading equipment. The advance rate for the most efficient equipment combination is presented. Working time per week In Norwegian tunnelling, the tunnel is normally excavated in 5 days per week, two shifts per day and 10 hours per shift. This results in an average of 101 working hours per week during a year. AS Figure 3.1 shows, as the cross section increase, the advance rate is reduced, due to increase in volume of the blasted rock and number of holes. Using larger equipment with higher loading and drilling capacity; causes upward “jumps” in the beginning of each curve, indicating a higher advance rate. Figure 3.1 can also be used for choosing an efficient excavation method relevant to the cross section size. As can be seen for cross sections larger 16 m2, the load and haul method has higher advance rate than the track tunnelling method. When the cross section is larger than 30 m2, the advance rate for direct loading is higher than for the load and haul method. This implies suitable excavation methods as follows:

• • •

Tunnel less than 16 m2 Tunnel between 16 m2 and 30 m2 Tunnel larger than 30 m2

: track tunnelling : trackless tunnelling, load and haul : trackless tunnelling, direct loading

In comparison with the 1995 model (NTNU 1995b), the advance rate model has been updated to 48 mm drill holes. In general, the advance rate is increased for large cross

23

3 The NTNU model sections. Due to the combination of positive and negative effects of changes in the round cycle elements, the advance rate is not changed significantly for small cross sections. Some positive changes are: reduced charging time due to use of emulsion explosives instead of ANFO for trackless tunnelling, and increased loading capacity for trackless tunnelling larger than 50 m2 and reduced ventilation time. Some negative changes are: increased niche excavation time for small cross sections and increased scaling time, especially for poor blastability conditions.

3.3 Cost model The cost model is based on detailed cost calculation for excavation operations as follows:

• • • • • • • •

Drilling Charging Scaling Loading Hauling Additional work Labour Niches

The report 2C 05 (NTNU 2007c) shows in more detail the cost items included and not included in the excavation costs. Equipment and material prices, labour wages and the expected lifetime for equipment are the main input for the cost calculation (Appendix B). Depreciation, interest, repairs, downtime cost and power consumption are included in the cost of each operation. Drilling costs include drilling jumbo and drill steel costs. Drill bit, drill rod, shank adapter and coupling are considered as drill steel. The charging costs include explosives and initiation system costs, the charging machine and charging equipment operator costs are included as part of the explosives price. Hauling costs include muck transport, roadway/rail and dump site costs. Transport by subcontractor is assumed for trackless tunnelling, while detailed cost estimation is performed for track tunnelling. Additional work refers to ventilation, electrical installations and water supply. The cost of each operation is estimated without labour. The labour costs are estimated as a separate item covering all the tunnelling operations in the tunnel and at the surface. 24

3 The NTNU model In addition, 10 % extra for unforeseen and uncertain assumptions is added to the sum of detailed costs (see Figure 3.4). As in the time model, the rock support is not included in the standard excavation costs, in order to provide the basis to compare the costs for different tunnel cross sections. The amount of rock support depends a lot on the rock conditions and tunnel application. In the printed edition of the cost model (NTNU 2007c), all input and background details are not shown. In the simulation tool (Chapter 4), all the input data and calculation processes are shown, and the user has opportunity to give his own experience data. 3.3.1 Cost calculations The complete detailed cost calculation for each operation is presented in the background material of the1988 model (NTNU 1988). The general machine cost model is described in the Project Report 15A-92 (NTNU 1992). The machine cost includes:

• • • • •

Depreciation Interest Repair Downtime Operating costs (Service, fuel and tyres)

To calculate the machine cost, the machine lifetime is important. The calculation is based on economic useful lifetime, tB in Figure 3.2, where the cumulated average unit cost is at a minimum value. The economic useful life time is derived mathematically from the cumulated average unit costs shown in Figure 3.2, for details see NTNU 1992.

tB =

IA tL = (1 + k ) ⋅ β 1+ k

tB

=

economic useful lifetime

IA

=

depreciation basis

k

=

downtime factor

β

=

repair factor

tL

=

economic lifetime

25

3 The NTNU model

NOK/eh

Depreciation and interest cost

Dow n

time c

osts

Repair costs Operation costs tB

eh

Fig. 3.2 Economic useful lifetime, tB

The total time of the machine used in the tunnel or the time until a major overhaul, is considered as a percentage of economic useful lifetime, varying for different machines. The economic lifetime, downtime factor and time for major overhaul is given in Table 3.1. Table 3.1 Economic lifetime, downtime factor and major overhaul time Machine

Economic lifetime (eh)

Downtime factor

Major overhaul time

6000 eh/hammer

0.5

75 % tB

9000 - 14500

0.5

50 % tB

Excavator

20000

0.5

50 % tB

Haggloader

9000

0.5

66 % tB

Locomotive

13000 - 16000

0.5

75 % tB

Shuttle car

18000 - 19500

0.2

100 % tB

Drilling jumbo Wheel/track loader

In the cost model the machines cost are depreciated over the economic useful lifetime. As a result of this, machines will have a rest value until tB is reached. The mathematical

26

3 The NTNU model expression of the rest value is derived from Figure 3.2 (see NTNU 1992 for more details) as follows:

R = 100 ⋅ (

tB − x 2 ) tB

R

=

rest value

x

=

part of tB

The non-depreciated value or rest value is shown in Figure 3.3, which is the total machine price, IA, minus all depreciation. The depreciation and interest are calculated until major overhaul time, tR, in Figure 3.3. For example, when tR is 75 % of tB, 94 % of IA is depreciated and considered for interest calculation, and when tR is 50 % of tB, 75 % of IA is depreciated. The downtime costs are costs that appear in addition to the repair costs when a machine stops. This can be costs regarding loss of production and other machinery participating in the production. Downtime costs are determined as a percentage of repair cost, downtime factor is given in Table 3.1. The equipment such as fans, ventilation duct, transformers, cable, water pumps, pipes and rails are depreciated linearly, 50 % on lifetime and 50 % on reuse basis.

A

R e s t v a lu e

I

I R

t

t R

L ife tim

B

e

Fig. 3.3 Rest value as a function of lifetime

27

3 The NTNU model 3.3.2 Cost model summary

The summary of the detailed cost model in the form of unit excavation cost as a function of tunnel cross section area is given in Figure 3.4, the assumptions are as follows: • • • • • •

48 mm drillhole diameter Round length 5.0 m Tunnel length 3 km, horizontal adit ANFO/emulsion with 5 % dynamite Medium drillability and blastability 10 % extra for unforeseen

NOK/m

1 Track tunnelling 2 Trackless tunnelling, Load&haul 3 Trackless tunnelling, Direct loading

17000

15000

13000

11000

9000

3

7000 2 1

5000

3000 0

10

20

30

40

50

60

70

80

90

100

110

120

Cross section, m

Fig. 3.4 Unit excavation cost as a function of tunnel cross section area

28

130 2

3 The NTNU model The machine combination is chosen according to the tunnel cross section area and excavation method, i.e. track tunnelling for cross section up to 16 m2 and trackless tunnelling, load and haul for cross section 16 – 30 m2 and trackless tunnelling, direct loading for cross section between 30 - 120 m2. The unit cost per tunnel metre is increased as the tunnel cross section area increases. A change in the excavation method causes downward “jumps” because of increase of the advance rate, indicating more cost-effective tunnelling. The curve in Figure 3.4 also shows the most economic excavation method for different tunnel cross section areas. For cross sections larger than 16 m2, the trackless load and haul method is cheaper than track tunnelling. For cross sections larger than 30 m2, the trackless direct loading is the most economic excavation method. In general, estimated cost by the present model is lower than the 1995 model (NTNU 1995c). As an example, the standard cost for a 60 m2 and 5 km long tunnel is 11300 NOK/m. the estimated cost for the same tunnel by the 1995 model corrected for price level is 12300 NOK/m. In comparison to the 1995 model, the following change and developments have been included in the model in addition to new machine and equipment prices and lifetimes: •

The cost model is developed for 48 mm drillhole diameter; the 1995 model was based on 45 mm drillhole.

•

In general, the advance rate is increased due to more efficient machines.

•

The correction factor for job training and tunnel length effect is applied on standard advance rate. In the 1995 model, the correction was applied on cost items such as drilling, charging, scaling and loading only.

•

The investment tax is reduced from 7 % to 0 %, which is also included in the general price level correction factor.

•

The percentage of primer (dynamite) for charging is reduced from 10 % to 5 %.

In Figure 3.5, the cost per actual solid cubic metre as a function of tunnel cross section area is presented. The degressive trend shows the role played by the confinement and equipment with higher capacity. As the cross section increases, the confinement is reduced and equipment with higher capacity can be employed. This leads to dramatically reduced unit cost.

29

3 The NTNU model

NOK / asm

3

1000

800

600

400

200

0 0

20

40

60

80

100 Cross section, m

Fig. 3.5 Cost of actual solid cubic metres as a function of tunnel cross section area

30

120 2

3 The NTNU model

3.4 Field data The field data are used to verify and update the blast design and advance rate models, especially the advance rate model. For this purpose, the field data are classified into different subcategories of round cycle time, i.e. drilling, charging, loading and hauling, ventilation and scaling, as presented in Appendix A. The data are collected from finished or ongoing projects in Norway. The data represent some basic principles of the models, such as drillhole diameter, drillhole length, parallel hole cut and type of explosives and detonators. The data are mainly collected through master’s theses at the Department of Civil and Transport Engineering of NTNU (Nordbotn 2004, Flage 2002, Austlid & Andreassen 1998, Thorkildsen & Vilhelmshaugen 1996, Nilsen 1998, Log 1998, Sørstrøm 2005, Paulsen 2005, Andersen 2000). The department has a more or less continuous involvement with the tunnelling industry in Norway for data acquisition and field studies. The master’s students are part of this link between the department and tunnel projects. In general the field data are average of more than one, up to several observations or measurements. The numbers of data vary substantially between various operations and tunnel sites. Due to variation in number of data the variation is not shown in the figures, only the average data are shown. In the following figures, the model data are estimated based on parameters corresponding to actual conditions regarding drillability, blastability, etc. the main purpose of the this is to verify the goodness of the model and also to explain deviation between new field data and model. For simplification, the linear regression has been used to compare the field data and model. For some of the figures the non-linear regression would obviously give a better fit. 3.4.1 Blast design

Figures 3.6 and 3.7 show the actual data versus model predicted data for blast design results. It can be seen that the trend of the predicted values for both number of holes and explosives consumption is the same as for the actual data. Since the model gives the necessary number of holes and explosives consumption, the actual values are generally higher than the model values, especially for explosives consumption. We found that a shorter uncharged length and spillage are the main reasons for higher explosives consumption in the field. Also, the actual number of holes and explosives

31

3 The NTNU model consumption is decided by the tunnel crew, they tend to include some extra holes and explosives for “security” of successful blasts. Generally, for medium and larger cross sections, the correlation between the actual and model data is better than for small cross sections; these refer to the model background data which are mostly from medium to large cross sections.

Number of holes

200

Field data Model Linear (Field data) Linear (Model)

150

100

50

0 0

20

40

60

80

100

120

Cross section, m

2

Fig. 3.6 Number of drilling holes, field data vs. model

6 Explosives consumption, kg/asm

3

Field data Model Linear (Field data) Linear (Model)

5 4 3 2 1 0 0

20

40

60

80

100 Cross section, m

Fig. 3.7 Bulk explosives consumption, field data vs. model

32

120 2

3 The NTNU model 3.4.2 Advance rate

The actual time consumption versus model predicted time consumption for major operations is shown in Figures 3.8 to 3.12. The rig time and incidental lost time are included in the drilling, charging and loading time. Net penetration rate and loading capacity in the model are assumed to be identical to the field performance data. For drilling, the actual time consumption is lower than the model (Figure 3.8), one reason is 10 % incidental lost time for all cross sections, while the incidental lost time should be less than 10 % for large cross sections. The 10 % incidental lost time has been measured in tunnels with medium cross sections. For charging (Figure 3.9), the model predicted times are less, but the amount of data is too few for large cross sections. The actual loading times are higher than the model (Figure 3.10), indicating higher incidental lost time in reality. Regarding ventilation and scaling times, there is a quite consistent trend between actual and model values (Figures 3.11 and 3.12); the model gives higher ventilation time and lower scaling time. The higher model ventilation time is to cover longest ventilation time for safety reasons. The higher explosives consumption in reality (Figure 3.7) and different contour quality are reasons for higher actual scaling time. In Figure 3.11, zero ventilation time indicates the ventilation is not time-determinant and is done in parallel to other operations like rigging. In general, except for the drilling and ventilation times, the actual times for charging, loading and scaling are higher than the model predicted times. These differences are of minor interest since they somewhat neutralise each other in the total round cycle time consumption, as can be seen in Figure 3.13, where the actual and predicted standard round cycle time consumption is presented. Figure 3.13 indicates better correlation for medium to large cross sections than small cross sections. This is chiefly due to the model background data being mostly from tunnels with medium to large cross sections, and partly due to 10 % incidental lost time which is also based on the field measurements in medium cross sections. On the other hand, the amount of field data for small cross sections is fewer than for larger cross sections, which may partly affect the results.

33

3 The NTNU model 300

Field data

Drilling time, min

Model Linear (Field data) Linear (Model)

250

200

150

100

50 0

20

40

60

80

100

120

Cros s s ection, m 2

Fig. 3.8 Drilling time, field data vs. model

Charging time, min

200

Field data Model Linear (Field data) Linear (Model)

150

100

50

0 50

75

100

125

150 Num ber of holes

Fig. 3.9 Charging time, field data vs. model

34

175

3 The NTNU model

Loading time, min

300

Field data Model Linear (Field data) Linear (Model)

250

200

150

100

50 0

20

40

60

80

100 Cross section, m

120 2

Fig. 3.10 loading time, field data vs. model

Ventilation time, min

50 Field data Model Linear (Field data) Linear (Model)

25

0 0

20

40

60

80

100 Cross section, m

120 2

Fig. 3.11 Ventilation time, field data vs. model

35

3 The NTNU model

Scaling time, min

250

Field data Model Linear (Field data) Linear (Model)

200

150

100

50

0 0

20

40

60

80

100 Cross section, m

120 2

Fig. 3.12 Scaling time, field data vs. model

Round cycle time, min

800

Field data Model

700

Linear (Field data) Linear (Model)

600 500 400 300 200 0

20

40

60

80

100 Cros s s ection, m 2

Fig. 3.13 Standard round cycle time, field data vs. model

36

120

3 The NTNU model

3.5 Model developments The first version of the model was published in 1975 (NTNU 1975a,b). From the first version the model has been updated and developed five times in 1979, 1983, 1988, 1995 and in 2005 through the author thesis work. In the following, a brief description of the major developments in the model is given, including the latest developments in 2005. The development of the model is parallel to the introduction of new equipment, methods and explosives. The trend of predicted time and cost is presented from the first to the latest version. 3.5.1 Overview from 1975 to 1995

1975 The model was published for two different combinations of drillhole diameter and length. 34 mm drillhole with 2.1 m standard drilled length and 45 mm drillhole with 2.7 m standard drilled length. The necessary drilling and explosives consumption were given for both diameters, supplemented by a correction factor when the drilled length is varied. The model applied to tunnel cross sections up to 80 m2. The net penetration rate was 30 – 140 cm/min for medium drillability and different rock drills (pneumatic and hydraulic drills), the maximum net penetration belongs to hydraulic rock drill Cop 1038. The gross loading capacity for trackless tunelling was 25 asm3/h for 16 m2 and 160 asm3/h for 80 m2. The excavation time and costs when using the hydraulic drilling jumbo were presented in the report Hydraulic Drilling in Tunnel (NTNU 1976) . 1979 The model was published based on 45 mm drillhole diameter, standard drilled length is increased from 2.7 m to 3.4 m. In addition to cartridged explosives, a specific charging curve for Anolit has also been included in the model. The net penetration for DRI = 50 was 70 – 155 cm/min, the maximum value belongs to Cop 1038 rock drills. The normalised gross loading capacity for hydraulic excavator is given; the maximum capacity was 180 asm3/h for Brøyt X50 and 35 ton truck. The scaling time is separated into scaling from the muck pile and use of scaling jumbo.

37

3 The NTNU model 1983 The standard drilled length was increased from 3.4 m to 3.7 m; the drillhole diameter remains 45 mm. The maximum loading capacity for an excavator is increased to 235 asm3/h for Brøyt X50 and 35 ton truck. 1988 The rock blastability index, SPR is introduced to the model for distinguishing between good, medium and poor blastability. In the previous models, good and poor blastability were distinguished by description and samples. The net penetration rate for DRI = 50 was 70 – 200 cm/min, the maximum value belongs to the Cop 1440 rock drill. The maximum loading capacity for an excavator is increased to 250 asm3/h for Brøyt X52WF and 35 ton truck. 1995 The blast design model was published as a separate volume (NTNU 1995a). The model was presented for 45 mm and 64 mm drillhole diameter and standard drilled length was increased from 3.7 m to 5 m. Many details and descriptions were added to the blast design model such as the tunnel cross section design, details of rock blastability index, details of cut design, firing pattern and details of drilling pattern in the contour, the row nearest to the contour, easer and invert holes. Emulsion is introduced as explosives in blast design model. The net penetration rate curves were updated, the penetration reached to 230 cm/min for medium drillability, DRI = 50 and Cop 1838 rock drill. The drilling time is divided to more sub-operations with separate equations to estimate time consumption for each operation. The charging time was updated and the charging time for emulsion explosives is added to the charging curve. The gross loading capacity curves are updated, the maximum loading capacity increased to 260 asm3/h, that is belongs to Volvo L330E and 35 ton truck. The time consumption for rock support, including rock bolt and shotcrete are added to the model. 3.5.2 Latest developments, 2005

The model is developed based on new field data and up-to-date technology, costs and new regulations representing 2005 level. New field data are collected by the various

38

3 The NTNU model master students (see Section 3.4), as well as information from contractors and suppliers (see Preface and data in Appendix B). The following describes the work and contribution by the author. The model is updated to the most commonly used 48 mm drillhole diameter, which affects the complete model from blast design to advance rate and costs; representing a large effort of re-modelling, cross-checking, quality control, updating of graphs, etc. In addition, as an alternative to paper edition, for fast calculation with possibility to vary input by user experienced data; a simulation tool, TunSim is developed in Excel. Some main developments in the model are given as follows: Blast design The tunnel cross section design is revised by the new standard (Norwegian Public Road Administration 2004), the rock blastability index, SPR, equation for emulsion explosives and SPR values tested on the samples in the laboratory are added to the model. Cut design, including the necessary large hole area and distance to the first and following holes are updated for the 48 mm drillhole diameter. The number of drilled holes, cartridged and bulk explosives consumption and correction factor for drilled length are also updated based on the 48 mm drillhole diameter. Many explanations with revised and updated examples are added to make the model more comprehensive and easy for application. Advance rate The net penetration rate and correction factor for different hole diameters are presented based on 48 mm drillholes. The unit time for bit changing has increased to 3 min, the equations in the model are presented in symbolic form and examples are updated. The charging time is updated for 48 mm drillholes. According to field data, the ventilation time is reduced and the ventilation time for emulsion explosives is added to the model. The overbreak factor and rig time for track transport, loading capacities and scaling time are updated. A method for estimating extra excavation time for niches is added to the model. The net, standard and gross weekly advance rates are introduced, the correction for jobtraining and tunnel length is added to weekly advance rate. The niche time is included for all operations in the model, some practical changes are made regarding number of hammers and charging lines and new combinations are given for loading and hauling machines.

39

3 The NTNU model Costs As the blast design and advance rate models are updated, consequently the cost model should be updated. The equipment, tool and material prices and lifetimes are collected from the related suppliers and companies. Using the new input, the whole cost model is updated; this includes drilling, charging, scaling, loading, hauling, additional work, labour and niches. The related curves and correction factors are presented based on June 2005 price level. The cost is given in detailed, summary and total construction sections. Simulation tool The simulation tool (TunSim) is developed in Excel and can be used for fast and exact calculation of time and cost. The tool is capable of estimating the time and cost for all tunnel cross sections and excavation methods. The main inputs are tunnel geometry, blast design parameters, rock properties and equipment type and number. The simulation tool can be used as an alternative to paper versions for blast design, advance rate and cost estimates. The users can also change the input by their own experienced data, a feature not available in the paper version. Furthermore, the tool may be used for risk analysis in time or cost. Further details on the simulation tool are given in Chapter 4. 3.5.3 Time and cost trends

The development of the net penetration rate is shown in Figure 3.14. The net penetration is presented for best hammer and medium drillability. Figure 3.14 can also be considered as the development of the hydraulic drilling hammers from beginning to latest model. The penetration rate is increased over two times during the past three decades. The development of gross loading capacity is shown in Figure 3.15. The loading capacity is given for different excavation methods. The typical cross sections of 15 m2, 25 m2 and 80 m2, correspond to track tunnelling, load and haul and direct loading methods respectively. First, the effect of tunnel size on loading capacity is obvious. As the cross section increases the higher loading capacity is expected. Second, for large cross sections the loading capacity is more continuously increased from 160 asm3/h to 280 asm3/h (a 75 % increase). For track tunnelling and load and haul methods the loading capacity remained unchanged for many years and increased around 60 %.

40

3 The NTNU model 350

Net penetration, cm/min

300 250 200 150 100 50 0 1975

1979

1983

1988

1995

2005 Year

Fig. 3.14 Development of net penetration rate

15 m2

25 m2

80 m2

3

Gross loading capacity, asm /h

300 250 200 150 100 50 0 1975

1979

1983

1988

1995

2005 Year

Fig. 3.15 Development of gross loading capacity

To investigate time and cost trends or productivity and efficiency of the drill and blast method, a 60 m2 tunnel with 3 km length is chosen as a typical two-way road tunnel. Medium rock blastability and drillability are assumed. The advance rate per week and unit costs per tunnel metre without rock support installation are estimated from 1975 to 2005 by the NTNU models.

41

3 The NTNU model Weekly advance rate is shown in Figure 3.16. The weekly advance rate is normalised based on 100 hours working time per week. The advance rate is increased from some 50 m/week to 81 m/week during the past 30 years, indicating a 60 % increase in production rate. The excavation method itself does not change significantly, development in the equipment capacity and efficiency, tools and materials leads to that the current tunnel will be excavated faster and the excavation time is shortened. For example, in 1975, the maximum penetration rate for medium drillability was 140 cm/min whilst in the recent model of 2005, it reaches 300 cm/min and the maximum loading capacity for same size cross sections is increased from 160 asm3/h to 280 asm3/h.

90 80

m / week

70 60 50 40 30 20 10 0 1975

1979

1983

1988

1995

2005 Year

Fig. 3.16 Development of weekly advance rate for a 60 m2 tunnel

The uncorrected excavation costs are shown in Figure 3.17, the unit cost is presented in Norwegian kroner per metre of tunnel. The cost in Figure 3.17 is not corrected for price level; only representing the unit excavation costs in each year. To find the comparable figures, the cost must be corrected for price level. The Department of Civil and Transport Engineering at NTNU has published a cost index for construction equipment (NTNU 2006) since 1978. The corrected excavation costs are shown in Figure 3.18. The trend of the excavation cost is not consistent compared to the time trend (Figure 3.16). One reason is different working hours per week; which is 112 hours in 1975, 75 hours in 1979 and 1983 models, and 100 hours for the rest. Based on 100 working hours, the costs for 1975 model should be higher and for the 1979 and 1983 models, lower than the presented values. 42

3 The NTNU model In spite of this minor deviation, the excavation cost has decreased from 16000 NOK/m to 10200 NOK/m, indicating a 36 % reduction. The technological development is the main reason for the decreasing cost trend. Changes in the investment tax and interest rate are included in the price level correction.

12000

NOK / m

10000

8000

6000

4000

2000

0 1975

1979

1983

1988

1995

2005 Year

Fig. 3.17 Development of uncorrected excavation cost for a 60 m2 tunnel

18000 16000

NOK / m

14000 12000 10000 8000 6000 4000 2000 0 1975

1979

1983

1988

1995

2005 Year

Fig. 3.18 Development of excavation costs for a 60 m2 tunnel, price level June 2005

43

3 The NTNU model

44

4 Simulation tool

4.1 Introduction Today spreadsheet calculations are widely used in different field of science and technology. In order to facilitate fast calculation of the advance rate and cost models, a new simulation tool is developed in the Excel software. The simulation tool, TunSim is based on the development of prediction models. A typical user of the simulation tool will utilise the models as follows: • • •

As an alternative to the paper version for fast and exact calculation Analysing the effect of variation in one or more of the input parameters or to analyse risk in time or costs by default or own experience data Estimate the advance rate and costs at different levels. On the top level by the tunnel cross section and length only; On a more detailed level by replacing defaults by own experience data

The excavation time and cost for any drill and blast tunnel are the final result of the simulation tool. On the top level, the user only needs to input the tunnel cross section and length, all other time and cost input data have built-in default or recommended values, and the excavation time and cost can be estimated. If necessary, the users can change the time and cost input by their own experience data. The simulation is done in 20 different Excel sheets, which are: • •

Input data Advance rate model

45

4 Simulation tool • • • • • • • • • • • • • • • • • •

General cost input Drilling cost Explosive and detonator cost Scaling cost Loading cost Contract transport cost Roadway cost Track transport cost Rail cost Tip cost Ventilation design Ventilation cost Electrical installations cost Water supply cost Labour cost Niches cost Cost summary Output

The general specifications of the simulation tool are as follows: • • • • • • • • • • •

Applicable for tunnel cross section from 6 m2 to 120 m2 Applicable for any tunnel length between 100 m and 10000 m Applicable for different excavation methods; which are track tunnelling, trackless tunnelling load and haul and trackless tunnelling direct loading Considering rock blastability, drillability and rock wear properties Applicable for drillhole length up to 6 m Applicable for different drillhole diameter; i.e. 45, 48, 51, 57 and 64 mm Applicable for different empty hole diameter; i.e. 76, 89, 102, 115 and 127 mm Built-in calculation for blast design and ventilation design parameters Considering the drilling jumbo rock drill type and number of hammers Considering different explosives types; i.e. cartridged, ANFO and emulsion explosives Considering different loading and transport equipment combinations

The list of input data is given in Table 4.1, each input has a default value and if necessary, the users can change the values by using the drop-down lists or by entering

46

4 Simulation tool their own experience data. Most of the data are used in both the time and cost model. To consider user value instead of default/model value, the calculation is based on an extra hidden column after the user column. When the user has an entry instead of the model value, the user value is replaced to this extra column by combination of IF and ISNUMBER or IF and ISBLANK functions. The results or output of the time and cost simulations are given in Table 4.2, the results are presented for excavation time and cost. The simulation tool can be easily used in drill and blast tunnelling industry by owners, contractors and consulting companies in all phases of a tunnel project.

Table 4.1 Simulation tool input General input data

Unit

Model

User

Tunnel data Tunnel cross section

m2

60.0

Tunnel length

m

3,000

Drillhole length

m

5.00

Drillhole diameter

mm

48

Diameter large drillholes

mm

102

Rock mass data Blastability

Medium

Drillability

Medium

DRI

49

Rock wear quality, VHNR

550

General Skill level Excavation method

High Trackless

Drilling, charging Type of drilling hammers Number of drilling hammers Explosives type

COP 1838 3 Emulsion

Loading, transport Loading machine

Volvo L330E

Transport machine

Dump truck

47

4 Simulation tool Table 4.2 Simulation tool output General output results

Unit 2

Value

Tunnel cross section

m

60.0

Tunnel length

m

3,000

Drillhole length

m

5.00

Drillhole diameter

mm

Excavation method

48 Trackless

Standard weekly advance rate

m/week

76.3

Gross weekly advance rate

m/week

76.3

Standard costs

NOK/m

10,511

Rock support costs

NOK/m

0

Total costs

NOK/m

10,511

4.2 Advance rate model On the basis of the advance rate model (NTNU 2007b), the input is divided into seven groups i.e. blast design, drilling and charging, ventilation, loading and hauling, scaling, niches and excavation time. The list of input for each group is given in Table 4.3; the value of each input is calculated based on the main input data presented in Table 4.1 and by using the blast design and/or the advance rate model. To do this the blast design and advance rate data are analysed in Excel, by using Excel curve fitting tools (trendline option). The proper equations are derived for each set of data. The equations are in the form of linear, polynomial or power, depending on best fit according to regression values (R-square). Many Excel functions or combinations of them have been used in the simulation tool, like IF, nested IF, LOOKUP and FORECAST. Due to variation in input parameters, like possible drillhole diameter from 45 to 64 mm; the FORECAST function has been used for interpolation or extrapolation. The LOOKUP function has frequently been used to find corresponding results according the input value, e.g. allocating the proper value of poor, medium or good blastability, loading capacity for a specific loader or penetration rate for different rock drills.

48

4 Simulation tool Table 4.3 Input data for advance rate Advance rate input data

Unit

Model

User

Blast design Number of drillholes, standard round length

89

Correction for drilled length

1.00

Number of large drillholes

3

Drilling, charging Penetration rate 45 mm drillhole

cm/min

228

Correction for drillhole diameter

%

96.0

Correction for large hole diameter

%

40.9

Time for moving per hole

min

0.75

Rod adding Rod adding time per hole

No min

0.0

Bit changing factor

0.022

Lack of simultaneousness factor

0.066

Number of charging lines Charging time basic round length

2 min

Correction for drilled length

1.00

Use of service platform on drilling jumbo for charging Rig time drilling, charging, blasting

43.1 No

min

21.0

min

10.6

Ventilation Ventilation break Loading, transport Normalised gross loading capacity

asm3/h

Factor of overbreak, excluding niches

267.5 1.150

Advance per round, pull

%

91.0

Rig time loading and hauling

min

17.8

Scaling Use of scaling jumbo Scaling time basic round length

Yes min

Correction for drilled length

51.8 1.00

Niches Volume of each niche

m3

85

Distance between niches

m

300

49

4 Simulation tool Advance rate input data

Unit

Model

User

Excavation time Hours per shift

h/shift

Number of shift per day

shift/day

2

Number of days per week

day/week

5

Working time per week

h/week

Number of weeks per year

week/year

Correction for jobtraining and tunnel length

10.1

101 44 1.00

Unforeseen time consumption

%

0.0

The calculations for net, standard and gross weekly advance rates (NTNU 2007b) and excavation time is given in Table 4.4. The method of calculation for each item is according to the advance rate model which is presented in NTNU 2007b and Chapter 3. As mentioned earlier, the user has possibility to change all the input, since most of the advance rate calculated items (Table 4.4) are input for the cost model, the calculated values also can be revised by the user. Some of the simulation tool results for advance rate have been used in the Chapters 3 and 4, like Figure 3.4, weekly advance rate and Figures 4.1- 4.4.

Table 4.4 Calculation for advance rate Advance rate calculation

Unit

Number of charged holes

Model 89

Penetration rate charged holes

cm/min

218.8

Penetration rate large holes

cm/min

93.1

Drilling time charged holes

min

67.8

Drilling time large holes

min

6.7

Time for moving

min

23.7

Time for rod adding

min

0.0

Time for bit changing

min

10.5

Time for lack of simultaneousness

min

6.5

Total drilling time

min

115.3

Charging time

min

43.1

Incidental lost time

min

19.9

Sum of drilling, charging, blasting time

min

199.3

Ventilation break

min

10.6

50

User

4 Simulation tool Advance rate calculation

Unit

Model

Actual volume per round

asm3

313.9

Loading time

min

70.4

Incidental lost time

min

9.8

Sum of loading and hauling time

min

98.0

Scaling time

min

51.8

Net round cycle

min

359.7

Net weekly advance rate

m/week

Extra time for niches

min

1.7

Correction for tunnel length and job-training

min

0.0

Standard round cycle

min

361.4

Standard weekly advance rate

m/week

Extra time for rock support

min

0.0

Unforeseen time consumption

min

0.0

Gross round cycle

min

361.4

Gross weekly advance rate

m/week

Excavation time

year

User

76.6

76.3

76.3 0.9

4.3 Cost model The simulation of the excavation cost is based on detailed cost calculation for all excavation operations as follows: • • • • • • • • • • •

Drilling Explosives and detonators Scaling Loading Contract transport for trackless tunnelling Track transport for track tunnelling Roadway/rails Tip Additional work (ventilation, electrical installation, water supply and misc.) Labour Niches 51

4 Simulation tool The general input data necessary for cost calculation is given in Table 4.5. The data forms the first sheet in cost calculation.

Table 4.5 General input for detailed cost calculations General cost input data

Unit

Investment tax

%

0.0

Interest rate

%

5.0

Electricity price

NOK/kWh

0.65

Fuel price

NOK/l

4.00

Adit slope

Model

Horizontal

Adit length

m

300

Distance to tip

m

300

Swelling factor for blasted rock Percentage of unforeseen costs Price level correction factor

User

1.65 %

10 1.0

Detailed cost calculation for each operation is carried out in separate Excel sheets. Each sheet is divided into three main parts, the input data, data from the advance rate/time model and a cost calculation part. In the two first parts, all necessary input is given. In the last part, the calculation of the cost items is performed. The unit cost in Norwegian kroner per metre of tunnel (NOK/m) is the result of each sheet, the sum of unit costs including 10 per cent extra for unforeseen, constitutes the total excavation cost. Since the cost model considers the gross weekly advance rate, time consumption for rock support and unforeseen will influence the excavation costs. Hence, one should be careful of adjusting the percentage of unforeseen costs proportional to unforeseen and rock support time consumption. Tables 4.6 and 4.7 present the input and cost calculation for drilling. The input and detailed cost calculation for the rest of the operations are given in Appendix C. In general, material and equipment prices and expected lifetime are main input factors for the cost calculation, through follow-up from different suppliers, contractors and data from previous versions of the cost model, the necessary data for different operation has been established and used in the new cost model. The equipment and material prices and

52

4 Simulation tool expected lifetime, including other data for cost calculation for different operations are presented in Appendix B.

Table 4.6 Drilling cost input data Drilling cost input data

Unit

Model

User

Drilling jumbo data Downtime factor

0.5

Percentage of life time used in tunnel

%

75.0

Time between tunnel sites

%

15.0

Jumbo price

NOK

Economic life time (in percussion hours)

h/hammer

Fixed repair costs

NOK/eh

300

Service costs

NOK/eh

150

Electricity consumption

kWh/eh

225

Drill bit price

NOK

600

Drill rod price

NOK

3800

Shank adapter price

NOK

1800

Coupling price

NOK

600

Drill bit lifetime

dm

437

Drill rod lifetime

dm

3500

Shank adapter lifetime

dm

4200

Coupling lifetime

dm

3500

Storing and freight costs

%

10.0

6,700,000 6,000

Drill steel data

Number of regrinds per bit Price for one regrinding

10 NOK

30.0

Time model data Penetration rate

cm/min

218.8

Advance per week

m/week

76.3

Number of weeks per year

week/year

44

Pull per round

%

91

Equivalent drillmetre per round

dm/round

480

Effective hours per round

eh/round

1.92

53

4 Simulation tool Table 4.7 Drilling cost calculation Drilling cost calculation

Unit

Value

Economic lifetime

dm/hammer

Economic useful lifetime of jumbo

dm

1,929,436

Time used in tunnel

dm

1,447,077

787,689

Depreciation factor

0.9375

Average rest value

0.4375

Number of rounds per week

round/week

16.8

Number of years for interest payment

year

4.70

Drillmetre per effective hour

dm/eh

Depreciation cost

NOK/dm

4.30

Interest cost

NOK/dm

0.50

Fixed repair cost

NOK/dm

1.20

Variable repair cost

NOK/dm

1.70

Downtime cost

NOK/dm

1.50

Electricity cost

NOK/dm

0.60

Service cost

NOK/dm

0.60

Jumbo cost

NOK/dm

10.40

Drill bit cost

NOK/dm

1.40

Bit regrinding cost

NOK/dm

0.70

Drill rod cost

NOK/dm

1.10

Shank adapter cost

NOK/dm

0.40

Coupling cost

NOK/dm

0.20

Storing and freight costs

NOK/dm

0.30

Drill steel cost

NOK/dm

4.05

Total drilling cost

NOK/dm

14.50

NOK/m

249.9

1526.00

Cost input data are stored in Excel in form of tables or in some few cases in form of equations. Excel functions like IF, LOOKUP and FORRECAST, or combinations of them have been used to allocate a proper entry to each input. The method of calculation of cost items is according to cost model presented in Chapter 3. The complete cost estimation presented in NTNU 2007c is based on the results of simulation tool. Since the Excel calculation is deterministic for a set of input, SENSIT (Sensitivity Analysis Add-in 2006) has been used to get the result as a function of

54

4 Simulation tool specific variables such as tunnel cross section area. The add-in tool facilitates application of the simulation tool for sensitivity analysis or risk analysis in time or cost.

4.4 Sensitivity analysis Sensitivity analysis can be performed at various levels, sensitivity analysis over all tunnel cross sections or over a specific tunnel with defined specifications. The purpose can be the overall excavation time and cost or a specific element of the excavation time or cost. This section is meant to assess the influence of rock properties on the excavation time and cost. Rock blastability and drillability are two main rock properties that are incorporated in the time and cost model. The number of holes and explosives consumption depends on the blastability. Hence, the blastability will influence the drilling and charging time and costs. The scaling time and cost are also dependent on the rock blastability. The drillability has a direct influence on the drilling time and cost. The influence of the rock blastability and drillability on the excavation time and cost are shown in Figures 4.1 - 4.4. The main input parameters are as follows: • • •

Tunnel length: 3 km Drill hole diameter: 48 mm Drill hole length: 5 m

Figures 4.1 and 4.2 show the influence of rock blastability. The impact of good and poor blastability on the excavation time and cost is 22 % and 15 % respectively. The time could vary 22 % and cost 15 %, the excavation time is more sensitive to the blastability than excavation cost. There is always a risk of assessing the wrong blastability, if good blastability is assumed when poor is correct the time and cost estimates will be too optimistic. The difference between good and poor drillability in Figures 4.3 and 4.4 is 10 % and 13 %, i.e. the rock drillability could fluctuate the excavation time and cost up to 10 % and 13 % respectively. The influence of good and poor blastability and drillability in percentage of time and cost is summarised in Table 4.8. The percentages are increase in time and cost from good to poor conditions.

55

4 Simulation tool Table 4.8 Influence of rock mass conditions on the excavation time and cost Rock conditions

Impact on time

Impact on cost

Blastability

22 %

15 %

Drillability

10 %

13 %

The sensitivity analysis of rock conditions indicates: •

The importance of the rock conditions and how the site investigation and assessment of the rock conditions could lead to a better estimation of tunnel project time and cost.

•

Relatively, the blastability is more important than drillability, since the blastability has more effect on the excavation time and cost.

•

The impact of the rock conditions on the time and cost terms provides the basis for economic estimation and risk evaluation.

The fluctuation of advance rate in Figures 4.1 and 4.3 is due to use of equipment with higher loading or drilling capacity as tunnel cross section area increases. Further details are given in Section 3.2.5.

56

4 Simulation tool 115 Good blastability m/week

105

Poor blastability Trendline

95 85 75 65 55 45 10

30

50

70

90

110 Cross section, m

130 2

Fig. 4.1 Influence of rock blastability on advance rate

20000 Poor blastability

NOK/m

18000

Good blastability Trendline

16000 14000 12000 10000 8000 6000 10

30

50

70

90

110 Cross section, m

130 2

Fig. 4.2 Influence of rock blastability on excavation cost

57

4 Simulation tool 115 Good drillability m/week

105

Poor drillability Trendline

95 85 75 65 55 45 10

30

50

70

90

110 Cross section, m

130 2

Fig. 4.3 Influence of rock drillability on advance rate

20000

NOK/m

Poor drillability 18000

Good drillability

16000

Trendline

14000 12000 10000 8000 6000 10

30

50

70

90

110

130 2

Cross section, m

Fig. 4.4 Influence of rock drillability on excavation cost

58

5 Conclusions and recommendations for further work 5.1 Conclusions Through the thesis, the only widely used drill and blast prediction model has been updated on the basis of the most recent technology, equipment and new field studies. The thesis consists of four volumes and one software program, and covers updated blast design, advance rate and cost models to be used for planning, design, time scheduling and cost estimation. The models are applicable through all phases of drill and blast tunnelling for small to large cross section areas. In addition, the Excel-based simulation tool (TunSim), a new engineering tool; facilitates the use of the prediction models as follows: • • • •

Fast and exact calculation of time and cost, as an alternative to the paper versions Possibility to vary all the input data Parametric studies and risk analyses on various input data related to excavation time or costs Built-in calculation for blast design and ventilation design

The whole thesis can be considered as a toolbox for tunnelling industry to be used by owners, consultants, contractors and suppliers. The models are practical and easy to use. The blast design model gives the design of cut holes, drilling pattern and explosives consumption. The final blast design output, i.e. number of holes or specific drilling and specific charging can be easily optimised by the guidance graphs.

59

5 Conclusions and recommendations for further work Comparison of the blast design model with a Swedish model shows that even with different methodology the results are close to each other. In general, the NTNU model gives higher number of holes and lower explosives consumption than the Swedish model (Zare and Bruland 2006a). By applying the advance rate and cost model for the following basic parameters, predicted advance rates and costs will be as given below: • • • •

Medium rock drillability and blastability 5 m round length and 48 mm drillhole diameter 100 hours working time per week 3 km tunnel length

The weekly advance rate without rock support installation varies from 100 m/week for 10 m2 tunnel to 54 m/week for 120 m2 tunnel, and the excavation costs without rock support varies from 6500 NOK/m for 10 m2 tunnel to 17500 NOK/m for 120 m2 tunnel, depending on equipment combination and number of drilling hammers. The costs are based on June 2005 price level (Zare and Bruland 2006b, 2007a). Considering the fact that the blast design and advance rate models are empirical and the results of such estimation models may not always exactly correspond to field performance data; analysing the field data against model prediction data shows acceptable correlation between the actual field data and the model. By means of the simulation tool, the impact of rock blastability and drillability is investigated. From good to poor blastability, the excavation time will increase by 22 % and the excavation costs will increase by 15 %. The drillability has relatively lower influence so that from good to poor drillability, the excavation time and cost will increase by 10 % and 13 % respectively. The investigation shows the importance of the rock conditions on the time scheduling and cost estimation. In line with developments in tunnelling, the productivity and efficiency of the drill and blast method have increased. The investigation shows a substantial increase in the productivity and reduction in the costs. For a 60 m2 road tunnel, the advance rate is increased from some 50 m/week to 81 m/week during the past 30 years, indicating a 60 % increase in production rate. And the excavation cost is decreased from some 16000 NOK/m to 10200 NOK/m, indicating a 36 % reduction from 1975 to 2005 (Zare and Bruland 2007a).

60

5 Conclusions and recommendations for further work In comparison to TBM, the drill and blast method is efficient and very cost effective in hard rocks and larger cross sections in various rocks (Zare and Bruland 2007b). The method has better flexibility to deal with the geological risks with substantially lower investment.

5.2 Recommendations for further work The current model is based on the 5 m round length and for longer rounds the model is corrected by correction factors. In some tunnels in Scandinavia the excavation of tunnels with 6 m round length has slowly started and maybe this will dominate in the future (Tunnelling and Trenchless Construction 2005). Hence, it is important to update the models based on 6 m round length when a sufficient amount of field data is available, especially for the blast design model. The amount of incidental lost time in the advance rate model is set to 10 % of the total time consumption of each operation. This percentage is obtained from the field studies in medium cross section tunnels. The incidental lost time for small and large tunnels should be investigated. The rock blastability index, or the SPR equation, is simplified for a specified degree of rock mass fracturing. Further research and investigation is required to incorporate the degree of fracturing in the rock blastabiliy index. This will enhance reliability and application of the index in various rock masses. As discussed previously, the blastability has a substantial impact on the excavation time and costs. The rock support time and cost default values in the simulation tool are set to zero. The user has the possibility to include calculated time and cost of the rock support – calculated from paper versions or own experience – into the simulation tool. In particular, it is important to include the time consumption of the rock support which has a direct impact on the excavation cost items. In further work, rock support items may be included to the simulation tool, but it is somehow difficult since the type and amount of the rock support depends a lot on the rock conditions and tunnel application. The results of blast design, advance rate and cost models are deterministic for a set of specified input. The simulation tool can be used for parametric studies or evaluation of risk/uncertainty in time and costs. The risk evaluation based on statistics and probability has also been used in tunnelling for time and cost analyses (Isaksson 2002, Haas and Einstein 2002).

61

5 Conclusions and recommendations for further work In addition, it is possible to further develop the simulation tool to be able to perform routine risk analysis. For this purpose; for at least some of the main input or all of them, a relevant probability distribution such as normal distribution is considered; then by using the Monte Carlo simulation method for example, the results will also be a probability distribution of time or costs.

62

References

Andersen, S.S., 2000: Production field study at Bakkatunnelen, Master’s thesis, NTNU, Trondheim. (in Norwegian) Atlas Copco, 2006: Drilling development, Available from < www.atlascopco.com>. Atlas Copco, 2004: Face Drilling, third edition, Atlas Copco Rock Drills AB, Sweden. Atlas Powder Company, 1987: Explosives and rock blasting, Dallas, TX, USA. Austlid, V., Andreassen, T.H.K., 1998: Production field study at Oslofjordtunnelen, Master’s thesis, NTNU, Trondheim. (in Norwegian) Barton, N., Grimstad, E., 1994: Rock mass conditions dictate choice between NMT and NATM, Tunnels & Tunnelling, October, pp 39-42. Blindheim, O.T., 2005: Clearing the air on long tunnel drives, Tunnels & Tunnelling International, Vol. 37 (3), pp 31-35. Clark, G.B., 1987: Principles of rock fragmentation, John Wiley & Sons, USA. Dyno Noble, 2005: Technical information; explosives, available from < www.dynonobel.com>. European Construction Technology Platform, 2005: Strategic research agenda for the European underground construction sector, available from . Flage, E., 2002: Production field study at Hagantunnelen, Master’s thesis, NTNU, Trondheim. (in Norwegian) Girmscheid, G., Moser, S., 2001: Fully automated shotcrete robot for rock support, in Computer-Aided Civil and Infrastructure Engineering, Vol. 16 (3), pp 200-215, Blackwell, Malden.

63

Girmscheid, G., Schexnayder, C., 2002: Drill and blast practices, Practice periodical on construction design and construction, Vol. 7 (3), pp 125-133. Haas, C., Einstein, H., 2002: Updating the Decision Aids for Tunnelling, Journal of Construction Engineering and Management, Vol. 128 (1), pp 40-48. Hermann, R., Elvoy, J., 2004: Automatic charging of emulsion explosives to increase safety, productivity and quality, In Health and Safety in Norwegian Tunnelling. Publication No 13, Norwegian Tunnelling Society. Isaksson, T., 2002: Model for estimation of time and cost based on risk evaluation applied on tunnel projects, PhD thesis, Division of Soil and Rock Mechanics, Royal Institute of Technology, Sweden. Johansen, J., Mathiesen, C.F., 2000: Modern trends in tunneling and blast design, Balkema, Rotterdam. Korsman, U., Nieminen, P., Salminen, P., 2006: New intelligent drilling jumbos for accurate, fast and cost-efficient tunnelling, 19th Canadian Tunnelling Conference, Vancouver. Kuczyk, G., 2002: Long blast round technology at the underground research laboratory, World Tunnelling, Vol. 15 (9), pp 432-434. Log, B., 1998: Production field study at Brønnøy kalk, Master’s thesis, NTNU, Trondheim. (in Norwegian) Lima, J., Blindheim, O.T., 2004: Development in ventilation methods, In Health and Safety in Norwegian Tunnelling. Publication No 13, Norwegian Tunnelling Society. Martin, P.D., Fitz, M.M., Friant, J.E., 1989: Development of an automated drill and blast system, RETC, pp 788-807. Mining & Construction, 2004: The development of the COP 3038, Mining & Construction No.3, pp 26-27. Neiminen, P., 2003: Blasting into the 21st century, Tunnels & Tunnelling International, Vol. 35 (7), 43-44. Nilsen, R., 1998: IT in tunnelling, Master’s thesis, NTNU, Trondheim. (in Norwegian) Nord, G., 2003: Use and misuse of the logging system, Tunnels & Tunnelling International, Vol. 35 (7), pp 46-47. Nordbotn, A.O., 2004: Production field study at OPS E39 Klett-Bårdshaug, Master’s thesis, NTNU, Trondheim. (in Norwegian)

64

Norwegian Public Roads Administration, 2004: Handbook 021; Road Tunnels, Directorate of public roads, pp 139. Norwegian Tunnelling Society, 2004: Norwegian Tunnelling, Publication No. 14, Oslo. NTNU, 2007a: 2A-05 Drill and Blast Tunnelling; Blast Design, Department of Civil and Transport Engineering, Trondheim, prepared by Shokrollah Zare, Vol. 2 of PhD thesis. NTNU, 2007b: 2B-05 Drill and Blast Tunnelling; Advance Rate, Department of Civil and Transport Engineering, Trondheim, prepared by Shokrollah Zare, Vol. 3 of PhD thesis. NTNU, 2007c: 2C-05 Drill and Blast Tunnelling; Costs, Department of Civil and Transport Engineering, Trondheim, prepared by Shokrollah Zare, Vol. 4 of PhD thesis. NTNU, 2006: Cost index of construction machinery, published monthly since 1978, Department of Civil and Transport Engineering, Trondheim. NTNU, 1998: Report 13A-98 Drillability; Test methods, Department of Civil and Transport Engineering, Trondheim. NTNU, 1995a: Report 2A-95 Tunnelling; Blast Design, Department of Civil and Transport Engineering, Trondheim. NTNU, 1995b: Report 2B-95 Tunnelling; Prognosis for Drill and Blast, Department of Civil and Transport Engineering, Trondheim. NTNU, 1995c: Report 2C-95 Tunnelling; Costs for Drill and Blast, Department of Civil and Transport Engineering, Trondheim. NTNU, 1992: Report 15A-92 Heavy Construction Machinery; Costs, Performance and Maintenance, Department of Civil and Transport Engineering, Trondheim (in Norwegian). NTNU, 1988: Tunnelling; Costs for Drill and Blast - background material, Department of Civil and Transport Engineering, Trondheim, (in Norwegian). NTNU, 1976: Report 3-76 Hydraulic Drilling in Tunnel, Department of Civil and Transport Engineering, Trondheim. (in Norwegian) NTNU, 1975a: Report 2-75 Tunnelling; Prognosis for Drill and Blast, Department of Civil and Transport Engineering, Trondheim. (in Norwegian) NTNU, 1975b: Report 3-75 Tunnelling; Costs for Drill and Blast, Department of Civil and Transport Engineering, Trondheim. (in Norwegian)

65

Olofsson, S. O., 2002: Applied explosives technology for construction and mining, Applex AB, Arla. Paulsen, M.Ø., 2005: Aukra onshore plant – Construction of subsurface space, Master thesis, NTNU, Trondheim. (in Norwegian) Rønn, P.E., Johannessen, O., 1995: Long rounds in tunnelling – the future, Norwegian rock blasting conference, Oslo. (in Norwegian) Sandvik Tamrock Corp., 1999: Rock excavation handbook for civil engineering. Scott, A., 1996: Blastability and blast design, proceeding of the fifth international symposium on rock fragmentation by blasting - Fragblast 5, Balkema, Rotterdam. Sensitivity Analysis Add-in for Excel, 2006: Available from < www.treeplan.com >. Singh, S.P., 2005: Blast damage control in jointed rock mass, Fragblast, Vol. 9 (3), pp 175-187. Singh, S.P., Xavier, P., 2005: Causes, impact and control of overbrek in underground excavations, Tunnelling and Underground Space Technology, Vol. 20, pp 63-71. Sørstrøm, M., 2005: Production field study at Sørdalstunnelen, Master thesis, NTNU, Trondheim. (in Norwegian) Taguchi, T., Sasaki, S., Ariki, T. and Kimura, Y., 2005: Developments and field tests of granular emulsion explosives, Science and Technology of Energetic Materials, 66 (5), pp 393-397. Thorkildsen, E., Vilhelmshaugen, L., 1996: Explosives, ventilation and grouting at Romeriksporten, Master’s thesis, NTNU, Trondheim. (in Norwegian) Tunnelling and Trenchless Construction, 2005: TTC Nordic Focus, Tunnelling and Trenchless Construction, November, pp 16-20. Zare, S., Bruland, A., 2006a: Comparison of tunnel blast design models, Journal of Tunnelling and Underground Space Technology, Vol. 21 (5), pp 533-541. Zare, S., Bruland, A., 2006b: Estimation model for advance rate in drill and blast tunnelling, Intern. symp. on utilization of underground space in urban areas, Egypt. Zare, S., Bruland, A., 2007a: Progress of drill and blast tunnelling efficiency with relation to excavation time and costs, World Tunnel Congress 2007, Prague. Zare, S., Bruland, A., 2007b: Assessment of TBM and D&B based on excavation time and costs, Submitted to 3rd Iranian rock mechanics conference, Tehran.

66

Published papers

- Comparison of tunnel blast design models Journal of Tunnelling and Underground Space Technology, Vol. 21 (5), pp 533-541. - Estimation model for advance rate in drill and blast tunnelling International symposium on utilization of underground space in urban areas, November 2006, Sharm El-Shikh, Egypt. - Progress of D&B tunnelling efficiency with relation to excavation time and costs Proceeding of the 33rd ITA World Tunnel Congress, May 2007, Prague. - Assessment of TBM and D&B based on excavation time and costs To be published in proceeding of 3rd Iranian Rock Mechanics Conference, October 2007, Tehran.

67

68

Tunnelling and Underground Space Technology incorporating Trenchless Technology Research

Tunnelling and Underground Space Technology 21 (2006) 533–541

www.elsevier.com/locate/tust

Comparison of tunnel blast design models Shokrollah Zare *, Amund Bruland Department of Civil and Transport Engineering, Norwegian University of Science and Technology, Trondheim, Norway Received 11 May 2005; received in revised form 2 September 2005; accepted 12 September 2005 Available online 2 November 2005

Abstract Blast design has direct influence on the time consumption and construction cost of drill and blast tunnels. Two tunnel blast design models based on parallel hole cut, NTNU blast design model and Swedish blast design model, are discussed and evaluated in this article. Both models suggest smooth blasting and in both models cut design depends on drill hole length and empty hole diameter. For lifter and stoping holes the Swedish model gives higher burden values, indicating a lower number of holes. Generally, The NTNU model gives longer uncharged length which indicates lower explosives consumption.  2005 Elsevier Ltd. All rights reserved. Keywords: Tunnel blast design; Parallel hole cut; Drill and blast tunnelling; Drilling pattern; Charging

1. Introduction Tunnel blasting is a much more complicated operation than bench blasting because the only free surface that initial breakage can take place toward is the tunnel face. Because of the high degree of constriction or fixation, larger charges will be required, leading to a considerably higher specific charge than in bench blasting (Persson et al., 2001). The basic principles for the method of charge calculation are those developed by Langefors and Kihlstrom (1978), first time published in 1963. The most important operation in the tunnel blasting procedure is to create an opening in the face in order to develop another free surface in the rock. This is the function of the cut holes. Cuts can be classified in two groups:  Parallel hole cuts.  Angle hole cuts.

*

Corresponding author. Tel.: +47 73 59 47 27; fax: +47 73 59 70 21. E-mail address: [email protected] (Sh. Zare).

The first group is most used in operations with mechanised drilling, whereas those of the second have fallen in disuse due to the difficulty in drilling (Jimeno et al., 1995). As to drilling, this has become more mechanised in the last decades, based on the development of hydraulic jumbos, with one or more booms, the trend has been toward parallel hole cuts as they are easier to drill, do not require a change in the feed angle and the advance is not as influenced by the width of the tunnel, as happens with angle cuts. In current drill and blast tunnelling, bulk explosives, i.e., ANFO and emulsion are widely used in the blasting operation and cartridged explosives are less used. The two blast design models to be investigated are the NTNU and Swedish models. The NTNU blast design model developed by the Department of Civil and Transport Engineering at NTNU (1975, 1995) is an empirical blast design model based on the parallel hole cut. The first version of the model was published in 1975. From the first publication, the model has been updated four times (1979, 1983, 1988 and 1995). The Swedish model is also based on the parallel hole cut. The Swedish model started with Langefors and Kihlstrom (1963) and has been further developed afterwards. Holmberg published the complete blast design model in 1982

0886-7798/$ - see front matter  2005 Elsevier Ltd. All rights reserved. doi:10.1016/j.tust.2005.09.001

69

534

Sh. Zare, A. Bruland / Tunnelling and Underground Space Technology 21 (2006) 533–541

(Holmberg, 1982) and recently updated by Persson et al. (2001).

2. NTNU model The tunnel blast design model is described in the Project Report 2A-95 (NTNU, 1995).The tunnel face divides into cut, stoping (easers), lifters (invert), row nearest contour and contour. Smooth blasting with double contour blasting is recommended, i.e., the charging density in the contour and row near the contour is reduced. The design for each part depends on the following rock and geometry parameters, which should be evaluated and determined in advance:    

Rock mass blastability. Drill hole diameter. Drill hole length. Skill level of the tunnel crew.

For each detonation interval, one has to control that the existing opening gives space for necessary expansion of the rock that will be blasted. This must be done by detailed calculations for the first two detonation intervals; visual evaluation is enough for the remaining holes. The cut is designed for an average pull of 90% of drillhole length for 45 mm drillhole diameter and 96% for 64 mm drillhole diameter. 2.2. Drilling pattern The drilling pattern for a specific tunnel depends on the following parameters:      

Drill hole diameter. Drill hole length. Rock mass blastability. Tunnel cross-section. Look-out angle. Skill level of the tunnel crew.

2.1. Cut design In the parallel hole cut used in the NTNU model, the blasting starts against an opening that is established by drilling one or more empty (large) holes. Three standard parallel hole cuts are shown in Fig. 1. The empty hole cut presupposes that the rock, which is blasted at each detonation interval, must have space for expansion (at least 80%) to secure full throw out. This requires precise drilling and correct firing sequence. The necessary area of empty holes is given in Fig. 2. The recommended distance between an empty hole and the first charged hole is shown in Table 1. When placing the other cut holes, the burden is set in relation to the basic width for the established opening (Fig. 3). The basic width is the width of the existing opening perpendicular to the direction of blasting, W1 or W2 in Fig. 3, basic width for hole number one or two. The recommended burden in Fig. 3 must be checked for enough expansion space, especially for hole number two in the cut, where Fig. 3 may give too high value for burden.

After the cut has been designed, the design of the drilling pattern should follow the sequence:    

Contour. Row nearest the contour. Lifters. Stoping.

For the contour, the row nearest the contour and the lifter holes there are guiding values for burden and spacing and there are also guiding stoping area values for stoping holes. Guiding values for burden, spacing and stoping area for 5 m drilled length are shown in Table 2. For the contour the values are given as intervals. The lowest values are for 20 m2 tunnels, the highest for 120 m2 tunnels. For other cross-sections, the values may be interpolated. The burden and spacing are given at the bottom of the round. At the face, eccentricity at the bottom of the holes (e.g., look-out angle) must be subtracted.

Fig. 1. Large hole cut for 45 mm drill holes, numbers indicates millisecond detonators interval (NTNU, 1995).

70

Sh. Zare, A. Bruland / Tunnelling and Underground Space Technology 21 (2006) 533–541

535

480

400

127mm

10 2mm

440

76 mm

Area, cm2

Empty hole diameter

3

360

8

320

7

4

280

64mm blasthole

6 2

3

240

45mm blasthole

5

200 4 2

160 3

1

120

Poor blastability

2

80

1

Good blastability

1

40 -2

-1

0

1

2

3

4

5

6

7

8

9

10

Drilled length, m Fig. 2. Necessary empty hole area for parallel hole cut (NTNU, 1995).

Table 1 Guiding distances between an empty hole and the first charged hole (NTNU, 1995)

dg

Diameter charged hole (mm)

Diameter empty hole, dg (mm)

Distance, a

45

76 102 127 76 102 127

1.5–2.0 Æ dg

64 a

For drilled length different from the basis (5 m), the values must be corrected by a correction factor for drilled length (Kbl in Figs. 4 and 5). For correction, the inverse of Kbl should be multiplied with the area (S · B). 2.3. Charging Necessary consumption of explosives for cartridged and bulk explosives are given in guidance graphs for planned tunnel cross-section. See Figs. 4, and 5 for ANFO. Tunnel rounds are usually charged with ANFO or emulsion, cartridged explosives may also be used; when water is a problem, ANFO can not be used efficiently. The contour is normally charged with special contour charges, e.g., tube charges or detonating cord. ANFO or emulsion may also

2.0–2.5 Æ dg

be used in the contour and the row nearest contour when mechanised charging systems are used. The charging density in the contour with double contour blasting is 20–25% and in the row nearest the contour it is 40–60% of normal charging density. The double contour refers to the contour and the row nearest the contour. Necessary charging depends on the following parameters:      

Drill hole diameter. Drill hole length. Rock mass blastability. Type of explosives (cartridge or ANFO/emulsion). Tunnel cross-section. Skill level of the tunnel crew.

71

536

Sh. Zare, A. Bruland / Tunnelling and Underground Space Technology 21 (2006) 533–541

Fig. 3. Guiding burden as function of basic width W of existing opening (NTNU, 1995).

Table 2 Guiding values for burden B, spacing S and stoping area Fs (NTNU, 1995) Type of hole

45 mm Drillhole

64 mm Drillhole

Burden (m)

Spacing (m)

Burden (m)

Spacing (m)

Contour

Good blastability Poor blastability

0.8–1.0 0.7–0.9

0.7–1.0 0.6–0.9

1.0–1.2 0.9–1.1

0.9–1.2 0.8–1.0

Row nearest contour

Good blastability Poor blastability

1.0 0.9

1.1 1.0

1.3 1.1

1.4 1.2

Lifters

Good blastability Poor blastability

1.0 0.8

1.0 0.8

1.3 1.1

1.3 1.1

Good blastability Poor blastability

Fs = 1.6 m2 Fs = 1.2 m2

Stoping (easers) Fs = S · B S/B = 1.2

In the NTNU experience ANFO and emulsion explosives have approximately the same charging density. Uncharged length (UL) is defined as a function of the drill hole length L, uncharged length for the cut and lifter holes is UL = 0.1L and for other holes UL = 0.3L. 2.4. Firing pattern Firing pattern must be planned so that each hole or group of holes, gets as favourable confinement and throw conditions as possible. That is ensured by trying to establish a smaller version of the final cross-section shape around the cut, and then enlarging this shape. It is also essential to check that the rock blasted at every interval number has space for expansion. The general sequence is cut, stoping, row nearest the contour, contour, lifter and finally corner holes of the lifter.

Fs = 2.6 m2 Fs = 1.8 m2

3. Swedish model This chapter is generally based on (Persson et al., 2001) and (Holmberg, 1982) where further details may be found. The tunnel face is divided into five separate sections as shown in Fig. 6. Cut, two stoping sections, contour and lifters. Each will be treated separately during calculations. Four-section cut type is used as a parallel hole cut. Design calculation depends on the following parameters:       

Length of drillhole. Diameter of drillhole. Linear charge concentration. Maximum burden. Type of explosive. Rock constant (Langefors and Kihlstrom, 1978). Fixation factor (Langefors and Kihlstrom, 1978).

72

Sh. Zare, A. Bruland / Tunnelling and Underground Space Technology 21 (2006) 533–541

537

Fig. 4. Necessary charging for ANFO in 45 mm drillholes and correction factor for drilled length (NTNU, 1995).

3.1. Cut holes The four-section cut is used as a dominant type of parallel hole cut (Fig. 7). Drillhole length depends on the empty hole diameter and there is direct relation between the drillhole length and the empty hole diameter as shown in Fig. 8. The resulting advance per round (pull) is assumed to be 95%. In the case with two empty holes in the cut instead of one, the equivalent diameter must be used in the calculations. The distance between the empty hole and the blastholes in the first quadrangle should not be more than 1.7 times the diameter of the empty hole to obtain breakage and a satisfactory movement of the rock. Breakage conditions differ very much depending upon the explosive type, structure of the rock and distance between the charge hole and the empty hole (Persson et al., 2001). As shown in Fig. 9, for burden larger than 2/, where / is empty hole diameter, the break angle is too small and a plastic deformation of the rock between the two holes is produced. Even if the burden is less than /, but the charge concentration is high, a sintering of the fragmented rock and cut failure will occur. For this reason, the recommended burden is B1 = 1.5/. For the first quadrangle the recommended burden should be checked by the empirical equation (Persson

et al., 2001, p. 221) and may be modified based on actual charge concentration or other parameters. In the equation, the burden depends on linear charge concentration, drillhole diameter, empty hole diameter, rock constant and type of explosive. The four holes in the first quadrangle are placed with the same distance from the empty hole (Fig. 7). To calculate the rest of the quadrangles (B2 to B4), it is considered that a rectangular opening already exists (Fig. 10) and linear charge concentrations are known. The burden will be calculated by the equation (Persson et al., 2001, p. 222) where burden depends on rectangular opening, linear charge concentration, drillhole diameter, rock constant and explosive type. For satisfactory breakage the calculated burden should fulfil two conditions:  B 6 2A to prevent plastic deformation.  B > 0.5A to reduce aperture angle to less than 90. The uncharged length of the cut holes is equal to 10 times of the drillhole diameter. 3.2. Lifters and stoping holes The burden for the lifters and stoping holes is in principle calculated with the same formula as for bench blasting.

73

538

Sh. Zare, A. Bruland / Tunnelling and Underground Space Technology 21 (2006) 533–541

Fig. 5. Necessary charging for ANFO in 64 mm drillholes and correction factor for drilled length (NTNU, 1995).

Fig. 6. Different tunnel sections (Persson et al., 2001).

The bench height is just exchanged for the advance, and a higher fixation factor is used due to the gravitational effect and to a greater time interval between the holes (Persson et al., 2001, p. 224). The burden depends on the linear charge concentration, fixation factor, rock constant and explosive type. A condition that must be fulfilled is B 6 0.6H where H is drillhole length. The same fixation factor (f = 1.45) is used for lifters and stoping holes in section B (breakage direction horizontally and upwards, Fig. 6). The fixation factor for stoping holes in section C (breakage direction downwards, Fig. 6) is reduced to f = 1.20.

Fig. 7. Four-section cut (Persson et al., 2001).

The spacing value of the lifter holes are equal to burden value (S/B = 1) and for both types of stoping holes the spacing is 1.25 times the burden values (S/B = 1.25). Like the cut holes, the uncharged length of the lifter and stoping holes is 10 times the drillhole diameter. The linear

74

Sh. Zare, A. Bruland / Tunnelling and Underground Space Technology 21 (2006) 533–541

539

since it is time-consuming charging work. Usually the same concentration is used both in the bottom and in the column (Persson et al., 2001). 3.3. Contour holes

Fig. 8. Hole depth as a function of empty hole diameter for four-section cut (Persson et al., 2001).

If smooth blasting were not to be used, the burden and spacing would be calculated according to stoping holes breaking downward. For smooth blasting, a spacing to burden ratio (S/B) of 0.8 should be used and the spacing between the contour holes is calculated from S = kd where the constant k is in the range of 15–16 and d is the drillhole diameter. The charge concentration is also a function of the drillhole diameter d. For hole diameter up to 150 mm, the following equation is used: Charge concentration ¼ 90d 2 ; where d is expressed in metres and the charge concentration in kg/m. In smooth blasting the total hole length must be charged to avoid the collar being left unbroken. 4. Comparison of the models 4.1. Cut design

Fig. 9. Blasting results for different relations between the burden and the empty hole diameter (Persson et al., 2001).

Fig. 10. Geometry for blasting toward a rectangular opening (Persson et al., 2001).

charge concentration in the column and the bottom charge (1.25B) may differ; the column charge can be reduced to 70% (of the bottom charge) for the lifter holes and 50% for the stoping holes. This is, however, not always common

In both models parallel hole cut with empty hole(s) is used. The necessary empty hole area in the NTNU model depends on drill hole length, drill hole diameter and rock blastability (Fig. 2). In the Swedish model, the diameter of the empty hole only depends on drill hole length (Fig. 8). In both models drillhole length has direct relation to the diameter of empty or large hole(s) in the cut. For 5 m drilled length the NTNU model gives 165–270 cm2 necessary area (depending on diameter and blastability) while the Swedish model gives 250 cm2. In the NTNU model the distance between an empty hole and the nearest charged hole (Table 1), depends on drill hole diameter and diameter of the empty hole. In the Swedish model this distance depends on the diameter of the empty hole (Fig. 9). NTNU model gives 1.5–2.5 times empty hole diameter depending on the diameter of the charged hole while the Swedish model gives 1.5 times the empty hole diameter. Comparison of necessary empty hole area and distance between empty hole and the nearest charged hole shows that the NTNU model more precisely determines these values. The two models give values in the same range. The design of the other cut holes in both models depends on dimension of existing opening, basic width in the NTNU model (Fig. 3) and opening width in the Swedish model (Fig. 10). Range of calculated burden in the Swedish model must be 0.5A < B 6 2A while according Fig. 3, NTNU model suggests 0.5A < B < 1.5A. So both models give more or less the same results. The NTNU model uses at least 80% allowable expansion to secure full throw out in each firing sequence. This

75

540

Sh. Zare, A. Bruland / Tunnelling and Underground Space Technology 21 (2006) 533–541

gives more possibilities for design of any type of parallel hole cut with one or more empty hole(s). 4.2. Drilling pattern In both models smooth blasting is recommended in the contour. The spacing to burden ratio for both models is presented in Table 3. With the same burden values, both models suggest equal or close to equal spacing values. The burden value in the NTNU model depends on blastability, drillhole diameter and length, and for the contour holes also tunnel cross-section. The burden values for 45 mm drill hole with 5 m length are given in the Table 4. In the Swedish model lifters and stoping holes are treated like bench blasting with higher fixation factor. The same formula is used to calculate the burden of the lifters and stoping holes with different S/B ratio and fixation factor. In smooth blasting the burden for the contour holes is calculated based on the spacing value. So the basis for contour calculation is different from the other holes. In the Swedish model the burden for the lifters and stoping holes depends on the linear charge concentration (drillhole diameter), fixation factor, rock constant, explosive type and S/B ratio. The burden values for 45 mm drillholes are given in the Table 5. The other assumptions for lifters and stoping holes are as follows:  Rock constant c = 0.4.  ANFO density = 900 kg/m3. Table 3 S/B relationship Hole type

Lifters Stoping Row nearest contour Contour

Model NTNU

Swedish

1 1.2 1.1 0.9

1 1.25 – 0.8

Table 4 Burden values in metres for the NTNU model (D = 45 mm, L = 5 m) Blastability Good

Poor

Lifters Stoping Row nearest contour Contour (average value)

1 1.15 1 0.8–1 (0.9)

0.8 1 0.9 0.7–0.9 (0.8)

Table 5 Burden values in metres for the Swedish model (D = 45 mm, c = 0.4)

Lifters Stoping, horizontally breaking Stoping, downward breaking Contour (average value)

The Swedish model recommends the contour burden in the range of medium blastability in the NTNU model. For the other holes the Swedish model gives considerably higher burden values especially when using ANFO as explosive. The following reasons illustrate the differences:  The different rock parameters in the models. In the Swedish model, rock constant, the amount of explosive needed for loosening one cubic meter of rock, under Swedish conditions c = 0.4 is predominant in blasting operations. In the NTNU model, rock mass blastability; depends on the rock sonic velocity, rock density, detonation velocity of explosive and charging density as well as rock mass fracturing (NTNU, 1995, pp. 13–15).  The bench blasting concept that is used in the Swedish model for the lifters and stoping holes calculation with different fixation factor may need some other correction or modification for use in tunnel blasting.  Indirect dependency of the burden value in the NTNU model to explosives type and charging density, is expressed only by rock mass blastability.  The different uncharged length of the drillholes in each model. The Swedish model does not take into consideration the drillhole length, while in the NTNU model when the drillhole length is different from the base (5 m) the burden must be corrected. This correction decreases the burden when the drillhole is longer than 5 m and increases the burden when the drillhole is less that 5 m. For example for 45 mm drillhole when the drillhole is 3 m correction increases the burden 3–4%. This correction has only minor effect on the general burden comparison. 4.3. Charging

Hole type

Hole type

 ANFO linear charge concentration = 1.43 kg/m.  Cartridge diameter = 32 mm.  Cartridge density = 1450 kg/m3.  Cartridge linear charge concentration = 1.17 kg/m.  Cartridge relative weight strength with respect to ANFO = 1.09.

Explosive ANFO

Cartridged explosive

1.3 1.2 1.3 0.8–0.9 (0.85)

1.25 1.1 1.23 0.8–0.9 (0.85)

As described in the Swedish model, in holes other than the contour, more emphasis is made on explosive type and charging density which are interdependent parameters with burden in the drilling pattern calculations. The NTNU model considers these parameters in the rock blastability evaluation (NTNU, 1995, pp. 13–15). In the Swedish model, except for the contour, where the total drillhole length is charged with less charging density, the uncharged length is equal to 10 times the drillhole diameter. For a 45-mm drillhole the uncharged length is 0.45 m. In the NTNU model the uncharged length depends on drillhole length. For cut and lifter holes 0.1L, for contour and stoping 0.3L. The uncharged length for 5 m drillhole is from 0.5 to 1.5 m. This indicates longer uncharged length

76

Sh. Zare, A. Bruland / Tunnelling and Underground Space Technology 21 (2006) 533–541

A comparison of the four-section cut design and total blast design result are summarized in Tables 6 and 7.

Table 6 Four-section cut burden and side length, m Quadrangle

First quadrangle burden B1 Second quadrangle burden B2 Third quadrangle burden B3 Fourth quadrangle burden B4 Fourth quadrangle side length

Model NTNU

Swedish

0.13 0.16 0.3 0.55 1.25

0.12 0.16 0.37 0.62 1.42

Table 7 Total blast design results Item

Total number of blast holes Specific drilling (dm/m3) Specific charging (kg/m3)

541

4.7. Cost estimation A cost comparison between two models is not within the scope of this paper. However, the fact that the Swedish model uses less drilling and more explosives then the NTNU model, indicates that the cost difference should not be significant between the two models. 5. Conclusions

Model NTNU

Swedish

45 2.5 1.7

40 2.2 1.9

and with the same charging density, the NTNU model gives lower explosives consumption in each hole. 4.4. Look-out angle In both models the burden and spacing are at the bottom of the hole or round and must be corrected at the face according the look-out angle or drilling deviation. 4.5. User-friendliness Since most data and values in the NTNU model are presented in graphs and tables, the model is easier for application for any tunnel cross-section, drillhole diameter and drillhole length. There are guidance graphs to check the final blast design outputs for any tunnel cross-section, i.e., number of holes, specific drilling and specific charging. 4.6. Comparison with an example An example of blast design with the Swedish model is presented for a tunnel with 19.5 m2 cross-section in (Holmberg, 1982), the main input are as follows:  Drillhole diameter = 45 mm.  Drillhole length = 3.2 m.  Empty hole diameter = 102 mm.  Explosive type = cartridged explosives.  Explosive density = 1200 kg/m3.  Rock constant = 0.4. For the NTNU model the same input is used and medium blastability is assumed. Although in the blastability definition the Swedish granite is evaluated as good blastability, considering all Swedish rock types ,the rock constant c = 0.4 is assumed be equivalent to medium blastability in this example for comparison.

Both models are developed by experience or empirical data. In both models parallel hole cut with empty hole(s) is used as cut type. In both models cut design depends on drillhole length which determines the empty hole diameter. Both models give approximately the same value for the empty hole diameter, distance between empty hole and the nearest charged hole and burden for other cut hole based on established opening. The NTNU allowable expansion enables possibility for design of any type of parallel hole cut with one or more empty holes. In both models, spacing to burden ratio for all holes are the same or close to each other. Also both models suggest smooth blasting with more or less the same burden values. In the Swedish model, lifters and stoping holes are treated like bench blasting with higher fixation factor. For these parts, the Swedish model gives higher burden values, especially when using ANFO as explosive. The reason mainly comes back to the methodology of each model. The higher burden values indicate the lower number of holes in the Swedish model. The recommended uncharged length of the drillhole in the two models is different, in the Swedish model it depends on the drillhole diameter and in the NTNU model it depends on drillhole length. Generally, the NTNU model gives longer uncharged length which indicates lower explosives consumption. References Holmberg, R., 1982. Charge Calculations for Tunnelling, Underground Mining Methods Handbook. Society of mining engineers, New york, pp. 1580–1589. Jimeno, C.L., Jimeno, E.L., Carcedo, F.J.A., 1995. Drilling and Blasting of Rocks. Balkema, Rotterdam. Langefors, U., Kihlstrom, B., 1978. The Modern Technique of Rock Blasting, third ed. Almqvist & Wiksell Forlag AB, Stockholm. NTNU, 1975. Project Report 2-75 TUNNELLING – Prognosis for Drill and Blast, NTNU. Department of Civil and Transport Engineering, Trondheim. NTNU, 1995. Project Report 2A-95 TUNNELLING – Blast Design, NTNU. Department of Civil and Transport Engineering, Trondheim. Persson, P.A., Holmberg, R., Lee, J., 2001. Rock Blasting and Explosives Engineering, sixth printing. CRC Press, USA.

77

Estimation Model for Advance Rate in Drill and Blast Tunnelling Shokrollah Zare, Amund Bruland Department of Civil and Transport Engineering, Norwegian University of Science and Technology (NTNU), Trondheim, Norway.

ABSTRACT This paper deals with the NTNU advance rate model for drill and blast tunnelling. The estimation model for advance rate is based on round cycle time consumption, and comprises drilling, charging, blasting, ventilation, loading and hauling, scaling and rock support. The basis of the model is State-ofthe-art technology and equipment with Norwegian tunnelling experience. Weekly advance rate as a function of tunnel cross section and equipment combination is presented, by applying the model for 5 m drillhole length, 48 mm drillhole diameter, parallel hole cut, medium blastability and drillability, medium rock wear quality and 101 working hours per week. Weekly advance rate without rock support installation varies from 100 m/week for 10 m2 tunnel to 54 m/week for 120 m2 tunnel, depending on equipment combination and number of drilling hammers.

1. INTRODUCTION Various applications of the drill and blast method in underground excavation necessitates updated design and planning models for blast design, time scheduling and cost estimations. The Department of Civil and Transport Engineering at NTNU has published models for such purposes since 1975 and has developed models for blast design, advance rate and excavation costs (NTNU, 1995a; NTNU, 1995b; NTNU, 1995c). The major part of around 5000 km of subsurface excavation in Norway has been excavated by the drill and blast method (Norwegian Tunnelling Society, 2004) and from these a substantial amount of field data have been used to develop the models. Numerous authors have emphasised influence of time on construction cost, e.g. Newitt (2005) and Grimstad (1999). The model is applicable for small to large tunnel cross sections with different excavation method, i.e. track tunnelling, trackless load and haul and trackless direct loading. Tunnel geometry, blast design parameters such as drillhole diameter and length and type of explosives, as well as rock properties, i.e. blastability, drillability and wear quality are considered in the model. The paper deals with basic parts and results of the model, the complete model is presented in (NTNU, 2006).

79

2. ESTIMATION MODEL FOR ADVANCE RATE The model is based on the round cycle time consumption. The round cycle is divided into four major operations: I II III IV

Drilling and charging Ventilation Loading and hauling Scaling and rock support

The operations I and III are divided into three different categories of time: A. Fixed lost time (rig time) All «unproductive» operations regularly repeated from round to round are collected here. The time consumption is also fixed in the sense that it is almost independent with regard to variations in round length, the number of crew members and the number of drilling hammers. Example: Driving the drilling jumbo to and from the face. B. Proportional operational time Proportional operational time is productive time, such as drilling and loading time. The time used is almost proportional to specific drilled metres and amount of broken rock, and inversely proportional to the number and performance of the drilling hammers or size and capacity of the loader. C. Incidental lost time The incidental lost time covers the technically dependent lost time occurring at random during tunnelling operations, for example machine breakdown. Personal time and delays connected to change of shifts are also included. A lost time of 6 minutes per hour is regarded as normal for well organized tunnelling. This constitutes 11.1% of A + B, i.e. 10% of total time consumption. 2.1 Drilling and charging The overall drilling time per round includes: o o o o o

Drilling of charged and empty/large holes Moving between holes Changing of bits Lack of simultaneousness Rod adding ( for drilled length longer than approximately 6 m )

A set of equations are presented (NTNU, 2006) to calculate time consumption of the above items. Number of holes, drillhole diameter, drillhole length, type and number of rock drills and Rock Drilling Index, DRI (NTNU, 1998) are main influencing parameters on drilling time. The charging time depends on: o Number of holes o Drillhole diameter and length o Explosives type and number of charging lines

80

For a 63 m2 tunnel with 92 charged holes of 48 mm diameter and 5 m length, when the holes are charged by two charging lines, the charging time varies from 60 to 85 min per round depending on explosives type. 2.2 Ventilation Ventilation break is a necessary break in the round cycle for diluting and removing of the blasting fumes. This is the time from blasting of the round until the concentration of nitrous gases (NOX) at the tunnel face is under the Threshold Limit Value, TLV = 2 ppm. Ventilation break varies from 5 to 30 min depending on tunnel cross section and explosives type. 2.3 Loading and hauling In general, loading and hauling time is dependent on the volume of blasted rock per round and loading capacity of loading and hauling equipment combination. The loading time is calculated by dividing the volume of blasted rock to loading capacity. Tunnel cross section, excavation method and combination of loading and hauling equipments are main influencing parameters on loading capacity. The normalized loading capacity for different equipment combinations is given based on field studies and normalization (NTNU, 2006). For a 63 m2 tunnel and 5 m round length, the loading and hauling time may vary from 100 to 150 min per round. 2.4 Scaling and rock support The scaling time covers time for scaling the round and checking of the rock face to allow further work. Time for rock support is not included in the scaling time. The scaling time depends on tunnel cross section, scaling method and rock blastability. For medium blastability (NTNU, 1995a) and 5 m round length the scaling time varies from 10 to 90 min depending on tunnel cross section. When using continuous rock support with bolts and/or shotcrete, it is common to include rock support time in the round cycle. By utilizing the time when there is no excavation, the shotcrete can be sprayed without being time-determinant. Time consumption for polyester anchored bolts as a function of number of bolts per round is given in (NTNU, 2006) both for bolt drilling time and bolt mounting time.

3. WEEKLY ADVANCE RATE For the weekly advance rate it is differentiated between net, normal and gross advance rate. The net advance rate is understood as the advance rate achieved for well organized tunnelling excluding time for blasting of niches, correction for job-training and tunnel length, rock support etc. The net round cycle time is the sum of the major operations (I- IV) excluding time consumption for rock support. The normal advance rate is determined based on net round cycle time with additional time consumption for blasting of necessary niches and correction for tunnel length and job-training effect. Correction factor for tunnel length and job-training effect is given in (NTNU, 2006). The gross advance rate is determined on the basis of the normal round cycle with additional time consumption for rock support and unforeseen, depending on site conditions. The normal weekly advance rate as a function of tunnel cross section for 3 km tunnel length is shown in Fig.1. The following assumptions are considered:

81

150 1 2 3 4 5 6 7 8

140 130 m/week 120

Haggloader - Shuttlecar, 2 drilling hammers Cat 972G - Load&haul, 2 drilling hammers Cat 980G - Load&haul, 3 drilling hammers Cat 973C - Truck, 3 drilling hammers Cat 966G/972G - Truck, 3 drilling hammers Cat 980G - Truck, 3 drilling hammers Volvo L330E - 35t dump truck, 3 drilling hammers Volvo L330E - 35t dump truck, 4 drilling hammers

110 1

100

3 5

2

90

4

6

7

80 8

70 60 50 40 0

10

20

30

40

50

60

70

80

90

100

110

120

Cross section, m

130 2

Fig.1 Normal weekly advance rate as a function of tunnel cross-section Rock conditions Rock blastability, rock drillability and rock wear quality influence the advance rate. The numbers of charged and empty holes are dependent to rock blastability. The drilling time is dependent to rock drillability and rock wear quality. Normal weekly advance rate is calculated for medium blastability, SPR = 0.47, medium drillability, DRI = 49, and medium rock wear quality, VHNR = 550. Blast design parameters The NTNU blast design model (NTNU, 1995a; Zare and Bruland, 2006) is used to calculate number of charged and empty holes for different tunnel cross sections. Parallel hole cut is assumed, the diameter of the empty hole(s) are 102 mm and the diameter of charged holes are 48 mm. The round length is 5m and the advance per round is assumed to be 91% of drilled length. Drilling and charging Nowadays, computerized drilling jumbos with different level of automation are widely used in drill and blast tunnelling (Atlas Copco, 2004). The drilling time is dependent on type and number of drilling hammers. 2 drilling hammers for cross section less than 20 m2, 3 drilling hammers for cross sections 20- 80 m2 and 4 drilling hammers for cross sections larger 80 m2 is assumed. The penetration rate is calculated based on the field performance data for COP 1838 rock drills.

82

Charging is assumed to take place after drilling. 2 charging lines for cross sections less than 80 m2 and 3 charging lines for cross sections greater than 80 m2, ANFO explosives for track tunneling and emulsion explosives for trackless tunnelling are assumed. Loading and hauling It is assumed that the numbers of hauling units are sufficient to fully utilize the loading equipment. Advance rate for the most efficient equipment combinations is presented. Working time per week In Norwegian tunnelling, the tunnel is normally excavated in 5 days per week, two shifts per day and 10 hours per shift. This results in an average of 101 working hours per week during a year.

4. APPLICATION FOR A TYPICAL ROAD TUNNEL A road tunnel with the following inputs is considered. o o o o

Tunnel geometry: 63 m2 cross section and 3 km length Blast design parameters: 92 charged holes of 48 mm and 3 large holes of 102 mm Rock support: 15 polyester anchored bolts per round Round length: 5 m

The normal and gross round cycle time are estimated to 371 and 431 min, which result in 74.3 m normal weekly advance rate and 64 m gross weekly advance rate. Distribution of the gross round cycle is shown in Fig.2.

Rock support 14% Drilling 33%

Scaling 12%

Loading,hauling 24%

V entilation 3%

Charging 14%

Fig.2 Distribution of the gross round cycle time for 63 m2 road tunnel The above results are subject to the assumptions specified in Section 3, in case of varying capacities and/or different conditions for each operation, the result may significantly vary. E.g., different drilling rate, varying working time per week, etc. Some results of such variations in advance rate are given in Tables 1 to 3.

83

Table1. The influence of drilling rate Item Drilling rate (m/min) Drilling time (min) Gross advance rate (m/week)

2.2 145 63.9

3 119 67.7

Ratio 1.36 0.8(1.22) 1.06

Table2. The influence of loading capacity Item Loading capacity (asm3/h) Loading time (min) Gross advance rate (m/week)

270 102 63.9

220 120 61.3

Ratio 0.8 1.18 0.96

144 88.6 9.6

Ratio 1.42 1.39 0.7(1.37)

Table3. The influence of working time Item Working time per week (hours) Gross advance rate (m/week) Excavation time(month)

101 63.9 13.2

REFERENCES Atlas Copco , 2004. Face Drilling, third edition, Atlas Copco Rock Drills AB, Sweden. Grimstad, E., 1999. Experiences from excavation under high rock stress in the 24.4 km long Laerdal tunnel, www.ngi.no/english/files/excavation_high_stress_-_eg_bangalo99.pdf. Newitt, J.S., 2005. Construction Scheduling: Principles and Practices, Pearson/Prentice Hall, U.S.A. Norwegian Tunnelling Society, 2004. Publication No. 14: Norwegian Tunnelling, Oslo. NTNU, 1995a. Report 2A-95 TUNNELLING Blast Design, NTNU, Department of Civil and Transport Engineering, Trondheim. NTNU, 1995b. Report 2B-95 TUNNELLING Prognosis for Drill and Blast, NTNU, Department of Civil and Transport Engineering, Trondheim. NTNU, 1995c. Report 2C-95 TUNNELLING Costs for Drill and Blast, NTNU, Department of Civil and Transport Engineering, Trondheim. NTNU, 1998. Report 13A-98 DRILLABILITY Test methods, NTNU, Department of Civil and Transport Engineering, Trondheim. NTNU, 2006. Report 2B-05 DRILL AND BLAST TUNNELLING Advance Rate, NTNU, Department of Civil and Transport Engineering, to be published in 2006. Zare, S., Bruland, A., 2006. Comparison of tunnel blast design models, Journal of Tunnelling and Underground Space Technology, Vol. 21/5, pp 533-541.

84

Progress of drill and blast tunnelling efficiency with relation to excavation time and costs S. Zare & A. Bruland

Department of Civil and Transport Engineering, Norwegian University of Science and Technology (NTNU), Trondheim, Norway

ABSTRACT: The productivity and efficiency of the drill and blast method have been investigated by means of NTNU prediction models for excavation time and costs since 1975, when the first time and cost models were published. The technological developments have improved the productivity in terms of increased advance rate and reduced excavation costs. For a typical 60 m2 road tunnel assuming 100 working hours per week, the normal advance rate without rock support installation is increased from 50 m/week to 80 m/week, i.e. 60 % improvement and the excavation costs are reduced from 16000 NOK/m to 10200 NOK/m, indicating a 36 % reduction from 1975 to 2005.

1 INTRODUCTION The drill and blast method has for long time been used for excavation of underground spaces in rock for various applications. The method may be applied in tunnels, rock caverns and mines. The major part of underground excavation in Norway is drill and blast (Norwegian Tunnelling Society 2004). The Department of Civil and Transport Engineering at NTNU has recently published updated models for blast design and excavation time and costs (NTNU 2006a,b,c). The models are developed based on field studies and technological progress of different elements involved in the tunnel excavation process. The models can be used through all phases of tunnel projects for: • Economic dimensioning • Choice of alternative • Time planning • Cost analysis, tender, budget and cost control • Choice of excavation method and equipment The first version of the models was published in 1975 and has later been updated and revised five times. From the first publication, substantial development has been achieved in drilling jumbos, drilling tools, explosives efficiency, loading and hauling equipment as well as HES regulations. The paper reviews the latest versions of the NTNU prediction models for excavation time and costs, as well as technological developments. Then

development of the productivity and cost efficiency of the method by means of empirical prediction models for excavation time and costs has been investigated since 1975. 2 TIME AND COST MODELS 2.1 Time model The complete time or advance rate model is published in NTNU 2006b. The model is based on the round cycle time consumption. The round cycle is divided into four major operations: I II III IV

Drilling and charging Ventilation Loading and hauling Scaling and rock support

The operations I and III are divided into three different categories of time, i.e. fixed lost time (rig time), proportional operational time and incidental lost time. The time consumption for each operation is given in NTNU 2006b for various tunnel cross sections. By applying the model for 3 km tunnel length, 5 m drillhole length, 48 mm drillhole diameter, parallel hole cut, medium rock drillability and blastability and 100 working hours per week, the weekly advance rate without rock support installation varies from 100 m/week for 10 m2 tunnel to 54 m/week for 120 m2 tunnel, depending on equipment combination 85

and number of drilling hammers (Zare & Bruland 2006b). 2.2 Cost model The cost model is based on detailed cost calculation for excavation operations, i.e. drilling, charging, scaling, loading, hauling, additional work, niches and labour. The complete cost model is published in NTNU 2006c. Equipment and material prices, labour wages and expected life time for equipment are main input for detailed cost calculation. Depreciation, interest, wages, repairs and power consumption are included in the costs of each operation. Linear or degressive depreciation method is assumed, depending on equipment type and repair costs over life time. In addition, 10 % extra for unforeseen costs and uncertain assumptions is added in the end to the excavation costs. As in the time model, rock support cost is not included in the excavation costs, in order to provide the basis to compare the costs for different tunnel cross sections. The amount of rock support much depends to the rock conditions and tunnel application. The equipment combinations are chosen according to the tunnel cross section and excavation method, i.e. track tunnelling for cross sections up to 16 m2 and trackless tunnelling for cross sections between 16 m2 and 120 m2. For rough estimation, the excavation costs for a tunnel with the same assumptions in the time model, varies from 6500 NOK/m for 10 m2 tunnel to 17500 NOK/m for 120 m2 tunnel, depending on equipment combination and number of drilling hammers. The costs input are based on June 2005 price level. 3 TECHNOLOGICAL DEVELOPMENT 3.1 Drilling Currently, computerised hydraulic drilling jumbos with different level of automation are widely used in drill and blast tunnelling (Atlas Copco 2004, Sandvik Tamrock Corp. 1999). The current generation of drilling jumbos is designed for high productivity, quality drilling, and comfortable working conditions for operators. The drill plan is stored in the drilling jumbo computer and therefore the face does not need to be marked up; the navigation is done by laser alignment and is very precise in deciding the position of the drilling jumbo. The excavated tunnel profiles can be scanned by electronic instruments, whereby overbreak and under-break can be recorded to be used for the optimisation of the drilling process in the coming rounds (Nord 2003).

3.1.1 Rock drills Obviously, invention of the hydraulic rock drills in the 1970s was a great progress in drilling speed, economy and improvement of the operator’s working conditions. Figure 1 supports the fact and shows general development of rock drills and jumbos.

Figure 1. Drilling development 1905 – 2005, drilled metres per hour and operator (Atlas Copco 2006).

As an example, one may look at the Cop 3038. The latest rock drill from Atlas Copco is designed to increase the penetration rate in hard rock tunnelling. According to Atlas Copco; the Cop 3038 is 50 % faster than its predecessor Cop 1838 ME, the time saved with Cop 3038 compared to Cop1838 ME on 3-boom jumbos in a 90 m2 tunnel could be about 40 minutes per round. The high penetration rate may be utilized in limited space to replace 3-boom to 2boom jumbos (Mining & Construction 2004). The Cop 3038 delivers the same energy per percussive blow as the Cop 1838 ME, but the frequency has been increased from 60 to 102 Hz. In the Table 1 the development of Atlas Copco hydraulic rock drills for face drilling is presented. Table1. Development of Atlas Copco hydraulic face drilling (Mining & Construction 2004). Type Year Impact power COP 1038HD 1973 12 kW COP 1238ME 1983 12 kW COP 1440 1986 21 kW COP 1838ME 1990 18 kW COP 1838HF 2000 22 kW COP 3038 2004 30 kW

rock drills for Impact rate 42 – 60 Hz 40 – 60 Hz 70 Hz 60 Hz 73 Hz 102 Hz

86

3.1.2 Drill steel Drill bits, drill rods and shank adapters may be referred to as the drill steel, which has been concurrently developed with introduction of more powerful hydraulic rock drills. New conical thread system, Magnum SR35, is devised to cope the weakness of the old threads. The advantages of the new thread system are intended to solve rod breakage at the bit end, less tendency for deviation when collaring, straighter holes, and longer service life (Atlas Copco 2004). The rod length has also increased, even though the 18 feet rods are common in civil tunnelling and gives around 5 m drilled length, the tendency towards longer round is increased (Nieminen 2003) and some tunnels have been excavated with 6 m round length (Tunnelling and Trenchless Construction 2005). 3.2 Charging and explosives Emulsion explosives combined with non electrical initiation system like NONEL have increased the safety of the charging and blasting operations and become more efficient. The pumpable emulsion explosives, SSE (Site Sensitised Explosives) are not explosives until pumped in the hole, which means safety in transportation and handling (Johansen & Mathiesen 2000, Atlas Copco 2006). The modern emulsion explosives are oxygenbalanced, producing a minimum of noxious fumes and far less smoke (Olofsson 2002), provides better working conditions and less ventilation requirements. The emulsion explosives are water resistant and can be used when water is a problem and ANFO can not be used efficiently. In addition, the pumpable emulsion explosives can be pumped into the blastholes by a computer controlled system and the amount of explosives in each hole and total explosives consumption can be measured, which can be used to control the amount of explosives in different holes in the face, especially in the contour and row nearest the contour where less explosives is needed for smooth blasting (NTNU 2006a, Zare & Bruland 2006a). The measuring system may also help to control the uncharged length of blastholes for optimum blasting results. Recently, after a feasibility study, a prototype machine for automatic charging of emulsion explosives was completed and tested in Norway (Hermann & Elvoy 2004). The project is meant to reduce the time consumption for drilling and charging by 20 % with improved work safety. The charging system is mounted on the drilling jumbo as a separate boom and charging is accomplished automatic during drilling without human interface at the face, in compliance with the regulations that prohib-

its simultaneous manual charging and drilling at the face. 3.3 Ventilation The harmful gases and dust particles above the certain level of acceptance must be removed by the ventilation system. For the cost efficiency and improvement of the work environment, the intelligent ventilation system is developed and tested in two tunnels (Lima & Blindheim 2004). The project includes development of new duct materials and duct support system as well as recording of air quality and automatic control of the ventilation fans. 3.4 Loading and hauling The loading and hauling capacity is dependent on the tunnel cross section size and excavation method. Using adequate size equipment as the cross section increases is an efficient way to increase loading capacity. By NTNU field studies in large cross sections (NTNU 2006b), a capacity up to 500 lm3/h is attainable when using wheel loader and dump truck combination and the loader is fully utilised, i.e. the loader is not waiting for a hauling unit. Crusher, conveyor belt and mucking trains are also employed in drill and blast tunnels. The average capacity is 300 m3/h (understood as loos cubic metre). 500 m3/h should be achievable by the method in the future (Girmscheid & Schexnayder 2002). The crusher and conveyor belt is used as part of “high performance drill and blast excavation concept” which uses a suspended platform system for higher performance. Even though, the crusher capacity is the limiting factor, by other benefits of the method, the total cycle performance has been increased by 30 % (Girmscheid & Schexnayder 2002). 3.5 Rock support Rock bolts and sprayed concrete (shotcrete) are two common types of rock support, both developed towards more durable; high strength, and quick installation by means of mechanised and automated equipment. CT-bolt, the newest bolt type in the market can first be used as an ordinary end-anchored bolt for temporary support and later on be grouted to a permanent bolt (Norwegian Tunnelling Society 2004). During the past two decades the sprayed concrete has faced a significant improvement; today fibrereinforced sprayed concrete by wet mix method is widely applied. Sprayed concrete robot is one of the modern equipment for applying sprayed concrete. The capacity by manual spraying is less than 5 to 8 m3/h, while using robot it reaches up to 20 m3/h (Girm87

4 TIME AND COST TRENDS To investigate time and cost trends or productivity and efficiency of the method, a 60 m2 tunnel with 3 km length is chosen as a typical two-way road tunnel. Medium rock blastability and rock drillability is assumed. The advance rate per week and unit costs per tunnel metre without rock support installation are estimated from 1975 to 2005 by the NTNU models. The NTNU model was first time published in 1975 and has been revised five times; in 1979, 1983, 1988, 1995 and 2005. Weekly advance rate is shown in Figure 2. The weekly advance rate is normalised based on 100 hours working time per week. The advance rate is increased from some 50 m/week to 81 m/week during the past 30 years, indicating a 60 % increase in production rate. The excavation method itself does not change significantly, development in the equipment capacity and efficiency, tools and materials leads to that the current tunnel will be excavated faster and the excavation time is shortened. For example in 1975 the maximum penetration rate for medium drillability was 140 cm/min whilst in the recent model of 2005, it reaches to some 300 cm/min and the maximum loading capacity for same size cross sections is increased from 160 sm3/h to 280 sm3/h.

Kroner per metre of tunnel. The cost in Figure 3 is not corrected for price level; only representing the unit excavation costs in each year. To find the comparable figures, the cost must be corrected for price level. The Department of Civil and Transport Engineering at NTNU has published a cost index for construction equipment (NTNU 2006d). The corrected excavation costs are shown in Figure 4. The trend of excavation cost is not consistent such as the time trend (Fig. 2). One reason is different working hours per week; which is 112 hours in 1975, 75 hours in 1979 and 1983 models, and 100 hours for the rest. Based on 100 working hours the costs for 1975 model should be higher and for the 1979 and 1983 models, lower than the presented values. In spite of this minor deviation the excavation cost is decreased from some 16000 NOK/m to 10200 NOK/m, indicating a 36 % reduction. The technological development is the main reason of decreasing cost trend. Changes in the investment tax and interest rate are included in the price level correction.

12000 10000 NOK / m

scheid & Moser 2001). The robot has better safety and quality control. By the NTNU models it possible to estimate excavation time and cost including rock support, but since the rock support much depend on the rock mass conditions, the time and costs which are presented in the paper do not include the rock support.

8000 6000 4000 2000 0 1975

1979

1983

1988

1995

2005 Year

Figure 3. Development of uncorrected excavation cost for a 60 m2 tunnel. 90 80 18000 16000

60

14000

50

NOK / m

m / week

70

40 30

12000 10000

20

8000

10

6000 4000

0 1975

1979

1983

1988

1995

2005 Year

2000 0

2

Figure 2. Development of weekly advance rate for a 60 m tunnel.

The uncorrected excavation costs are shown in Figure 3, the unit cost is presented in Norwegian

1975

1979

1983

1988

1995

2005 Year

Figure 4. Development of excavation costs for a 60 m2 tunnel, price level June 2005.

88

5 CONCLUSIONS In line with development in tunnelling, the productivity and efficiency of the drill and blast method has increased. Drilling is now using computerised drilling jumbos with various levels of automation, faster rock drills and more durable drill steel. Charging is accomplished safer and more efficient by using NONEL detonators and emulsion explosives which causes less pollution and is not considered as an explosive until placed in the holes. Ventilation may utilise intelligent ventilation system. The loading and hauling capacities are increased and advanced rock support equipment with higher performance is employed. By means of the NTNU prediction models, the investigation shows substantial increase in the productivity and reduction in the costs. For a 60 m2 road tunnel, the productivity has increased 60 % and the cost has been reduced by 36 % from 1975 to 2005. Furthermore, the time and cost trends confirm the importance of updated models for time and cost prediction. The NTNU models are unique and should be revised based on technological development in the excavation methods and equipment.

NTNU. 2006a. Report 2A-05 Drill and Blast Tunnelling; Blast Design. Trondheim: NTNU, Department of Civil and Transport Engineering, to be printed in 2007. NTNU. 2006b. Report 2B-05 Drill and Blast Tunnelling; Advance Rate. Trondheim: NTNU, Department of Civil and Transport Engineering, to be printed in 2007. NTNU. 2006c. Report 2C-05 Drill and Blast Tunnelling; Costs. Trondheim: NTNU, Department of Civil and Transport Engineering, to be prined in 2007. NTNU. 2006d. Cost index of construction machinery, published monthly since 1978. Trondheim: NTNU, Department of Civil and Transport Engineering. Olofsson, S. O. 2002. Applied explosives technology for construction and mining. Arla: Applex AB. Sandvik Tamrock Corp. 1999. Rock excavation handbook for civil engineering. Tunnelling and Trenchless Construction. 2005. TTC Nordic Focus. Tunnelling and Trenchless Construction. November 2005:16-20. Zare, S. & Bruland, A. 2006a. Comparison of tunnel blast design models. Journal of Tunnelling and Underground Space Technology 21(5): 533-541. Zare, S. & Bruland, A. 2006b. Estimation model for advance rate in drill and blast tunnelling. Intern. symp. on utilization of underground space in urban areas, 6-7 November 2006. Sharm El-Sheikh, Egypt. .

ACKNOWLEDGEMENTS The paper is prepared as part of the first author’s PhD research project financed by the Ministry of Science, Research and Technology of Iran. REFERENCES Atlas Copco. 2006. Available from < www.atlascopco.com>. Atlas Copco. 2004. Face Drilling third edition. Orebro: Atlas Copco Rock Drills AB. Girmscheid, G. & Moser, S. 2001. Fully automated shotcrete robot for rock support. in Computer-Aided Civil and Infrastructure Engineering 16: 200-215. Blackwell: Malden. Girmscheid, G. & Schexnayder, C. 2002. Drill and blast practices. Practice periodical on construction design and construction 7(3): 125-133. Hermann, R. & Elvoy, J. 2004. Automatic charging of emulsion explosives to increase safety, productivity and quality. In Health and safety in Norwegian Tunnelling. Publication No 13. Oslo: Norwegian Tunnelling Society. Johansen, J. & Mathiesen, C.F. 2000. Modern trends in tunneling and blast design. Rotterdam: Balkema. Lima, J. & Blindheim, O.T. 2004. Development in ventilation methods. In Health and safety in Norwegian Tunnelling. Publication No 13. Oslo: Norwegian Tunnelling Society. Mining & Construction. 2004. The development of the COP 3038. Mining & Construction No.3: 26-27. Neiminen, P. 2003. Blasting into the 21st century. Tunnels & Tunnelling International 35(7): 43-44. Nord, G. 2003. Use and misuse of the logging system. Tunnels & Tunnelling International 35(7): 46-47. Norwegian Tunnelling Society. 2004. Norwegian Tunnelling. Publication No. 14. Oslo.

89

Assessment of TBM and D&B based on excavation time and costs Shokrollah Zare*, Amund Bruland Department of Civil and Transport Engineering, Norwegian University of Science and Technology, NTNU, Trondheim, Norway.

Abstract By means of the latest NTNU prediction models for time and costs, the productivity and efficiency of the TBM and D&B options have been investigated. The weekly advance rate and unit excavation costs as a function of rock conditions and tunnel geometry are presented. As a result, the efficient and cost effective method for a tunnel is subject to mainly rock conditions, tunnel cross section area, length and tunnel application. Under Norwegian costing, for hard rock (poor boreability), the D&B method is cost effective whereas in small and long tunnels with good boreability the cost associated to the TBM is lower. Generally, the TBM has higher advance rate, and the rock conditions has a substantial impact on the TBM performance and costs. Key words: TBM tunnelling, drill and blast tunnelling, tunnel excavation time and costs

1- Introduction Selection of the excavation method is an important issue that may arise during planning of a tunnel project, depending chiefly on tunnel geometry, rock conditions, labour costs and local regulations. Both the TBM and the drill and blast methods have pros and cons related to excavation time and costs, safety and environment aspects, and risk evaluation, which all should be pointed out and analysed prior to the construction phase. The Department of Civil and Transport Engineering at NTNU has been involved both in D&B and TBM tunnelling since 1975 and has published models for design, advance rate and cost estimations. Norway has a mountainous topography with dominating hard rock, so far more than 5000 km of tunnels and caverns has been excavated, mostly for hydropower projects, out of this, around 260 km tunnel have been bored by TBM [1]. The experience and use of the latest tunnelling technology provides the basis of the model development. The paper reviews the latest NTNU models and outlines influencing rock parameters used in the models. By applying the model, the efficiency of the two methods is assessed for small and large cross sections. The assessment will focus on the excavation and construction time and costs. The influence of the various rock conditions and tunnel geometry are investigated.

*

Corresponding author. Tel.: +47 73 59 47 27; fax: +47 73 59 70 21 E-mail address: [email protected]

91

2- NTNU models The Department of Civil and Transport Engineering at NTNU has published models for TBM and drill and blast tunnelling since 1975. The models have later been further developed and updated based on new field studies and technological development five times since the first edition. The latest edition of the TBM model was published in 1998, consisting of six volumes [2,3,4,5,6,7] covering design and construction, models for advance rate and costs, geology, pre-investigation, back-mapping and the boring process. In the model rock mass boreability and machine parameters play a major role in the TBM performance and efficiency. The net penetration rate, cutter life and ultimately excavation time and cost depend heavily on the rock mass and machine parameters. The model for drill and blast tunnelling consists of three volumes including blast design, advance rate and costs. The 5th edition of the model was published in 1996, the latest edition to be published in 2007 [8,9,10,11]. The model provides a practical tool for design, time scheduling and costs. The models are applicable for small to large cross sections (10 - 120 m2), with different excavation methods. The rock mass conditions and equipment efficiency are also considered in the model. By the technological development the productivity and efficiency of the D&B method is substantially increased [12]. The models can be used through all phases of tunnel projects for: • Economic dimensioning • Choice of alternative • Time planning • Cost analysis, tender, budget and cost control • Choice of excavation method and equipment 3- Rock conditions 3-1- TBM model

The most important rock mass and rock type parameters are the degree and orientation of rock mass fracturing, drillability (Drilling Rate Index DRI) and abrasiveness (Cutter Life Index CLI). The systematic rock mass fracturing (joints and fissures) are classified into seven different classes based on the average spacing between weakness planes varying from 5 to 160 cm, the homogenous rock with no fracture is classified with indefinite spacing. The angle between the tunnel axis and the planes of weakness may vary from 0 - 90 °, the angle about 60 ° is found to be more favourable for penetration rate. The estimated penetration rate is increased by a factor of five from homogenous to well fractured rock mass [13]. The drillability and abrasiveness is determined through three laboratory tests [14, 15]. The tests yield the Brittleness Value S20, Sievers J-value SJ and Abrasion Value Cutter Steel AVS. The S20 is a measure for rock resistance to impact or energy required to crush the rock. SJ is a measure for rock resistance to indentation or surface hardness and AVS is a measure of abrasion of the cutter ring. The results of S20 and SJ values determine the DRI value and SJ and AVS determine the CLI. The drillability, DRI is used to estimate net penetration rate. Abrasiveness or CLI is used to estimate cutter life. For homogeneous or non-fractured rock mass, the estimated

92

penetration rate increases by a factor of two from extremely low (DRI = 25) to extremely high (DRI = 82) values [13]. A set of rock conditions is chosen to estimate excavation time and costs of bored tunnels as summarised in Table 1. The boreability refers to both rock type and rock mass properties involved in the boring process. As an example schist and phyllite may be considered as good boreability, limestone as a medium and granite and gneiss as a poor boreability. Table 1 Rock conditions for TBM Boreability Fracture class, fissures Average spacing, cm Fracture class no. Angle, degrees Drilling Rate Index, DRI Cutter Life Index, CLI Quartz contents, %

Good St III 10 3 20 65 25 25

Medium St II - І 20 – 40 1.5 20 49 10 - 12 25

Poor St І - 0 40 - ∞ 0.5 20 35 6-7 25

3-2- Drill and blast model

Rock type blastability and drillability are two main properties in the drill and blast model. The rock blastability is given by the rock blastability index, SPR [8] which is “the amount of explosives (kg/m3) needed to break the rock to a certain degree of fragmentation, where 50 % of the blasted rock size is under 250 mm (d50 = 250 mm)”. The blastability index can be determined by an equation that depends on rock and explosives properties such as anisotropy of rock, density, sonic velocity of rock, charging density and detonation velocity. The index is meant to aid the evaluation of blastability and assumes access to laboratory data from a representative sample of the particular rock. In the model, a classification is used to distinguish between various rocks, i.e. good, medium and poor blastability; referring to SPR = 0.38, 0.47 and 0.56 respectively. The number of holes and explosives consumption depends on the blastability. Hence, the blastability will influence the drilling and charging time and costs. The penetration rate of the drilling hammers directly depends on the DRI, therefore the DRI influences the drilling time and costs. Expected lifetime of the drill bits is estimated using Vickers Hardness Number Rock VRNR [14, 9], which is a measure of the abrasiveness, equivalent to the quartz contents of rock type. The assumptions of the good/poor rock conditions for estimation of the excavation time and costs are presented in the Table 2. Table 2 Rock conditions for drill and blast Rock conditions Rock mass blastability Blastability index, SPR Drilling Rate Index, DRI Rock wear quality, VHNR

Good good 0.38 65 250

Medium medium 0.47 49 550

Poor poor 0.56 35 850

93

4- Excavation time and costs Generally, the excavation time or cost is a function of excavation method, rock conditions, tunnel cross section area and length, and equipment capacity. In Figure 1, the weekly advance rate based on 101 working hours per week is presented for tunnels with 3 km length. Since the rock support much depends on the rock mass conditions and tunnel application, rock support is not included in the weekly advance rate and costs shown here. The input data are for 3.5, 6 and 9 m TBM or 9.6, 28.3 and 63.6 m2 cross section areas representing tunnels from a typical water tunnel to a typical two-way road tunnel. The figure demonstrates the importance of the rock conditions especially for TBM excavation. From poor to good boreability, the weekly advance rate increases 3 – 4 times for a TBM, while for drill and blast, the corresponding increase is 20 – 40 % when the rock conditions is varied. 300

m/week

TBM-good boreability TBM-poor boreability

250

D&B-good rock conditions 200

D&B-poor rock conditions

150 100 50 0 0

10

20

30

40

50

60

70

Cross section, m2

Figure 1 Weekly advance rate, tunnel length 3 km

In the good boreability, the TBM has notably higher advance rate whereas in the poor boreability the weekly advance rate of the TBM drops even below that of the drill and blast. The weekly advance rate curves for TBM as a function of tunnel cross section are degressive, the D&B curves are more linear. Indicating when the cross section increases; by reduction in RPM (revolution per minutes) the net penetration of the TBM is dramatically reduced while in D&B due to possibility of using larger equipment with higher capacity the advance rate curves are more or less linear. The unit excavation costs in NOK/m are shown in Figures 2 and 3 for 3 km and 6 km tunnel length. The curves are based on the detailed excavation costs (excluding rock support) with 10 % extra for unforeseen and uncertain assumptions. The cost curves are the reflection of the weekly advance rate on the cost so that lower advance rate gives a higher cost and vice versa. From 3 to 6 km tunnel the unit cost of D&B is increased by 6 – 9 % (medium rock conditions) because of mainly transport and ventilation cost, but the TBM cost is reduced 17 – 6 % (medium boreability) since the 6 km is almost optimum length and lower set up costs.

94

When the cross section increases, the TBM cost is increased more rapidly than the D&B, i.e. the marginal cost (the extra cost when the tunnel cross section increases 1 m2) of TBM is higher than the D&B. The main reasons are reduction of the advance rate and increase of the equipment price. The marginal cost for TBM (medium boreability) and D&B (medium rock conditions) is estimated at 310 and 95 NOK/m.m2 respectively (tunnel length 6 km), which indicates the marginal cost of TBM is more than 3 times that of the D&B. The costs curves also show that the excavation cost of the D&B is lower than for TBM, especially in case of poor boreability or larger cross sections. The TBM has advantages in the smaller cross sections with favourable rock conditions. When the tunnel gets longer the cost benefit of the TBM is increased. 60000

NOK/m

TBM-good boreability TBM-poor boreability

50000

D&B-good rock conditions D&B-poor rock conditions

40000 30000 20000 10000 0 0

10

20

30

40

50

60 Cross section, m

70 2

Figure 2 Unit excavation costs, tunnel length 3 km (price level June 2005)

60000

NOK/m

TBM-good boreability TBM-poor boreability

50000

D&B-good rock conditions D&B-poor rock conditions

40000 30000 20000 10000 0 0

10

20

30

40

50

60 Cross section, m

70 2

Figure 3 Unit excavation costs, tunnel length 6 km (price level June 2005)

95

Furthermore, the TBM costs are very sensitive to the rock conditions. From the small to large cross sections the TBM costs is increased by a factor of 2 – 3 when the rock conditions vary from good to poor boreability. The factor drops to between 20 % - 30 % for D&B. 5- Total construction costs In the total construction costs, items such as rock support, site preparation and operation, administration, planning and interest during construction are considered. In contrast between TBM and D&B, many of the items are similar or have totally minor impacts. Some items need to be assessed carefully when comparing the methods with regard to the total construction costs as follows: • Rock support • Equivalent hydraulic cross sections (for water tunnels) • Adits (in long tunnels) • Environmental aspects • Risk analysis • TBM manufacturing and assembly Usually, the circular cross section of the bored tunnels has better stability, resulting in less rock support requirements than D&B, indicating the advantage of bored tunnels. However, the more favourable the rock conditions for tunnel boring are (i.e. more fractured media), the heavier rock support is required and the TBM needs to be well equipped for such poor conditions. For unlined water tunnels, due to more head losses in the D&B tunnels, considering equivalent hydraulic cross section, the D&B tunnels should be 1.7 (for 10 m2) to 1.55 (for 100 m2) times larger the TBM cross section [2], also a benefit of the tunnel boring in the unlined water tunnels. Tunnels may be longer for tunnel boring than for D&B; this may facilitate a more favourable tunnel system layout with fewer adits. Regarding environment and occupational aspects, the tunnel boring has many advantages such as less pollution and better safety, but reuse of the bored material is less. As described earlier, the rock boreability is the most important factor for penetration rate and costs. Therefore, the uncertainty or geological risk connected to the bored tunnels is larger compared to D&B. Furthermore, dealing with the possible weak zones during the boring will increase the risk of the planned construction time and costs. The D&B has a better flexibility to cope with such conditions. The TBM is normally designed and manufactured or rebuilt for a specific project. The time needed for manufacturing or rebuilding of the machine and back-up, transport and assembly at the site may take from four months to one year [16]. 6- Conclusions Selection of the most economic excavation method for tunnels is not always straightforward, requiring updated practical tools. NTNU predictions models are developed for such purposes to facilitate selection of the efficient and cost effective methods. Among a variety of parameters which may affect the excavation time and costs, the rock conditions and tunnel geometry impacts have been investigated.

96

By far, rock conditions are the decisive parameters for both methods; in particular the TBM method where from poor to good boreability, the advanced rate is increased by a factor of 3 - 4 and the excavation costs is reduced by a factor of 2 – 3. Whereas in the D&B the range of rock conditions influence is limited to 20 - 40 % on the advance rate and 20 - 30 % on the excavation costs. The weekly advance rate of TBM is notably higher than the D&B, especially for small cross sections. By increasing the cross section area, the advance rate of TBM is dramatically reduced. The D&B has a more flat and linear advance rate as a function of tunnel cross section area. In hard rock and larger cross sections, the D&B is the cost effective method, for small and long tunnels with appropriate boring conditions, the TBM has lower excavation costs. 7- References [1] Norwegian Tunnelling Society; Norwegian Tunnelling, Publication No.14, 2004. [2] NTNU; Report 1A-98 Hard Rock Tunnel Boring – Design and Construction, Department of Civil and Transport Engineering, 1998. [3] NTNU; Report 1B-98 Hard Rock Tunnel Boring – Advance Rate and Cutter Wear, Department of Civil and Transport Engineering, 1998. [4] NTNU; Report 1C-98 Hard Rock Tunnel Boring – Costs, Department of Civil and Transport Engineering, 1998. [5] NTNU; Report 1D-98 Hard Rock Tunnel Boring – Geology and Siteinvestigations, Department of Civil and Transport Engineering, 1998. [6] NTNU; Report 1E-98 Hard Rock Tunnel Boring – Performance Data and Back-mapping, Department of Civil and Transport Engineering, 1998. [7] NTNU; Report 1F-98 Hard Rock Tunnel Boring – The Boring Process, Department of Civil and Transport Engineering, 1998. [8] NTNU; Report 2A-05 Drill and Blast Tunnelling – Blast Design, Department of Civil and Transport Engineering, to be published 2007. [9] NTNU; Report 2B-05 Drill and Blast Tunnelling – Advance Rate, Department of Civil and Transport Engineering, to be published 2007. [10] NTNU; Report 2C-05 Drill and Blast Tunnelling – Costs, Department of Civil and Transport Engineering, to be published 2007. [11] Zare, S.; Prediction model and simulation tool for time and costs of drill and blast tunnelling, PhD Thesis, Department of Civil and Transport Engineering, to be published 2007. [12] Zare, S.; Bruland, A.; “Progress of D&B tunnelling efficiency with relation to excavation time and costs”, ITA World Tunnel Congress, Prague, 2007. [13] Bruland, A.; “Prediction model for performance and costs”, in Norwegian TBM Tunnelling, Publication No.11, Norwegian Tunnelling Society, 1998. [14] NTNU; Report 13A-98 Drillability – Test Methods, Department of Civil and Transport Engineering, 1998. [15] Blindheim, O. T.; Bruland, A.; “Boreability testing”, in Norwegian TBM Tunnelling, Publication No.11, Norwegian Tunnelling Society, 1998. [16] Bruland, A.; “Future demands and development trends”, in Norwegian TBM Tunnelling, Publication No.11, Norwegian Tunnelling Society, 1998.

97

98

Appendices

Appendix A:

Field data

Appendix B:

Equipment price and lifetime

Appendix C:

Cost model in simulation tool (TunSim)

99

100

Appendix A

Field data

Table 1 Drilling data

Diameter (mm)

Number

Drillhole length (m)

Type

Number

DRI

Penetration (cm/min)

Drilling, moving etc.

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

82 82 82 82 63.7 86.9 79 108 108 71 71 62.7 70.1 78.9 105 28.9 25.2 31.7

48 48 48 48 48 48 48 51 51 48 48 48 48 48 48 48 48 45

106 106 106 106 100 156 108 116 118 103 103 115 119 108 123 68 68 78

102 102 102 102 102 102 102 102 102 127 127 127 102 102 102 102 102 102

4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4 4

5.15 5.15 5.15 5.15 5.1 5.2 5.1 5.1 5.65 5 5 5.11 5.15 5.15 5.15 5.15 5.15 5.15

COP 1838 HC 90 COP 1838 COP 1838 HC 90 COP 1838 HC 90 HC 90 HC 90 HC 90 HC 90 HC 90 COP 1838 COP 1838 COP 1838 COP 1838 COP 1838 HL 550

3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3 3

50 50 57 57 35 55 50 47 47 65 65 65 42 42 42 42 42 52

242 202 225 241 160 210 214 177 194 229 249 213 230 230 230 230 230 152

141 127 138 123

147 155 157 124 110 143

Total

Number

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18

Rig time

Diameter (mm)

Drilling time (min)

Area (m2)

Hammer

Project number

Large holes

Data set no.

Blast holes

15 15 15 15 16 15 15 23 20 17 10 15

176 159 161 135 176 176 175 189 187 157 130 166 158 160 150 134 142 240

Table 2 Charging data

2.1 2.03 2.14 2.07 1.84

1.2 1.2 1.3 1.2

Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP

2 2 2 2 2 2

67 67 65 68 0 0

Total

ANFO ANFO ANFO ANFO Slurry Slurry

Rig time

106 106 106 106 100 156

Charging time

Type

48 48 48 48 48 48

Number of charging lines

Number

82 82 82 82 63.7 86.9

Detonator

Diameter (mm)

1 1 1 1 2 2

Primer percentage (%)

Area (m2)

1 2 3 4 5 6

Charging time (min)

Consumption (kg/asm3)

Project number

Explosives

Data set no.

Blast holes

84.3 84.6 76 75 43 105

101

Appendix A 7 8 9 10 11 12 13 14 15 16 17 18

3 4 4 5 5 6 7 7 7 7 7 8

Field data

79 108 108 71 71 62.7 70.1 78.9 105 28.9 25.2 31.7

48 51 51 48 48 48 48 48 48 48 48 45

108 116 118 103 103 115 119 108 123 68 68 78

Slurry ANFO ANFO Slurry Slurry Slurry Slurry Slurry Slurry Slurry Slurry ANFO

1.85 1.79

1.97 1.9 1.6 1.32 2.86 3.3 2.76

3.6

Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP Nonel LP

2.4 2.3 2.2 2.3 2.2 2.2 36

2 2 2 2 2 2 2 2 2 2 2 2

59

47 45 38 69 63 83 44 40

64 0 0 0 0 0 10 10 10 10 10

52 45 40.1 89 83 103 64 60 111

Table 3 loading and hauling data

Pull (%)

Drillhole length (m)

Equipment

1.15 1.15 1.15 1.15 1.15 1.15 1.11 1.11 1.11 1.15 1.12 1.12 1.12 1.12 1.16 1.247 1.14 1.15 1.15 1.15 1.21

96 96 96 96 96 96 90 90 90 92 92 91 92 91 94.0 94.7 97 97 94 94 86

5.15 5.15 5.15 5.15 5.15 5.15 5.1 3 5.2 5.1 5.1 5.65 5.1 5.65 5 5 5.11 5.15 5.15 5.15 5.15

Brøyt X43- Truck Brøyt X43- Dump Truck Brøyt X53-Truck Volvo L330 - Truck Volvo L330 - Dump Truck Volvo L330 - Dump Truck Cat 980G-Truck/Dump Truck Cat 980G-Truck/Dump Truck Cat 980F-Truck/Dump Truck Brøyt X53-Truck/Dump Truck Brøyt X53-Truck Brøyt X53-Truck Volvo L330 - Truck Volvo L330 - Truck Volvo L330 - Dump Truck Brøyt X53- Dump Truck Volvo L330 - Dump Truck Volvo L330 - Dump Truck Cat 972G - Truck Cat 972G - Truck Cat 963 - Truck

102

247.3 277.4 272.7 276.4 322.5 305.8 134.2 134.2 134.2 241 220 225 260 260 270 216 278 270 89 89

5 5 5 5 5 5

109 163 175 138 151 86 130 77 104

Total

Overbreak factor

82 82 82 82 82 82 63.7 63.7 86.9 79 108 108 108 108 71 73.9 62.7 70.1 28.9 25.2 31.7

Rig time

Area (m2)

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

Loading

Project number

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21

Capacity (asm3/h)

Data set no.

Loading time (min)

18 35 35 15 15 15

205 205 182 212 172 172 133 104 200 149 219 232 169 184 119

19

119 161 169 153

Appendix A

Field data

Table 4 Ventilation and scaling data

Explosives Type

Ventilation time (min)

Drillhole length (m)

82 82 82 82 63.7 63.7 86.9 79 108 108 71 71 62.7 70.1 78.9 104.5 28.1 25.2 31.7

ANFO ANFO ANFO ANFO Slurry Slurry Slurry Slurry ANFO ANFO Slurry Slurry Slurry Slurry Slurry Slurry Slurry Slurry ANFO

15 15 15 15 12 11 11 0 20 20 10 0 7.5 5 5 5 5 5 0

5.15 5.15 5.15 5.15 5.1 3 5.2 5.1 5.1 5.65 5 5 5.11 5.15 5.15 5.15 5.15 5.15 5.15

Blastability

Area (m2)

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

Use of scaling jumbo

Project number

1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19

Scaling from the pile

Data set no.

Scaling time (min)

72 13 78 21 104 65 130 16 125 140 59 60 68 80 89 117 83 77 92

Medium Medium Medium Medium Medium Medium Medium Good Medium Medium Medium Medium Good Medium Medium Medium Medium Medium Poor

103

Appendix A

104

Field data

Appendix B

Equipment price and lifetime

Rail-mounted drilling jumbo data Atlas Copco 2-boom drilling jumbo Hammer type Hammer lifetime Jumbo price Fixed repair cost Service cost Electricity consumption 3-boom drilling jumbo Hammer type Hammer lifetime Jumbo price Fixed repair cost Service cost Electricity consumption

Model

eh NOK NOK/eh NOK/eh kWh/eh

1838 6000 4,343,274 200 100 90

1838 6000 3,900,000 200 100 90

eh NOK NOK/eh NOK/eh kWh/eh

1838 6000 5,953,504 300 150 135

1838 6000 5,400,000 300 150 135

Wheel-mounted drilling jumbo data Atlas Copco

Model

2-boom drilling jumbo Hammer type Hammer lifetime Jumbo price Fixed repair cost Service cost Electricity consumption

eh NOK NOK/eh NOK/eh kWh/eh

1838 6000 5,414,000 200 100 150

1838 6,000 4,900,000 200 100 150

3-boom drilling jumbo Hammer type Hammer lifetime Jumbo price Fixed repaire cost Service cost Electricity consumption

eh NOK NOK/eh NOK/eh kWh/eh

1838 6000 7,442,000 300 150 225

1838 6,000 6,700,000 300 150 225

eh NOK NOK/eh NOK/eh kWh/eh

1838 6000 9,470,000 400 200 380

1838 6,000 8,500,000 400 200 380

4-boom drilling jumbo Hammer type Hammer lifetime Jumbo price Fixed repair cost Service cost Electricity consumption

The jumbo prices are reduced by10 % and rounded. Atlas Copco 4-boom jumbo price is estimated based on 2 and 3 boom jumbo prices. Rail-mounted jumbo price using 1838 hammers is estimated based on wheel-mounted jumbos.

105

Appendix B

Equipment price and lifetime

Drill steel price 45

Drillhole diameter (mm) 48 51 57

578 485 475 500

830 530 510 600

998 580 545 700

1,197 810 655 900

1,208 850 820 1,000

4,015 4,200 3,050 3,800

4,015 4,200 3,050 3,800

4,015 4,200 3,050 3,800

4,015 3,900 3,050 3,700

4,015 4,200 3,050 3,800

1,754 1,980 1,650 1,800

1,754 1,980 1,650 1,800

1,754 1,980 1,650 1,800

1,754 1,980 1,650 1,800

1,754 1,980 1,650 1,800

525 710 500 600

525 710 500 600

525 710 500 600

525 710 500 600

525 710 500 600

Supplier 2

Supplier 3

Model

10 -

10 30

64

NOK

Drill bit Supplier 1 Supplier 2 Supplier 3 Model Drill rod (18 feet) Supplier 1 Supplier 2 Supplier 3 Model

NOK

Shank adapter

NOK Supplier 1 Supplier 2 Supplier 3 Model NOK

Coupling Supplier 1 Supplier 2 Supplier 3 Model

Drill Bit regrinding data Supplier 1 Number of regrinding Price for one regrinding

NOK

5 16

Supplier 3 prices are reduced 50 % from published price list. The model prices are average and rounded.

106

10 45

Appendix B

Equipment price and lifetime

Drill steel lifetime 45

Drillhole diameter (mm) 48 51 57

64

dm

Drill rod Good drillability Supplier 1 Supplier 2 Supplier 3 Model

4,000 5,000 5,000 4,700

3,800 4,750 4,750 4,400

3,600 4,500 4,500 4,200

3,240 4,050 4,050 3,800

2,880 3,600 3,600 3,400

Supplier 1 Supplier 2 Supplier 3 Model

1,500 3,750 2,778 2,700

1,425 3,563 2,639 2,500

1,350 3,375 2,500 2,400

1,215 3,038 2,250 2,200

1,080 2,700 2,000 1,900

Supplier 1 Supplier 2 Supplier 3 Model

4,000 7,000 5,000 5,300

3,800 6,650 4,750 5,100

3,600 6,300 4,500 4,800

3,240 5,670 4,050 4,300

2,880 5,040 3,600 3,800

Supplier 1 Supplier 2 Supplier 3 Model

2,000 5,250 2,778 3,300

1,900 4,988 2,639 3,200

1,800 4,725 2,500 3,000

1,620 4,253 2,250 2,700

1,440 3,780 2,000 2,400

Supplier 1 Supplier 2 Supplier 3 Model

6,000 3,500 5,000 4,800

5,700 3,325 4,750 4,600

5,400 3,150 4,500 4,400

4,860 2,835 4,050 3,900

4,320 2,520 3,600 3,500

Supplier 1 Supplier 2 Supplier 3 Model

2,000 2,625 2,778 2,500

1,900 2,494 2,639 2,300

1,800 2,363 2,500 2,200

1,620 2,126 2,250 2,000

1,440 1,890 2,000 1,800

Poor drillability

dm

Shank adapter Good drillability

Poor drillability

dm

Coupling Good drillability

Poor drillability

The underlined numbers are taken from suppliers info. The rest of the data are estimated based on direct relation between penetration rate and lifetime (dm), assuming constant life in hours for different diameters. The percent of reduction in penetration rate are estimated based on Fig 2.4 in PR 2A-95. The supplier 2 poor drillability are estimated based on 25% reduction in penetration rate from good drillability (DRI=65) to poor drillability (DRI=37) according to Fig 2.3 in PR 2A-95. The model data are averaged and rounded.

107

Appendix B

Equipment price and lifetime

Explosives and initiation system data Supplier 1

Supplier 2

Model

20.48

18 18 18 7 8

Explosives price Dynamite Dynorex Kemix A ANFO Emulsion

NOK/kg NOK/kg NOK/kg NOK/kg NOK/kg

20.20 20.20 20.20 7.70 ~8.50

19.43 7.14 8.93

NONEL detonator NONEL LP Detonator, 4.8 m NONEL LP Detonator, 6 m NONEL LP Detonator, 7.8 m NONEL Tube NONEL Tube,150m Miscellaneous

NOK each NOK each NOK each NOK/m NOK/round NOK/round

16.65 17.90 19.96 1.17

19.74 1.15

Electrical detonator Millisecond detonator, 4 m Millisecond detonator, 6 m Millisecond detonator, 10 m Halfsecond detonator Wire Wire, 25 m Miscellaneous

NOK each NOK each NOK each NOK each NOK/m NOK/round NOK/round

12.60 15 24.74

12,71 15.02 25.10

1.50

1.68

18 1 150 30

13.50 13.50 1.50 40 30

The model prices are average, ~10% off and rounded.

Wheel loader data Type

Economic lifetime Fixed repair cost Service cost

NOK (1000) eh NOK/eh NOK/eh

Fuel consumption

l/eh

Loader price

Life of tyres

NOK (1000) eh

Operating weight

ton

Price of one set of tyres

108

Cat 966G II

Cat 972G II

Cat 980G II

Cat 988G

Volvo L330E

1,700

1,900

2,500

4,100

3,300

9,600 95 31

9,900 99 33

10,650 107 38

13,650 133 50

14,400 141 53

27

30

36

62

68

126

126

167

220

220

2,000

2,000

2,000

2,000

2,000

23

25

30

50

55

Appendix B

Equipment price and lifetime

Track-type loader data Type Loader price Economic lifetime Fixed repair cost Service cost

Cat 963C 1,500,000 9,150 115 30

Cat 973C 2,400,000 10,350 131 35

NOK eh NOK/eh NOK/eh

Fuel consumption

l/eh

23

41

Operating weight

ton

20

28

The loader and tyre prices are reduced 10 % and rounded. Lifetime is estimated based on 9600 eh for 23 ton loader and 14400 eh for 55 ton loader (80 % of Cat info). Fixed repair cost and service cost are taken from 1988 model and increased 50 % for price level correction. For Cat 972G and Volvo 330 fixed repair and service cost are estimated by other wheel loader data vs. operating weight. For fuel consumption and tyre life, see Cat performance handbook edition 35.

Excavator data Brøyt ED 600T Excavator price Economic lifetime Fixed repair cost Service cost Electricity consumption Fuel consumption

NOK eh NOK/eh NOK/eh kWh/eh l/eh

Brøyt ED 600T 4,500,000 20,000 130 25 120 30

Model 4,100,000 20,000 130 25 120 0

Cost of helping loader

NOK/eh

-

1,000

Excavator price Economic lifetime Fixed repair cost Service cost Electricity consumption Fuel consumption

NOK eh NOK/eh NOK/eh kWh/eh l/eh

Brøyt ED 1000T 5,500,000 20,000 160 30 150 40

Model 5,000,000 20,000 160 30 150 0

Cost of helping loader

NOK/eh

-

1,000

Excavator data Brøyt ED 1000T

Excavator price is rounded and reduced by 10 %. Cost of helping loader is taken from 1988 model and increased 50 % for price level correction.

109

Appendix B

Equipment price and lifetime

Contract transport data Price ( NOK/lm3) 7.40 8.40 9.20 10.20 10.90 11.70 12.40 13 13.50 14 16.90 19.80 22.50 25 27.60 30.20 32.70

Distance in metres (up to) 100 200 300 400 500 600 700 800 900 1000 2000 3000 4000 5000 6000 7000 8000

Roadway data Material and placement price (at 3 km transport length) NOK/m3 Maintenance price NOK/eh The prices are taken from 1988 model and increased 50 % for price level correction.

90 1050

Diesel locomotive data Locomotive Weight Locomotive price Economic lifetime Fixed repair cost Service cost Fuel consumption

ton NOK eh NOK/eh NOK/eh l/eh

15 1,160,000 13,000 35 25 23

20 1,280,000 14,500 42 29 27

25 1,450,000 16,000 50 33 30

m3 NOK eh NOK/eh NOK/eh kWh/eh

9 910,000 18,000 110 26 22

11.5 950,000 18,750 115 27 22

14 990,000 19,500 120 28 22

Shuttlecar data Capacity of each car Price of each car Economic lifetime Fixed repair cost Service cost Electricity consumption

110

Appendix B

Equipment price and lifetime

Haggloader data Type Loader price Economic lifetime Fixed repair cost Service cost Electricity consumption

NOK eh NOK/eh NOK/eh kWh/eh

8HR1 1,660,000 9,000 185 36 45

8HR2 1,820,000 9,000 185 36 45

Equipment price is taken from GIA company info and reduced 10%. Lifetime is estimated based on 1988 model data and increased 50 % . Fixed repair cost and service cost are taken from 1988 model and increased 50 % for price level correction. Electricity and fuel consumption are taken from GIA company info.. Locomotive fuel consumption is based on 70 % efficiency.

Rail data Rail price NOK/kg Rail weight kg/m Length of each section m Connection (fish plate) price NOK each Bolt price NOK each Sleeper price NOK each Space between sleeper m Switches price NOK each Freight price NOK/kg All data are taken from 1988 model (not corrected for price level)

1 25 6 80 20 100 0.9 21100 0.22

Tip and tip station data Trackless/track transport Cost of placing and treatment at tip Track transport Loader cost Tip station at surface Cost of foundation and concrete work Other cost Tip station at underground Cost of tip station excavation Cost of foundation and concrete work Other cost

NOK/asm3

NOK/h

3

725

NOK NOK

150,000 15,000

NOK/asm3 NOK NOK

128 150,000 22,500

The price is taken from 1988 model and increased 50 % for price level correction. The loader price is not corrected by price level factor.

111

Appendix B

Equipment price and lifetime

Ventilation data Korfmann

Model

580,000 870,000

522,000 783,000

Fan (2 x 55) 110 kW (2 x 110) 220 kW

NOK NOK

Duct diameter 1 m diameter 2 m

NOK/m NOK/m

140 300

130 270

diameter 1 m diameter 2 m

NOK each NOK each

130 230

120 210

NOK m NOK/m NOK each NOK/m NOK/m

13,000 20 -

13,000 20 15 8 12 8

Duct coupling

Freight price per fan Duct section length Duct repair cost Duct bolt price Duct wire price Other duct cost

Fan and duct prices are reduced by 10 %. Duct repair, bolt and wire prices are taken from 1988 model and increased 50 % for price level correction.

Electrical installations data Price of outside transformer Price of inside transformer Freight price per transformer Cable price Earth connection cable price High voltage connection price Other cable cost

NOK NOK NOK NOK/m NOK/m NOK each NOK/m

390,000 255,000 9,000 90 20 4,635 15

The prices (except cable price) are taken from 1988 model and increased 50 % for price level correction.

Water supply data Pump price Pipe price Bolt price Other water supply cost

NOK NOK/m NOK/m NOK/m

153,000 100 9 18

The prices (except pipe price) are taken from 1988 model and increased 75 % according to the pipe price.

112

Appendix B

Equipment price and lifetime

Miscellaneous cost Tunnel cross section area (m2) Horizontal adit Decline adit

6 81 170

NOK/m NOK/m

80 275 365

The prices are taken from 1988 model and increased 50 % for price level correction.

Labour cost Social security cost Wages of workers in tunnel Wages of workers outside the tunnel Wages of external workers Wages of other workers

% NOK/h NOK/h NOK/h NOK/h

50 315 210 210 210

Scaling cost Scaling from the pile Tunnel cross section area Good blastability Poor blastability

m2 NOK/round NOK/round

Using scaling jumbo

NOK/h

6 148 173

80 246 288 1000

The prices (except scaling jumbo) are taken from 1988 model and increased 50 % for price level correction.

General cost data Investment tax Real interest rate Electricity price Fuel price

% % NOK/kWh NOK/l

0 5 0.65 4

All prices are excluding tax. All repair and service costs are excluding wages.

.

113

Appendix B

114

Equipment price and lifetime

Appendix C

Cost model in simulation tool (TunSim)

In the Chapter 4 the simulation tool is explained and input, out put, advance rate model and drilling cost is presented. To avoid repetition in the appendix only cost model in the simulation tool is presented. General cost input General cost input data

Unit

Model

Investment tax Interest rate Electricity price Fuel price

% % NOK/kWh NOK/l

Adit slope Adit length Distance to tip

m m

0.0 5.0 0.65 4.00 Horizontal 300 300

Swelling factor for blasted rock Percent of Unforeseen costs Price level correction factor

User

1.65 %

10 1.00

Drilling cost Drilling input data Drilling jumbo data Downtime factor Percent of lifetime used in tunnel Time between tunnel sites for interest payment Jumbo price Economic lifetime (in percussion hours) Fixed repair costs Service costs Electricity consumption Drill steel data Drill bit price Drill rod price Shank adapter price Coupling price Drill bit lifetime Drill rod lifetime Shank adapter lifetime Coupling lifetime Storing and freight costs Number of regrinds per bit Price for one regrinding

Unit

% % NOK h/hammer NOK/eh NOK/eh kWh/eh

NOK NOK NOK NOK dm dm dm dm % NOK

Model

User

0.5 75 15 6,700,000 6,000 300 150 225

600 3800 1800 600 437 3500 4200 3500 10.0 10 30

115

Appendix C

Cost model in simulation tool (TunSim)

Time model data Penetration rate Advance per week Number of weeks per year Pull per round Equivalent drillmetre per round Effective hours per round

Cost calculation Economic lifetime Economic useful lifetime of jumbo Time used in tunnel Depreciation factor Average rest value factor Number of rounds per week Number of years for interest payment Drillmetre per effective hour Depreciation cost Interest cost Fixed repair cost Variable repair cost Downtime cost Electricity cost Service cost Jumbo cost

cm/min m/week week/year % dm/round eh/round

dm/hammer dm dm

round/week year dm/eh NOK/dm NOK/dm NOK/dm NOK/dm NOK/dm NOK/dm NOK/dm NOK/dm NOK/round NOK/m

Drill bit cost Bit regrinding cost Drill rod cost Shank adapter cost Coupling cost Storing and freight cost Drill steel cost

NOK/dm NOK/dm NOK/dm NOK/dm NOK/dm NOK/dm NOK/dm

Total drilling cost

NOK/dm NOK/round NOK/m

219 76.3 44 91 480 1.92

787,689 1,929,436 1,447,077 0.9375 0.4375 16.8 4.7 249.9 4.34 0.48 1.20 1.74 1.47 0.59 0.60 10.41 4998 1098 1.37 0.69 1.09 0.43 0.17 0.31 4.05 14.46 6944 1526

Explosives cost Explosives input data

Unit

Cartridged explosives Percent of dynamite (Dynomit) Percent of contour charge (Dynorex) Percent of stope charge (Kemix A)

% % %

116

Model

50 30 20

User

Appendix C Bulk explosives Type of primer Proportion of primer Primer price General Explosives price Detonator type Initioation system price

Cost model in simulation tool (TunSim)

% NOK/kg

NOK/kg NOK/round

Dynamite 5 18

8 NONEL 1782

kg/m3

1.44

Time model data Number of detonators (blastholes ) Correction for drillhole length Pull per round

%

89 1.00 91

Explosives and detonators cost Explosives cost Primer cost Initiation system cost Sum

NOK/m NOK/m NOK/m NOK/m

658 78 392 1128

Total explosives consumption

Scaling cost Scaling input data

Unit

Model

Scaling jumbo (excavator) cost Extra rigging time of scaling jumbo Scaling cost

NOK/hours min/round NOK/round

Time model data Use of scaling jumbo Scaling time Pull per round

min %

Yes 52 91

Scaling cost

NOK/m

226

User

1000 10 1030

Loading cost Loading input data

Unit

Model

General Downtime factor Percent of lifetime used in tunnel Time between tunnel sites for interest payment Additional use of loader per round Loader price Economic lifetime

% % min NOK eh

0.5 50 10 0 3,300,000 14,400

User

117

Appendix C

Cost model in simulation tool (TunSim)

Fixed repair cost Service cost Electricity consumption Fuel consumption

NOK/eh NOK/eh Kwh/eh l/eh

Wheel loader Price of one set of tyres Life of tyres

NOK eh

Excavator Cost of helping loader

NOK/eh

Time model data Loading time (loading + rig time) Advance per week Number of working weeks per year Pull per round Overbreak factor

min m/week week/year %

141 53 0 68

220,000 2,000

0

88 76.3 44 91 1.15

Cost calculation Economic useful lifetime Time used in tunnel Depreciation factor Average rest value factor Corrected loading time Number of years for interest payment Depreciation cost Interest cost Fixed repair cost Variable repair cost Downtime cost Fuel cost Electricity cost Service cost

min/round year NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh

Depreciation of tyres Interest of tyres

NOK/eh NOK/eh

121 6

Cost of helping loader

NOK/eh

0

Loader cost

NOK/eh NOK/round NOK/m

eh eh

NOK/m3

118

11,758 5,879 0.7500 0.5833 88 6.0 421 98 141 94 117 272 0 53

1322 1944 427 6

Appendix C

Cost model in simulation tool (TunSim)

Contract transport Contract transport input data

Unit

Model 3

Basic cumulated cost for truck Cost reduction factor (when using dump truck)

NOK/lm

User

18.41 0.85

Time model data Overbreak factor

1.15

Cost calculation NOK/lm3

Muck transport Muck transport

NOK/sm

Transport cost

NOK/m

15.65

3

25.82 1782

Roadway cost Roadway input data

Unit

Model

3

Volume of road material

m /m

User

2.15 3

Material and placement price (at 3 km transport length) Maintenance time (at 3 km transport length) Maintenance price

NOK/m h/week NOK/h

Time model data Advance per week

m/week

90 2 1050

76.3

Cost calculation Correction for transport length ≠ 3 km Material and placement cost Correction for maintenance length ≠ 3 km Maintenance cost

NOK/m NOK/m

1.1 209 0.70 19

Roadway cost

NOK/m

228

Tip cost, trackless tunnelling Trackless transport tip input data

Unit

Cost of placing and treatment of tip Tip machine utilization factor

NOK/asm3

Time model data Overbreak factor

Tip cost

Model

User

3 0.77

1.15

NOK/m

159

119

Appendix C

Cost model in simulation tool (TunSim)

Track transport cost Track transport input data

Unit

General data Time between tunnel sites for interest payment Additional use per round

% min

Cycle time data Fixed time Driving speed Meet time (per meeting)

min km/h min

Locomotive data Downtime factor Percent of lifetime used in tunnel Locomotive weight Locomotive price Economic lifetime Fixed repair cost Service cost Fuel consumption Shuttlecar data Downtime factor Percent of lifetime used in tunnel Number of shuttlecars

20 5

% ton NOK eh NOK/eh NOK/eh l/eh

% m3 NOK eh NOK/eh NOK/eh kWh/eh

Time model data Transport time (loading time)

min

Cost calculation Locomotive cost Economic useful lifetime Time used in tunnel Depreciation factor Average rest value factor Corrected transport time Number of years for interest payment

120

14.1 15 2

0.5 75 20 1,280,000 14,500 42 29 27

0.2 100 4

Capacity of each car Shuttlecar price Economic lifetime Fixed repair cost Service cost Electricity consumption

Loading capacity Advance per week Number of working weeks per year Pull per round Overbreak factor

Model

11.5 950,000 18,750 115 27 22

70 3

asm /h m/week week/year %

eh eh

min/round year

268 76.3 44 91 1.15

11,839 8,879 0.9375 0.4375 75 11.5

User

Appendix C

Cost model in simulation tool (TunSim)

Depreciation cost Interest cost Fixed repair cost Variable repair cost Downtime cost Fuel cost Service cost Sum

NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh

135 36 42 54 48 108 29 452

Shuttlecar cost Economic useful lifetime Time used in tunnel Depreciation factor Average rest value factor Corrected Ttransport time Number of years for interest payment Depreciation cost Interest cost Fixed repair cost Variable repair cost Downtime cost Electricity cost Service cost Sum

min year NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh NOK/eh

17,116 17,116 1.0000 0.3333 75 22.1 222 82 115 185 60 29 27 720

Total cost

NOK/eh

1172

Transport cost

NOK/sm3 NOK/m

38 2615

eh eh

Rail cost Rail input data

Unit

Model

Rail price Rail weight Length of each track section Connection (fish plate) price Bolt price Sleeper price Space between sleepers Switches price Freight price

NOK/kg kg/m m NOK each NOK each NOK each m NOK each NOK/kg

1 25 6 80 20 100 0.9 21100 0.22

Location of tip station Extra track length in adit and tip Extra track length in meeting point Total track length Number of switches

m m m

User

Surface 700 250 3950 15

121

Appendix C

Cost model in simulation tool (TunSim)

Time between tunnel sites Rail lifetime Number of sites for depreciation Depreciation percent on lifetime Depreciation percent on sites

% %

Time model data Excavation time

year

0.89

Cost calculation Rail cost Connection (fish plate) cost Bolts cost Sleepers cost Switches cost Sum

NOK NOK NOK NOK NOK NOK

197,500 210,667 105,333 438,889 316,500 1,320,239

Maintenance factor Excavation time + time between sites Number of sites over lifetime Excavation length over lifetime Excavation length over sites Depreciation on lifetime Depreciation on sites Interest on lifetime Interest on sites Freight cost Maintenance cost Rail cost

year year

1 10 3 50 50

1.90 1.89 5.28 15842 9000 42 73 10 10 17 17 170

year m m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m

Tip cost, track tunnelling Track transport tip input data

Unit

Model

Location of tip station Tip machine utilization factor

Surface 0.77

Loader loading capacity Loader cost

asm3/h NOK/h

Tip station at surface Cost of foundation and concrete work Extra transport length by truck

NOK m

Transport cost

NOK/lm3

Cost of placing and treatment of tip Correction loading capacity (60 m movement) Other costs

122

NOK/asm NOK

104 725

150,000 0 0 3

0 0.7 15,000

User

Appendix C

Cost model in simulation tool (TunSim)

Tip station at underground Excavation volume of tip station

asm3

4000 3

Cost of tip station excavation Cost of foundation and concrete work

NOK/asm NOK

Transport cost by truck

NOK/lm3

Cost of placing and treatment of tip Other costs

NOK/asm NOK

Time model data Overbreak factor

128 150,000 11.7

3

3 22,500

1.15

Cost calculation Tip station at surface Cost of foundation and concrete work Cost of loader Transport cost Cost of placing and treatment of tip Other costs Sum

NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m

50 528 0 0 5 583

Tip station at underground Cost of tip station excavation Cost of foundation and concrete work Cost of reloading Cost of transport Cost of placing and treatment of tip Other costs Sum

NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m

171 50 369 1023 159 8 1779

Tip cost

NOK/m

583

Ventilation design Ventilation design input

Unit

Model

Loader fuel consumption Amount of explosives per round Dilution factor, F Tunnel slope, s Rolling resistance, fr Relative load factor, n Duct leakage, pv

l/eh kg/round

%

68 394 127.6 0.003 0.04 1.6 10

Maximum air velocity in duct Maximum fan pressure

m/s pa

24.4 5000

User

123

Appendix C

Cost model in simulation tool (TunSim)

Time model data Loading capacity Loading time

asm3/h min/round

267.5 70

Calculation of ventilation system Necessary air for rock pile, Qr

m3/s

6.14

3

Necessary air for loader, Ql

m /s

Necessary air at tunnel face, Qo=Qr+Ql

m3/s 3

22.71 28.86 .

Necessary air for transport, qb

m /s m

0.020

Total tunnel length including adit Theoretical ventilated transport length, Xk

m m

3300 4404

Q for Xk equal to total tunnel length

m3/s

Qxk

94

3

116

3

94

m /s

Necessary total fan capacity, Qt

m /s

Number of ducts Duct diameter Calculated pressure losses Calculated max. air velocity in duct

m pa m/s

2 1.6 3737 23.4

Calculated duct flow

m3/s

47.0

Number of fans Calculated fan capacity Necessary power of each fan

2 3

m /s kW/fan

47.0 176

Ventilation cost Ventilation input data Calculated capacity (Q) Calculated pressure (p) Necessary fan output Fan efficiency factor Necessary fan installed power Number of fans outside the tunnel Number of fans at tunnel face Number of ducts Duct diameter Fan power utilization factor Fan price Freight price per fan Duct price Duct coupling price Duct section length Duct maintenance cost Duct bolt price Duct wire price

124

Unit 3

m /s pa kW kW

m NOK NOK NOK/m NOK each m NOK/m NOK each NOK/m

Model 47 3,737 176 0.8 219 2 1 2 1.6 1.0 781,609 13,000 214 174 20 15 8 12

User

Appendix C

Cost model in simulation tool (TunSim)

Other duct costs

NOK/m

Time between tunnel sites Fan lifetime Number of sites for depreciation of fan Duct lifetime Number of sites for depreciation of duct Depreciation percent on lifetime Depreciation percent on sites

year year

% %

Time model data working time per week Advance per week Excavation time

h/week m/week year

Cost calculation Total outside fans price Extra fans at face for two-way ventilation Fan installing and foundation cost Total fan price Excavation time + time between sites Number of sites over lifetime Excavation length over lifetime Excavation length over sites Depreciation on lifetime Depreciation on sites Interest on lifetime Interest on sites Electricity cost Service and repair cost Freight cost Fan cost Duct length Total duct price Duct coupling price Duct and coupling price Excavation time + time between sites Number of sites over lifetime Excavation length over lifetime Excavation length over sites Depreciation on lifetime Depreciation on sites Interest on lifetime Interest on sites Maintenance cost Bolt and wire cost Other duct costs Duct cost Ventilation cost

year

NOK NOK NOK NOK year m m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m m NOK/m NOK each NOK year m m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m

8 1 10 5 6 3 50 50

101 76.3 0.89

1,563,218 781,609 468,965 2,813,793 1.89 5.28 15,842 15,000 89 94 22 22 378 91 9 705 3,300 428 348 1,469,820 1.89 3.17 9,505 9,000 77 82 12 12 17 15 9 214 918

125

Appendix C

Cost model in simulation tool (TunSim)

Electrical installation cost Electrical installations input data

Unit

Model

Price of transformer outside tunnel Price of transformer inside tunnel Freight price per trasformer Cable price Earth connection cable price High voltage connection price Other cable costs Number of transformers outside the tunnel Number of transformers inside the tunnel

NOK NOK NOK NOK/m NOK/m NOK each NOK/m

Time between tunnel sites Transformer lifetime No. of sites for depreciation of transformer No. of sites for depreciation of cable Depreciation percent on lifetime Depreciation percent on sites

year year

% %

1 20 5 4 50 50

Time model data Excavation time

year

0.9

NOK year

390,000 255,000 9,000 90 20 4,635 15 1 1

Cost calculation Total transformer price Excavation time + time between sites Number of sites over lifetime Excavation length over lifetime Excavation length over sites Depreciation on lifetime Depreciation on sites Interest on lifetime Interest on sites Freight cost Transformer cost

m m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m

645,000 1.9 10.6 31,684 15,000 10 22 5 5 6 48

Cable length Total cable price Excavation length over sites Depreciation on sites Interest on sites High voltage connection cost Other cable costs Cable cost

m NOK m NOK/m NOK/m NOK/m NOK/m NOK/m

3,300 363,825 12,000 30 6 13 17 65

Electrical installations cost

NOK/m

113

126

User

Appendix C

Cost model in simulation tool (TunSim)

Water supply cost Water supply input data

Unit

Pump price Pipe price Bolt price Other water supply costs Number of pumps Distance from water source to adit

NOK NOK/m NOK/m NOK/m m

% %

1 4 2 6 3 50 50

Time model data excavation time

year

0.9

NOK year

306,000 1.89 2.11 6,337 6,000 24.1 25.5 2.4 2.4 54.5

year

m m NOK/m NOK/m NOK/m NOK/m NOK/m m NOK year

User

153,000 100 9 18 2 700

Time between tunnel sites Pump lifetime Number of sites for depreciation of pump Pipe lifetime Number of sites for depreciation of pipe Depreciation percent on lifetime Depreciation percent on sites

Cost calculation Pump price Excavation time + time between sites Number of sites over lifetime Excavation length over lifetime Excavation length over sites Depreciation on lifetime Depreciation on sites Interest on lifetime Interest on sites Pump cost

year year

Model

Pipe length Pipe and bolt price Excavation time + time between sites Number of sites over lifetime Excavation length over lifetime Excavation length over sites Depreciation on lifetime Depreciation on sites Interest on lifetime Interest on sites Pipe cost

m m NOK/m NOK/m NOK/m NOK/m NOK/m

4,000 436,000 1.89 3.17 9,505 9,000 22.9 24.2 3.4 3.4 54.0

Other cost

NOK/m

24.0

Water supply cost

NOK/m

133

127

Appendix C

Cost model in simulation tool (TunSim)

Labour cost Labour input data

Unit

Social security cost

%

Wages of workers in tunnel Number of workers in tunnel per shift Number of locomotive drivers per shift

NOK/h

315 3 0.0

Wages of workers outside the tunnel Number of workers outside the tunnel Working time for outside workers

NOK/h

210 4 37.5

h/week

Model 50

Wages of external workers Number of external workers Working time for external workers

NOK/h

Wages of other workers Number of other workers Working time for other workers

NOK/h h/week

210 0 33.5

Time model data Working time per week Advance per week

h/week m/week

101 76.3

Cost calculation Tunnel workers Locomotive drivers Outside workers External workers Other workers Labour cost

NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m

1876.6 0.0 619.4 154.8 0.0 2651

h/week

User

210 1 37.5

Niches cost Niches input data

Unit

Drilling cost

NOK/m3

25.4

3

18.8

3

3.8

3

7.1

3

29.7

3

2.6

3

15.3

3

44.2

3

147.0

Explosive cost Scaling cost Loading cost Transport cost Tip cost Ventilation cost Labour cost Sum

128

Model

NOK/m

NOK/m

NOK/m NOK/m

NOK/m

NOK/m NOK/m

NOK/m

User

Appendix C

Cost model in simulation tool (TunSim)

Time model data Volume of each niche Distance between niches

m3 m

Niches cost

NOK/m

85 300 41

Cost summary Cost summary

Unit

Drilling cost Explosive cost Scaling cost Loading cost Transport cost Roadway cost Rail cost Tip cost Ventilation cost Electrical installations cost Water supply cost Miscellaneous cost Labour cost Niches cost

NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m NOK/m

1,526 1,128 226 427 1,782 228 0 159 918 113 133 223 2,651 41

Sum elemental costs Unforeseen cost Correction price level Standard costs Rock support costs Total costs

NOK/m NOK/m

9,555 956 1.00 10,511 0.0 10,511

NOK/m NOK/m NOK/m

Model

User

129

>H7C.,-"-'"),&"'-'*"&eg^ciZYkZgh^dc >H7C.,-"-'"),&"'-(."-ZaZXigdc^XkZgh^dc >HHC&*%("-&-&

>ccdkVi^dcVcY8gZVi^k^in