Constancia Technical Report Ni 43 101

Norsemont Mining Inc.

CONSTANCIA COPPER PROJECT Definitive Feasibility Study Technical Report NI 43-101 Revision 0 28 September 2009

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0

Important notice This Constancia Copper Project Technical Report (Technical Report or Report) has been prepared for Norsemont Mining Inc. (Norsemont) by GRD Minproc Limited (GRD Minproc). The Technical Report is intended to be used by Norsemont subject to the terms and conditions of its definitive feasibility study contract with GRD Minproc. That contract permits Norsemont to file this report as a Technical Report with Canadian Securities Regulatory Authorities pursuant to provincial securities legislation. Except for the purposes legislated under provincial securities laws, any other use of this Report by any third part is at that party’s sole risk. The Technical Report should be read in the context of the methodology, procedures and techniques used and GRD Minproc's assumptions, and the circumstances and constraints under which the Technical Report was prepared. The Technical Report is to be read as a whole, and sections or parts of it should not be read or relied upon out of context. GRD Minproc has, in preparing the Technical Report, followed methodology and procedures, and exercised due care consistent with the intended level of accuracy of the study on which it is based, using its professional judgment and reasonable care. However, no warranty is given regarding the accuracy of estimates and other assumed values. It should also be borne in mind that all figures contained in the Technical Report are valid only as at the date of the Technical Report and may vary thereafter. GRD Minproc does not take responsibility for or adopt as its own sections of the Report prepared by or for information supplied by third persons or Qualified Persons not employed or engaged by GRD Minproc. This notice, which is an integral part of the Report, must accompany every copy of the Report.

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Title Page Project Name: Title: Location:

Constancia Copper Project Definitive Feasibility Study Technical Report Province of Chumbivilcas, Department of Cusco, Peru

Effective Dates: Effective Date of Technical Report: Effective Date of Mineral Reserves: Effective Date of Mineral Resources:

28 September 2009 28 September 2009 28 September 2009

Qualified Persons: 

David D. (Dan) Greig, B.Sc, M.A.I.G. (1684-1722), employed by GRD Minproc Limited as Principal Geologist, was responsible for overall preparation of the report.



Lynn Widenbar, B.Sc. (Geology), M.Sc., MAusIMM, employed by GRD Minproc Limited as a Geological Consultant, was responsible for data quality verification (Section 14) and for resource modelling and reporting (Section 17).



Ross Oliver, B.Eng (Mining), Aus.I.M.M. 105137, employed by GRD Minproc Limited as Manager Mining & Geology, was responsible for the preparation of Section 17.14 on Mineral Reserves and Section 18.1 covering Mining.



Greg Harbort, B.Eng (Met), University of Qld, 1985; Ph.D. University of Qld, 2005; MAusIMM Principal Process engineer for GRD Minproc Limited, was responsible for Mineral Processing and Metallurgical Testwork (Section 16), and for Plant Operating Costs and site infrastructure operating costs (Section18.14).



Craig Cuttriss, B.Sc, M.Eng, R.P.E.Q., MAusIMM, Principal Process Engineer and Study Manager for GRD Minproc Limited,was responsible for GRD Minproc’s engineering aspects of the Project as contained in Section 18.4, Section 18.9, Section 18.11, Section 18.13 and Section 18.17.



Thomas F. Kerr, M.Sc., President of Knight Piésold and Co. USA, Registered Professional Engineer (Civil and Geotechnical), P.Eng., in British Columbia (#14906) and Ontario (#90407230) and P.E. in California (#C49260) and Alaska (#10969), was responsible for information relating to the site geotechnical investigations, and design and costing of the Tailings Storage Facility and water management systems as described in Section 18.2 (excluding Section 18.2.2), Section 18.6, Section 18.7 and Section 18.13.5.



Robert Cummings, M.Sc. Geol. Eng., Registered Professional Engineer in Arizona, Geotechnical Consultant and Principal of Saguaro Geoservices, was responsible for development of pit slope stability analyses and design parameters in Section 18.1.3 and Section 18.2.2.

Other Expert contributors: 

Carol Fries, Vice President of Health, Safety, Environment and Community Relations for Norsemont Mining Inc., responsible for the preparation of the Environmental Social Impact Assessment and the preparation of information provided in Sections 5 and 18.

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

Gaston Loyola, MAusIMM, Vice President Exploration for Norsemont Mining Inc. is responsible for Norsemont’s exploration, drilling, sampling and data quality activities as described in Sections 7 to 13, and for Norsemont’s geological modelling (Section 17).



Christopher Gilmour, B.Eng (Civil), R.P.E.Q. (7126), M.I.E.Aust, CP.Eng, Operational Manager for GRD Minproc Limited, Brisbane, was responsible for the preparation of the capital costs estimate as detailed in Section 18.13.



Alan Ekstrom, Dip.E.Eng, R.P.E.Q. (9013), M.I.E. Aust. Leading Electrical and Instrumentation Engineer for GRD Minproc Limited, responsible for electrical, instrumentation and controls engineering as described in Section 18.4.3 and Section 18.4.4.



Ross Oliver, B.Eng (Mining), Aus.I.M.M. 105137, employed by GRD Minproc Limited as Manager Mining and Geology, responsible for preparation of the pre-tax financial model (part of Section 18.16).



Jorge Picon financial consultant to Norsemont is responsible for preparation of the post-tax financial model (part of Section 18.16).



David Evans, M.App.Sc. University of New South Wales, Australia, Principal Hydrogeologist for MWH Peru, was responsible for hydrogeological components of the environmental impact assessment and feasibility study.



Jean Cho, Ph.D. Princeton. P.Eng. British Columbia, Canada (Registration number 24635), was responsible for the numerical hydrogeological modeling and impact analysis.



Luis Yafac, B.Sc. Eng., Univ. National de Ingeniera (Mining), Project Manager of SIGT S.A., a Columbian consulting company which provides services in design and supervision of roads and highways, supervised the design and costing for the access road to the project (Section 18.5.1).



Mario Rojas Garay, B.Sc. Eng Universidad Nacional de Ingeniería, Engineer in Projects and Power Generation and Transmission for Cesel Ingenieros S.A., undertook design and costing for the Constancia project sub-stations in Section 18.5.4.



Pablo Lozano Ramon Chavez, B.Sc. Eng Universidad Nacional de Ingeniería, electrical engineer and Head of studies for Cesel Ingenieros S.A., was responsible for the design and costing of the HT power line in Section 18.5.4.

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Table of Contents Important notice ............................................................................................................ i 1. 1.1 1.2

1.3

1.4

1.5

1.6

1.7

SUMMARY............................................................................................................. 19 INTRODUCTION...............................................................................................................................19 GEOLOGY AND MINERAL RESOURCES.......................................................................................21 1.2.1 Geological setting, mineralisation and alteration...............................................................21 1.2.2 Exploration.........................................................................................................................22 1.2.3 Mineral resource estimation ..............................................................................................23 1.2.4 Mineral resources ..............................................................................................................24 MINING .............................................................................................................................................25 1.3.1 Introduction ........................................................................................................................25 1.3.2 Pit optimisation and pit design...........................................................................................25 1.3.3 Mineral reserve ..................................................................................................................26 1.3.4 Mine and process schedules .............................................................................................27 1.3.5 Mine fleet assessment .......................................................................................................27 1.3.6 Mine capital costs ..............................................................................................................28 1.3.7 Mine operating cost ...........................................................................................................31 GEOTECHNICAL INVESTIGATIONS ..............................................................................................31 1.4.1 Site investigations..............................................................................................................31 1.4.2 Pit geotechnical design......................................................................................................32 1.4.3 Subsurface conditions .......................................................................................................35 1.4.4 Construction materials .......................................................................................................35 1.4.5 Seismic conditions .............................................................................................................35 METALLURGICAL TESTWORK.......................................................................................................36 1.5.1 Testwork sample selection ................................................................................................36 1.5.2 Comminution testwork .......................................................................................................36 1.5.3 Flotation testwork...............................................................................................................37 PROCESS DESCRIPTION AND PLANT DESIGN...........................................................................38 1.6.1 Throughput ........................................................................................................................40 1.6.2 Crushing ............................................................................................................................41 1.6.3 Grinding .............................................................................................................................41 1.6.4 Copper flotation .................................................................................................................42 1.6.5 Molybdenum flotation.........................................................................................................42 1.6.6 Copper thickening and filtration .........................................................................................42 1.6.7 Molybdenum thickening and filtration ................................................................................42 1.6.8 Tailing thickening ...............................................................................................................43 1.6.9 Tailing water reclaim..........................................................................................................43 1.6.10 Concentrate storage and loadout ......................................................................................43 1.6.11 Water services ...................................................................................................................43 1.6.12 Reagents ...........................................................................................................................43 1.6.13 On stream analysis and laboratory....................................................................................43 WASTE MANAGEMENT...................................................................................................................44 1.7.1 Waste rock facility (WRF) ..................................................................................................44 1.7.2 Tailings management facility (TMF) ..................................................................................45

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1.8

1.9

1.10

1.11

1.12 1.13

1.14

1.15 1.16

1.17

1.7.3 Topsoil and unsuitable material stockpiles ........................................................................46 INFRASTRUCTURE .........................................................................................................................46 1.8.1 Access roads .....................................................................................................................46 1.8.2 Water supply......................................................................................................................47 1.8.3 Power supply .....................................................................................................................48 1.8.4 Internal roads.....................................................................................................................50 1.8.5 Buildings ............................................................................................................................50 1.8.6 Construction and accommodation camp ...........................................................................51 1.8.7 Concentrate transport and shipping ..................................................................................51 WATER MANAGEMENT ..................................................................................................................51 1.9.1 Process water ....................................................................................................................52 1.9.2 Water balance....................................................................................................................53 1.9.3 Non-process water.............................................................................................................53 ENVIRONMENTAL AND SOCIAL CONSIDERATIONS...................................................................54 1.10.1 ESIA...................................................................................................................................54 1.10.2 Stakeholder mapping.........................................................................................................55 1.10.3 Status of land ownership in the communities of Uchucarco and Chilloroya......................56 1.10.4 Impact identification and evaluation ..................................................................................56 1.10.5 Environmental management plan......................................................................................56 1.10.6 Resettlement Action Plan (RAP) .......................................................................................57 1.10.7 Community relations plan ..................................................................................................57 1.10.8 Health, safety and environment (HSE) management and monitoring plan .......................58 1.10.9 Closure plan.......................................................................................................................59 PROJECT IMPLEMENTATION PLAN..............................................................................................60 1.11.1 Approach and strategy.......................................................................................................60 1.11.2 Quality assurance ..............................................................................................................61 1.11.3 Project implementation schedule.......................................................................................61 1.11.4 Civil construction fleet........................................................................................................63 PROJECT OPERATIONAL PLAN ....................................................................................................64 CAPITAL COST ESTIMATE .............................................................................................................65 1.13.1 Initial project capital ...........................................................................................................65 1.13.2 Sustaining capital...............................................................................................................66 OPERATING COST ESTIMATES ....................................................................................................66 1.14.1 Mine operating cost ...........................................................................................................67 1.14.2 Plant and infrastructure costs ............................................................................................67 1.14.3 General and administration ...............................................................................................68 1.14.4 Off-site charges .................................................................................................................68 1.14.5 Royalties ............................................................................................................................69 MARKETING, PRODUCT PRICING AND TREATMENT CHARGES ..............................................69 PROJECT FINANCIAL ANALYSIS...................................................................................................70 1.16.1 Background........................................................................................................................70 1.16.2 Summary ...........................................................................................................................73 1.16.3 Pre-tax analysis .................................................................................................................73 1.16.4 Post-tax analysis................................................................................................................74 CONCLUSIONS................................................................................................................................77 1.17.1 Project overview ................................................................................................................77

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1.18

1.17.2 Risks and opportunities .....................................................................................................77 RECOMMENDATIONS.....................................................................................................................79 1.18.1 Mineral resources ..............................................................................................................79 1.18.2 Mining and mineral reserves .............................................................................................79 1.18.3 Geotechnical and hydrogeological studies ........................................................................79 1.18.4 Process testwork and plant design....................................................................................80 1.18.5 Environmental and permitting............................................................................................80

2. 2.1 2.2 2.3 2.4 2.5 2.6

INTRODUCTION.................................................................................................... 81 BACKGROUND ................................................................................................................................81 SCOPE OF WORK ...........................................................................................................................81 SOURCES OF INFORMATION ........................................................................................................82 SITE INSPECTIONS.........................................................................................................................82 CONTRIBUTORS TO THE REPORT ...............................................................................................83 DISCLOSURE OF INTEREST ..........................................................................................................85

3.

RELIANCE ON OTHER EXPERTS .......................................................................... 86

4. 4.1 4.2 4.3 4.4

PROPERTY DESCRIPTION AND LOCATION ......................................................... 87 GENERAL LOCATION......................................................................................................................87 PERUVIAN MINING LAW .................................................................................................................87 CONSTANCIA MINING CONCESSIONS.........................................................................................88 MINERAL RIGHTS OWNERSHIP ....................................................................................................91 4.4.1 Rio Tinto purchase.............................................................................................................91 4.4.2 Mitsui Mining and smelting purchase ................................................................................93 SURFACE RIGHTS ..........................................................................................................................93 ENVIRONMENTAL REGULATIONS ................................................................................................95

4.5 4.6 5. 5.1 5.2 5.3 5.4

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ................................................................................................... 97 ACCESSIBILITY ...............................................................................................................................97 CLIMATE AND VEGETATION..........................................................................................................97 LOCAL RESOURCES AND INFRASTRUCTURE............................................................................98 GEOMORPHOLOGY ........................................................................................................................99

6.

HISTORY............................................................................................................. 100

7. 7.1 7.2

GEOLOGICAL SETTING ...................................................................................... 101 DISTRICT GEOLOGY.....................................................................................................................101 PROPERTY GEOLOGY .................................................................................................................102 7.2.1 Stratigraphy .....................................................................................................................103 7.2.2 Intrusions .........................................................................................................................104 7.2.3 Alteration..........................................................................................................................105 7.2.4 Structural geology............................................................................................................106

8.

DEPOSIT TYPES ................................................................................................. 108

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9. 9.1 9.2

MINERALISATION AND ALTERATION ................................................................ 109 CONSTANCIA.................................................................................................................................109 SAN JOSÉ ......................................................................................................................................110

10. 10.1 10.2 10.3

EXPLORATION.................................................................................................... 111 SURFACE MAPPING AND SAMPLING .........................................................................................111 GEOPHYSICS ................................................................................................................................111 EXPLORATORY DRILLING ...........................................................................................................117

11.

DRILLING ............................................................................................................ 119 11.1.1 Collar location ..................................................................................................................120 11.1.2 Rig setup..........................................................................................................................121 11.1.3 Downhole survey .............................................................................................................121

12. 12.1

12.4

SAMPLING METHODS AND APPROACH ............................................................. 122 DRILLHOLE SAMPLING METHODS NORSEMONT 2007-2008...................................................122 12.1.1 Sample collection.............................................................................................................122 12.1.2 Drillhole logging ...............................................................................................................123 12.1.3 Density measurements ....................................................................................................123 12.1.4 Sample preparation, analysis and security......................................................................123 CORE SAMPLING ..........................................................................................................................124 QAQC PROCEDURES ...................................................................................................................124 12.3.1 Field duplicates................................................................................................................124 12.3.2 Blanks ..............................................................................................................................124 12.3.3 Standards ........................................................................................................................125 12.3.4 Other QAQC samples......................................................................................................125 12.3.5 Referee laboratory ...........................................................................................................126 STATEMENT ON SAMPLE PREPARATION AND ANALYSIS ......................................................126

13.

SAMPLE PREPARATION, ANALYSES AND SECURITY ....................................... 127

14. 14.1

DATA VERIFICATION.......................................................................................... 128 NORSEMONT INTERNAL DATA VERIFICATION.........................................................................128 14.1.1 Collar location ..................................................................................................................128 14.1.2 Downhole survey .............................................................................................................128 14.1.3 QAQC data verification ....................................................................................................128 14.1.4 Database generation and validation ................................................................................128 GRD MINPROC DATA VERIFICATION .........................................................................................129 14.2.1 Drilling ..............................................................................................................................129 14.2.2 Sampling..........................................................................................................................129 14.2.3 QAQC data verification ....................................................................................................129 14.2.4 Collar location ..................................................................................................................138 14.2.5 Downhole survey .............................................................................................................138 14.2.6 Sample database integrity ...............................................................................................139 14.2.7 Independent samples ......................................................................................................139 14.2.8 Core recovery considerations..........................................................................................139

12.2 12.3

14.2

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15.

ADJACENT PROPERTIES.................................................................................... 141

16. 16.1

MINERAL PROCESSING AND METALLURGICAL TESTING ................................ 142 PRE-DFS TEST WORK ..................................................................................................................142 16.1.1 Iditec-Chile.......................................................................................................................142 16.1.2 Plenge Laboratories.........................................................................................................142 16.1.3 SGS Chile test work.........................................................................................................145 16.1.4 SGS Canada flotation variability tests .............................................................................146 DFS METALLURGICAL TESTWORK ............................................................................................148 16.2.1 Mineralogy .......................................................................................................................149 16.2.2 Comminution....................................................................................................................150 16.2.3 Flotation ...........................................................................................................................152 PROCESS DESIGN CRITERIA......................................................................................................164 16.3.1 Grinding ...........................................................................................................................164 16.3.2 Copper flotation ...............................................................................................................165 16.3.3 Molybdenum flotation.......................................................................................................166 16.3.4 Copper thickening and filtration .......................................................................................168 16.3.5 Molybdenum thickening and filtration ..............................................................................168 16.3.6 Tailings thickening ...........................................................................................................168 16.3.7 Concentrate storage ........................................................................................................169 16.3.8 Water services .................................................................................................................169 16.3.9 Reagents .........................................................................................................................170 PROCESS PLANT DESCRIPTION ................................................................................................170 16.4.1 Overview..........................................................................................................................170

16.2

16.3

16.4

17. 17.1 17.2 17.3 17.4

17.5

17.6 17.7

17.8

MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ............................ 172 INTRODUCTION.............................................................................................................................172 DATA PROVIDED TO GRD MINPROC..........................................................................................172 DATA PREPARATION....................................................................................................................173 SURFACE AND SOLID WIREFRAME DATA GENERATION........................................................174 17.4.1 Lithology model................................................................................................................174 17.4.2 Mineralisation domain model...........................................................................................180 17.4.3 Copper grade shell model................................................................................................184 17.4.4 Zinc grade shell model.....................................................................................................186 17.4.5 Skarn ore type domain.....................................................................................................188 17.4.6 Topography......................................................................................................................196 SAMPLE CODING ..........................................................................................................................198 17.5.1 Flagging by lithology domain ...........................................................................................198 17.5.2 Flagging by mineralisation domain ..................................................................................198 17.5.3 Flagging by copper and zinc grade shells .......................................................................198 DATA COMPOSITING ....................................................................................................................198 STATISTICAL ANALYSIS AND VARIOGRAPHY ..........................................................................200 17.7.1 Statistical analysis by domain..........................................................................................200 17.7.2 Outlier analysis (capping) ................................................................................................200 17.7.3 Variography .....................................................................................................................207 BLOCK MODEL CONSTRUCTION ................................................................................................210

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17.9

17.10 17.11 17.12 17.13 17.14 18. 18.1

18.2

18.3

18.4

17.8.1 Preparation ......................................................................................................................210 17.8.2 Lithological modelling (LITH) ...........................................................................................211 17.8.3 Mineralisation modelling (ZONE).....................................................................................212 17.8.4 Cu grade shell modelling (02CU) ....................................................................................214 17.8.5 Zn grade shell modelling (ZNZONE) ...............................................................................214 17.8.6 Sub-model consolidation .................................................................................................214 GRADE ESTIMATION ....................................................................................................................214 17.9.1 Domain control on estimation ..........................................................................................215 17.9.2 Search strategy................................................................................................................215 17.9.3 Unfolding..........................................................................................................................216 17.9.4 High grade capping..........................................................................................................217 17.9.5 Other parameters.............................................................................................................217 DENSITY ASSIGNMENT................................................................................................................217 RESOURCE CLASSIFICATION .....................................................................................................218 MODEL VALIDATION .....................................................................................................................222 MINERAL RESOURCE REPORTING ............................................................................................228 MINERAL RESERVES....................................................................................................................236 OTHER RELEVANT DATA AND INFORMATION .................................................. 237 MINING STUDIES...........................................................................................................................237 18.1.1 Introduction ......................................................................................................................237 18.1.2 Pit optimisation ................................................................................................................238 18.1.3 Pit design .........................................................................................................................243 18.1.4 Mineral Reserve...............................................................................................................251 18.1.5 Mine and process schedules ...........................................................................................253 18.1.6 Ore types and batching....................................................................................................258 18.1.7 Mine fleet assessment .....................................................................................................259 18.1.8 Mine operating cost .........................................................................................................262 18.1.9 Mine capital cost ..............................................................................................................265 GEOTECHNICAL STUDIES ...........................................................................................................268 18.2.1 Introduction ......................................................................................................................268 18.2.2 Open pit geotechnical investigations, design parameters and slope angles...................273 18.2.3 Geotechnical investigations at the TMF, PAG WRF and plant site.................................281 18.2.4 Borrow materials..............................................................................................................283 18.2.5 Natural hazards and slope stability..................................................................................283 18.2.6 Seismic risk analysis........................................................................................................284 HYDROGEOLOGICAL STUDIES...................................................................................................285 18.3.1 Introduction ......................................................................................................................285 18.3.2 Hydrogeological investigations and results .....................................................................286 18.3.3 Numerical seepage modelling and affects analysis ........................................................288 18.3.4 Open pit seepage and dewatering...................................................................................289 18.3.5 Tailings management facility seepage ............................................................................290 18.3.6 Waste rock facility seepage.............................................................................................290 PROCESS PLANT DESIGN ...........................................................................................................291 18.4.1 General layout and description........................................................................................291 18.4.2 Facility description ...........................................................................................................292

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18.5

18.6

18.7

18.8

18.9

18.10

18.11 18.12

18.13

18.4.3 Electrical ..........................................................................................................................303 18.4.4 Instrumentation and control .............................................................................................309 18.4.5 Buildings ..........................................................................................................................310 INFRASTRUCTURE .......................................................................................................................314 18.5.1 Access road .....................................................................................................................314 18.5.2 Water supply....................................................................................................................320 18.5.3 Accommodation camp .....................................................................................................321 18.5.4 Power supply ...................................................................................................................322 WATER MANAGEMENT ................................................................................................................326 18.6.1 Introduction ......................................................................................................................326 18.6.2 Process water ..................................................................................................................328 18.6.3 Water balance..................................................................................................................329 18.6.4 Non-process water...........................................................................................................331 WASTE MANAGEMENT.................................................................................................................332 18.7.1 Waste rock facility............................................................................................................332 18.7.2 Tailings management facility ...........................................................................................334 18.7.3 Topsoil and unsuitable material stockpiles ......................................................................339 PORT AND TRANSPORT ..............................................................................................................339 18.8.1 Introduction ......................................................................................................................339 18.8.2 Management of concentrates ..........................................................................................340 18.8.3 Transportation..................................................................................................................340 18.8.4 Port management ............................................................................................................340 PROJECT IMPLEMENTATION PLAN............................................................................................341 18.9.1 Approach and strategy.....................................................................................................341 18.9.2 Quality assurance ............................................................................................................343 18.9.3 Project construction fleet .................................................................................................344 18.9.4 Project implementation schedule.....................................................................................346 18.9.5 Permits and licences........................................................................................................348 PROJECT OCCUPATIONAL HEALTH, SAFETY, ENVIRONMENT AND SECURITY..................349 18.10.1 Occupational health and safety .......................................................................................349 18.10.2 Security ............................................................................................................................349 PROJECT OPERATIONAL PLAN ..................................................................................................350 ENVIRONMENTAL CONSIDERATIONS........................................................................................351 18.12.1 Legal framework ..............................................................................................................351 18.12.2 Environmental impact assessment..................................................................................352 18.12.3 HSE management and monitoring plan...........................................................................363 18.12.4 Closure plan.....................................................................................................................364 CAPITAL COST ..............................................................................................................................367 18.13.1 Summary .........................................................................................................................367 18.13.2 Mining capital cost estimate ............................................................................................369 18.13.3 Process plant and associated infrastructure capital cost estimate..................................369 18.13.4 Indirect cost estimate (EPCM).........................................................................................370 18.13.5 Waste management facilities and water infrastructure capital cost estimate..................372 18.13.6 Infrastructure capital cost estimate..................................................................................372 18.13.7 Owners civil construction costs .......................................................................................374 18.13.8 Owners cost estimate ......................................................................................................375

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18.17

18.13.9 Contingency.....................................................................................................................375 OPERATING COSTS......................................................................................................................376 18.14.1 Summary .........................................................................................................................376 18.14.2 Mining - operating cost estimate......................................................................................377 18.14.3 Process plant and associated infrastructure - operating cost estimate...........................378 18.14.4 General and administration - operating cost estimate.....................................................381 18.14.5 Off-site operating cost estimate.......................................................................................384 18.14.6 Royalty – cost estimate....................................................................................................384 MARKETING, TREATMENT CHARGES AND PRODUCT PRICING ............................................385 PROJECT FINANCIAL ANALYSIS.................................................................................................388 18.16.1 Summary .........................................................................................................................388 18.16.2 Key project assumptions..................................................................................................388 18.16.3 Pre-tax analysis ...............................................................................................................392 18.16.4 Post-tax Analysis .............................................................................................................398 RISK ASSESSMENT ......................................................................................................................402

19.

INTERPRETATION AND CONCLUSIONS............................................................. 406

20. 20.1 20.2 20.3 20.4 20.5 20.6 20.7

RECOMMENDATIONS FOR FURTHER WORK ..................................................... 409 RESOURCES .................................................................................................................................409 MINING ...........................................................................................................................................409 GEOTECHNICAL AND HYDROGEOLOGICAL STUDIES.............................................................409 METALLURGICAL TESTWORK.....................................................................................................410 INFRASTRUCTURE .......................................................................................................................411 ESIA AND PERMITTING ................................................................................................................411 PROJECT IMPLEMENTATION ......................................................................................................411

21.

REFERENCES...................................................................................................... 412

22.

DATE AND SIGNATURE PAGE ............................................................................ 413

23.

ADDITONAL REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION AND DEVELOPMENT PROPERTIES .................................................................... 415

24.

ILLUSTRATIONS................................................................................................. 416

25.

ANNEXURES ....................................................................................................... 417

18.14

18.15 18.16

List of Attachments Appendix 1 Permits Appendix 2 Picon & Asociados Tax Opinion

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List of Tables Table 1.1 Constancia Project Global Mineral Resource Estimate (0.25% Cu Cut-off)...........................24 Table 1.2 Constancia Project Global Mineral Resource Estimate (0.20% Cu Cut-off)...........................24 Table 1.3 Constancia Project Global Mineral Resource Estimate (0.15% Cu Cut-off)...........................25 Table 1.4 Constancia Project Mineral Reserves....................................................................................26 Table 1.5 Equipment Fleet, Capital and Hourly Operating Costs ..........................................................28 Table 1.6 Mine Capital Costs by Component ........................................................................................29 Table 1.7 Mine Capital, Sustaining and Replacement Costs ($ M)........................................................30 Table 1.8 Expected Elemental Recoveries ............................................................................................38 Table 1.9 Theoretical Maximum Throughput For Each Ore Type..........................................................40 Table 1.10 Plant Design Throughput Rate ............................................................................................41 Table 1.11 Civil Construction Fleet........................................................................................................64 Table 1.12 Constancia Project Capital Cost Estimate ...........................................................................65 Table 1.13 Summary Operating Cost Estimate .....................................................................................66 Table 1.14 Mine Operating Costs ..........................................................................................................67 Table 1.15 Plant and Associated Infrastructure Operating Cost Estimate .............................................68 Table 1.16 G&A Operating Cost Estimate .............................................................................................68 Table 1.17 Off-site Operating Costs ......................................................................................................69 Table 1.18 Royalty Costs.......................................................................................................................69 Table 1.19 Operating Cost Inputs to Financial Model............................................................................70 Table 1.20 Metal Price Assumptions in Cashflow Model .......................................................................71 Table 1.21 Production Schedule for Cashflow Model ............................................................................72 Table 1.22 Constancia Project After Tax Analysis Summary.................................................................73 Table 1.23 Tax and Depreciation Rates ................................................................................................75 Table 1.24 Post-tax Project Sensitivity Analysis ....................................................................................76 Table 2.1 Contributors to Constancia DFS Technical Report ................................................................84 Table 4.1 Constancia Concessions .......................................................................................................90 Table 4.2 Constancia Concession Grants .............................................................................................91 Table 4.3 Option Exercise Schedule .....................................................................................................92 Table 4.4 Private Lands Summary ........................................................................................................93 Table 5.1 Travel Distance and Time to Constancia ...............................................................................97 Table 11.1 Drilling Programmes by Year (in metres drilled) ................................................................119 Table 11.2 Drilling Programmes (to 6 June 2009) ...............................................................................120 Table 12.1 Recommended Values for Standards ................................................................................125 Table 16.1 Iditec-Chile Flotation Recoveries .......................................................................................142 Table 16.2 Plenge Laboratories Phase 1 Locked Cycle Test Results – Hypogene Ore ......................143 Table 16.3 Plenge Laboratories Phase 1 Supergene Locked Cycle Test Results...............................143 Table 16.4 Plenge Laboratories Phase 1 Locked Cycle Test Results – Skarn Ore .............................143 Table 16.5 Plenge Laboratories Phase 2 Locked Cycle Test Results (July 4, 2008)...........................144 Table 16.6 Plenge Laboratories Phase 2 Selected Locked Cycle Test Results ..................................145 Table 16.7 Historical SGS Chile Comminution Test Work ...................................................................145 Table 16.8 SGS Canada Rougher Flotation Variability Testing Conditions .........................................146 Table 16.9 SGS Canada Rougher Flotation Variability Testing Results – Supergene Ore..................146 Table 16.10 SGS Canada Rougher Flotation Variability Testing Results – Hypogene Ore.................147 Table 16.11 SGS Canada Rougher Flotation Variability Testing Results – Skarn Ore........................147

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Table 16.12 Ore Zone Samples...........................................................................................................148 Table 16.13 Collectors Evaluated in Supergene Rougher Flotation ....................................................152 Table 16.14 Conditions for Supergene Locked Cycle Tests ................................................................153 Table 16.15 Results from Variability Testing on Supergene Ore Samples ..........................................154 Table 16.16 Test Conditions for the Hypogene Locked Cycle Test Work............................................155 Table 16.17 Results From Hypogene Locked Cycle Tests..................................................................155 Table 16.18 Results from Variability Testing on Hypogene Ore Samples ...........................................156 Table 16.19 Deep Hypogene Samples Tested at 75 µm and 106 µm .................................................156 Table 16.20 Conditions Used in Locked Cycle Test Work on Medium-Zinc Skarn Ore .......................157 Table 16.21 Results from Medium-Zinc Locked Cycle Tests...............................................................158 Table 16.22 Simulated Molybdenum Grade/Recovery as a Function of Cleaning Stages...................159 Table 16.23 Anticipated Molybdenum Concentrate .............................................................................160 Table 16.24 Final Molybdenum Product Elemental Composition ........................................................161 Table 16.25 Pressure Filtration Results...............................................................................................164 Table 16.26 Summary of Design Ore Parameters – Grinding .............................................................165 Table 16.27 Summary of Copper Flotation Design Criteria .................................................................165 Table 16.28 Molybdenum Flotation Recovery and Concentrate Grade ...............................................166 Table 16.29 Summary of Molybdenum Flotation Design Criteria.........................................................167 Table 16.30 Copper Concentrate Thickening and Filtration Specification ...........................................168 Table 16.31 Molybdenum Concentrate Thickening and Filtration Specification...................................168 Table 16.32 Tailings Thickening Specification.....................................................................................169 Table 16.33 Concentrate Storage Specification ..................................................................................169 Table 16.34 Water Services Specification ...........................................................................................169 Table 16.35 Reagent Specification Summary......................................................................................170 Table 17.1 Quantities of SK2 Skarn in Model ......................................................................................195 Table 17.2 Variogram Model Parameters ............................................................................................208 Table 17.3 Model Prototype Parameters .............................................................................................210 Table 17.4 Search Parameters............................................................................................................216 Table 17.5 Density Assignment ...........................................................................................................218 Table 17.6 Resource Classification Criteria.........................................................................................219 Table 17.7 Statistical Comparison between Means of Composite and Model Data – Lithology Model............................................................................................................................223 Table 17.8 Statistical Comparison between Means of Composite and Model Data – Lithology Model............................................................................................................................224 Table 17.9 Statistical Comparison between Means of Composite and Model Data – Mineral Zonation Model.............................................................................................................224 Table 17.10 Constancia Project Global Mineral Resource Estimate 0.25% Cu Cut-off .......................229 Table 17.11 Constancia Project Global Mineral Resource Estimate 0.20% Cu Cut-off .......................229 Table 17.12 Constancia Project Global Mineral Resource Estimate at Various Cu Cut-off .................230 Table 17.13 Constancia Project Global Mineral Resource Estimate – Constancia Main Orebody ......231 Table 17.14 Constancia Project Global Mineral Resource Estimate – San José Orebody..................232 Table 17.15 Constancia Project Global Mineral Resource Estimate – Domain MP1...........................233 Table 17.16 Constancia Project Global Mineral Resource Estimate – Domain Skarn.........................234 Table 17.17 Constancia Project Global Mineral Resource Estimate – Domain Supergene.................235 Table 17.18 Constancia Project Global Mineral Reserve Estimate .....................................................236 Table 18.1 Pit Optimisation Parameters - NSR Calculation.................................................................239

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Table 18.2 Optimisation Parameters - Operating Costs and Throughput Rates..................................240 Table 18.3 Pit Optimisation Parameters - Concentrator Recoveries ...................................................241 Table 18.4 Pit Optimisation Parameters - Concentrate Grades...........................................................242 Table 18.5 Pit Design Inventories........................................................................................................247 Table 18.6 Constancia Project Global Mineral Reserve Estimate .......................................................251 Table 18.7 Constancia Project Mineral Reserve by Pit Stage .............................................................251 Table 18.8 Constancia Project Mineral Reserve by Ore Type .............................................................252 Table 18.9 Constancia Project Mineral Reserve by Operating Margin ................................................252 Table 18.10 Low Margin Mineralisation excluded from the Mineral Reserve.......................................253 Table 18.11 Annual Mining and Processing Schedule ........................................................................256 Table 18.12 Key Operating Cost Inputs...............................................................................................262 Table 18.13 Equipment Fleet and Hourly Costs ..................................................................................263 Table 18.14 Mining - Capital, Sustaining and Replacement Costs ($ M).............................................266 Table 18.15 Mine Capital Costs by Component ..................................................................................267 Table 18.16 Equipment Unit Capital Cost - Mining Fleet .....................................................................268 Table 18.17 Major Structures System in Constancia Pit Area .............................................................275 Table 18.18 Inter-ramp Slope Angles – Constancia Open Pit .............................................................280 Table 18.19 Construction Materials Summary.....................................................................................283 Table 18.20 Key Hydrogeologic Issues and Mitigative Measures........................................................285 Table 18.21 Design Return Periods for Hydraulic Structures ..............................................................317 Table 18.22 Access Road Summary of Capital Costs .........................................................................320 Table 18.23 Power Supply Capital Cost Estimate ...............................................................................325 Table 18.24 Construction Fleet Requirements ....................................................................................346 Table 18.25 Disturbance Areas ...........................................................................................................365 Table 18.26 Reclamation Cost Estimate Summary .............................................................................367 Table 18.27 Capital Cost Estimate Summary ......................................................................................368 Table 18.28 Mining Capital Cost Estimate Summary ..........................................................................369 Table 18.29 Plant Capital Cost Estimate Summary.............................................................................369 Table 18.30 Process Plant Area Summary Capital Cost .....................................................................371 Table 18.31 Waste and Water Infrastructure Capital Cost Estimate Summary ...................................372 Table 18.32 Access Road Capital Cost Estimate Summary ................................................................372 Table 18.33 Accommodation Camp Capital Cost Estimate Summary.................................................373 Table 18.34 HV Power Supply Capital Cost Estimate Summary .........................................................373 Table 18.35 Owners Civil Construction Fleet Capital Cost Estimate Summary ...................................374 Table 18.36 Owners Cost Estimate Summary.....................................................................................375 Table 18.37 Operating Cost Estimate Summary .................................................................................377 Table 18.38 Mining Operation Cost Estimate ......................................................................................378 Table 18.39 Process Plant Operating Cost Estimate...........................................................................379 Table 18.40 General and Administration Operating Cost Estimate Summary .....................................381 Table 18.41 General and Administration Operating Cost Estimate .....................................................383 Table 18.42 Transport and Off-site Treatment Charges ......................................................................384 Table 18.43 Off-site Operating Cost Estimate .....................................................................................384 Table 18.44 Royalty Cost Estimate Summary .....................................................................................385 Table 18.45 Prices and Charges Used in Constancia DFS .................................................................387 Table 18.46 Project After Tax Analysis................................................................................................388 Table 18.47 Production Schedule........................................................................................................389

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Table 18.48 Table 18.49 Table 18.50 Table 18.51 Table 18.52 Table 18.53

Operating Cost Summary ................................................................................................390 Key Project Metrics..........................................................................................................395 Project Sensitivity Analysis (pre-tax)................................................................................396 Tax and Depreciation Assumptions .................................................................................398 Key Post-tax Project Metrics............................................................................................399 Post Tax Project Sensitivity Analysis ...............................................................................400

List of Figures Figure 1.1 Constancia Project Location.................................................................................................19 Figure 1.2 Ultimate Pit – Plan View .......................................................................................................26 Figure 1.3 Annual Ore Production by Ore Type.....................................................................................27 Figure 1.4 Mine Operating Costs over Time ($/t)...................................................................................31 Figure 1.5 Overall Site Plan...................................................................................................................32 Figure 1.6 Proposed Design Sectors, Inter-ramp and Bench Slopes ....................................................34 Figure 1.7 Constancia Process Flowsheet ............................................................................................39 Figure 1.8 Constancia Plant Layout.......................................................................................................40 Figure 1.9 Project Implementation Schedule .........................................................................................62 Figure 1.10 Pre-tax Cashflow Sensitivity Analysis (NPV – 8% discount rate)........................................74 Figure 1.11 Post-tax NPV Sensitivity (8% discount rate).......................................................................77 Figure 4.1 Project Location....................................................................................................................87 Figure 4.2 Concession Boundaries........................................................................................................89 Figure 4.3 Surface Rights ......................................................................................................................94 Figure 7.1 Simplified Geology of the Andahuaylas-Yauri Area ............................................................102 Figure 7.2 Geological Map of the Constancia Deposit.........................................................................103 Figure 10.1 Constancia Project – Exploration Targets ........................................................................112 Figure 10.2 Pampacancha Geology Map ............................................................................................113 Figure 10.3 Pampacancha Prospect – Exploratory Drilling..................................................................114 Figure 10.4 Pampacancha Prospect – Main Drilling Intercepts ...........................................................115 Figure 10.5 Chilloroya South – Geological Map ..................................................................................116 Figure 14.1 Cu Blanks (%)...................................................................................................................130 Figure 14.2 Mo Blanks (ppm) ..............................................................................................................130 Figure 14.3 Ag Blanks (ppm) ...............................................................................................................131 Figure 14.4 Au Blanks (ppb) ................................................................................................................131 Figure 14.5 Pb Blanks (ppb) ................................................................................................................132 Figure 14.6 Zn Blanks (%) ...................................................................................................................132 Figure 14.7 Cu Standard MV700041 ...................................................................................................133 Figure 14.8 Cu Standard MV700040 ...................................................................................................134 Figure 14.9 Cu Standard MV700039 ...................................................................................................134 Figure 14.10 Cu Standard MV00038 ...................................................................................................135 Figure 14.11 Cu ICP Field Duplicate Correlation Plot..........................................................................135 Figure 14.12 High Grade Cu Field Duplicate Correlation Plot .............................................................136 Figure 14.13 Mo Field Duplicate Correlation Plot ................................................................................136 Figure 14.14 Ag Field Duplicate Correlation Plot.................................................................................137

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Figure 14.15 Zn Field Duplicate Correlation Plot .................................................................................137 Figure 14.16 CO-0501 Collar Monument.............................................................................................138 Figure 14.17 Core Recovery................................................................................................................140 Figure 14.18 Core Recovery Distribution in Mineralised Samples.......................................................140 Figure 16.1 Drop Weight (DWi) and Bond Work (BWi) Index Test Results for Supergene Ore...........150 Figure 16.2 Drop Weight (DWi) and Bond Work (BWi) Index Test Results for Hypogene Ore ............151 Figure 16.3 Drop Weight (DWi) and Bond Work (BWi) Index Test Results for Skarn Ore ...................152 Figure 16.4 Concentrate Settling Tests ...............................................................................................162 Figure 16.5 Tailing Settling Tests ........................................................................................................163 Figure 16.6 Predicted Plant Production Rates Years 1-16 ..................................................................171 Figure 16.7 Predicted Average Copper Recovery and Copper Concentrate Grade ............................171 Figure 17.1 Orthographic View Looking NE, Showing the Skarn and Marble Units.............................175 Figure 17.2 Plan View Showing the Constancia Skarn Geometry .......................................................176 Figure 17.3 Sectional View Showing the San José and Constancia Skarn Bodies .............................177 Figure 17.4 Plan View Showing the Geological Model for the Constancia Project ..............................178 Figure 17.5 Lithological Wireframes- Plan View ..................................................................................179 Figure 17.6 Lithological Model Looking North West ............................................................................180 Figure 17.7 Sectional View 100m Wide Looking NE ...........................................................................181 Figure 17.8 Plan View Showing the Surface Extent of Supergene Enrichment ...................................182 Figure 17.9 Mineralisation Domains - Plan View .................................................................................183 Figure 17.10 Mineralisation Domains, Looking Northwest...................................................................183 Figure 17.11 Plan View Showing the 0.2% Cu Grade Shell and Drill Strings ......................................185 Figure 17.12 Orthographic View Looking NE Showing the 0.2% Cu Grade Shell and Drill Strings .....185 Figure 17.13 Plan View Showing the 1500 ppm Zn Shell....................................................................186 Figure 17.14 Orthographic View Looking NE Showing the Relationship between Skarn Lithology (green) and the 1500 ppm Zn Shell (light blue) ............................................................187 Figure 17.15 Plan View Showing the Relationship between the Skarn Lithology (green) and the 1500 ppm Zn Shell (light blue)......................................................................................187 Figure 17.16 Orthographic View Looking North Showing the 1500 ppm Zn Shell ...............................188 Figure 17.17 Skarn Wireframe (blue) and the Supergene Wireframe (yellow) ....................................189 Figure 17.18 Skarn Wireframe (blue) and the Supergene Wireframe (yellow), with a better View of Overlap Areas ..........................................................................................................189 Figure 17.19 Plan View of Skarn and Supergene Wireframes.............................................................190 Figure 17.20 Supergene and Skarn, with High Zn Domains Highlighted.............................................191 Figure 17.21 Supergene and Skarn, with High Zn Domains Highlighted – Plan View .........................192 Figure 17.22 SK2 in Resource Model Blocks (highlighted in red)........................................................193 Figure 17.23 Plan View of SK2 in Resource Model Blocks (highlighted in red) ...................................193 Figure 17.24 SK2 Areas - 50x50x25 m Blocks ....................................................................................194 Figure 17.25 SK2 Areas - 100x100x40 m Blocks ................................................................................195 Figure 17.26 Topographic Contours and Features ..............................................................................196 Figure 17.27 Topographic Surface Plan ..............................................................................................197 Figure 17.28 Topographic Surface Model Looking Northwest .............................................................197 Figure 17.29 Histogram of Assay Interval Length................................................................................199 Figure 17.30 Histogram of Composite Interval Length ........................................................................199 Figure 17.31 Cu Log Probability Plot Overlay of Domains...................................................................201 Figure 17.32 Mo Log Probability Plot Overlay of Domains ..................................................................202

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Figure 17.33 Ag Log Probability Plot Overlay of Domains ...................................................................203 Figure 17.34 Au Log Probability Plot Overlay of Domains ...................................................................204 Figure 17.35 Zn Log Probability Plot Overlay of Domains ...................................................................205 Figure 17.36 Pb Log Probability Plot Overlay of Domains ...................................................................206 Figure 17.37 S Log Probability Plot Overlay of Domains.....................................................................207 Figure 17.38 Lithology Model – Plan View ..........................................................................................211 Figure 17.39 Lithology Model Northing Section ...................................................................................212 Figure 17.40 Mineral Zonation Model Plan ..........................................................................................213 Figure 17.41 Mineral Zonation Model Northing Section.......................................................................213 Figure 17.42 Data before Unfolding.....................................................................................................217 Figure 17.43 Data in Unfolded Space..................................................................................................217 Figure 17.44 Resource Classification Schematic Level 4300 mRL .....................................................220 Figure 17.45 Resource Classification Schematic Level 4200 mRL .....................................................220 Figure 17.46 Resource Classification Schematic Level 4100 mRL .....................................................221 Figure 17.47 Resource Classification Schematic Level 4000 mRL .....................................................221 Figure 17.48 Resource Classification Schematic - Transform Section ................................................222 Figure 17.49 Cu Grade Model Validation - Section 202050 East ........................................................226 Figure 17.50 Cu Grade Model Validation - Transform Section Cu Validation......................................227 Figure 17.51 Cu Grade Model Validation - Level Plan 4225 mRL .......................................................228 Figure 17.52 Constancia Resource Reporting by Confidence Classification .......................................231 Figure 18.1 Design Sectors and Proposed Inter-ramp and Bench Slopes ..........................................245 Figure 18.2 Ultimate Pit Design Plan View ..........................................................................................247 Figure 18.3 Constancia Stage 1 Design Plan View .............................................................................248 Figure 18.4 San José Pit Design Plan View ........................................................................................248 Figure 18.5 Constancia Stage2 Design Plan View ..............................................................................249 Figure 18.6 Constancia Stage 3 Pit Design Plan View ........................................................................249 Figure 18.7 Constancia Stage 4 Pit Design Plan View ........................................................................250 Figure 18.8 Constancia Ultimate and Staged Pit Designs Plan View ..................................................250 Figure 18.9 Mining by Pit Stage (Mt) ...................................................................................................254 Figure 18.10 Ore Mining by Material Type (Mt) ...................................................................................254 Figure 18.11 Float Ore Processing Plant Feed by NSR Range (Mt) ...................................................257 Figure 18.12 Closing Ore Stockpile Inventories (Mt) ...........................................................................257 Figure 18.13 Haul Truck Requirements...............................................................................................261 Figure 18.14 Operating Costs by Time ($/t).........................................................................................264 Figure 18.15 Mine Operating Costs Components................................................................................265 Figure 18.16 Geotechnical Site Investigation ......................................................................................270 Figure 18.17 General Geologic Map....................................................................................................271 Figure 18.18 Pit Structural Map ...........................................................................................................274 Figure 18.19 Design Sectors and Proposed Inter-ramp and Bench Slopes ........................................278 Figure 18.20 Relationship Between Inter-ramp Angle, Overall Angle, Bench Face Angle, and Catch Bench Width.......................................................................................................279 Figure 18.21 TMF Geotechnical Investigation and Geologic Map .......................................................281 Figure 18.22 Site Layout .....................................................................................................................292 Figure 18.23 Processing Plant Layout.................................................................................................293 Figure 18.24 Constancia Project Location and Access Road ..............................................................315 Figure 18.25 Typical Section of Upgraded Access road......................................................................316

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Figure 18.26 Figure 18.27 Figure 18.28 Figure 18.29 Figure 18.30 Figure 18.31 Figure 18.32 Figure 18.33 Figure 18.34 Figure 18.35 Figure 18.36 Figure 18.37 Figure 18.38 Figure 18.39 Figure 18.40 Figure 18.41 Figure 18.42 Figure 18.43 Figure 18.44 Figure 18.45 Figure 18.46 Figure 18.47 Figure 18.48 Figure 18.49

Water Management Schematic ......................................................................................327 General Water Balance Schematic.................................................................................330 PAG Waste Rock Facility Stage 5 (Ultimate) Loading Plan............................................333 Tailings Management Facility Final Configuration Plan ..................................................335 Tailings Management Facility – Typical Staged Embankment Sections.........................336 Delivery Model................................................................................................................342 Project Implementation Schedule ...................................................................................347 Wind Rose ......................................................................................................................354 Location of Water Quality Sampling Sites ......................................................................356 Capital Cost Estimate Probability Profile ........................................................................376 Year by Year Metallurgical Costs ...................................................................................380 Unit Process Operating Costs Per Tonne Milled ............................................................380 LME Long Term Copper Price Predictions (BH) .............................................................386 Copper Spot Price, July 2008-June 2009 .......................................................................386 Ramp up Recovery Assumptions ...................................................................................391 Plant Feed Grade ...........................................................................................................391 Payable Metal.................................................................................................................392 Estimated Cash Costs ....................................................................................................394 Pre-tax NPV Sensitivity (@8%) ......................................................................................397 Pre-tax IRR Sensitivity....................................................................................................397 Pre-tax NPV Sensitivity (@8%) ......................................................................................398 Post-tax NPV Sensitivity (@8%).....................................................................................401 Post-tax IRR Sensitivity ..................................................................................................401 Post-tax NPV Sensitivity (@8%).....................................................................................402

Page xviii


1.

SUMMARY

1.1

INTRODUCTION

Norsemont Mining Inc (Norsemont) is developing the Constancia Copper, Molybdenum, Silver Project in Southern Peru, approximately 100 km south of the city of Cusco (Figure 1.1). Figure 1.1 Constancia Project Location

GRD Minproc Limited (GRD Minproc) was appointed by Norsemont to undertake and manage a Definitive Feasibility Study (DFS) which has included a resource update, comprehensive metallurgical testwork, mine design, plant and infrastructure design and development of capital and operating costs. The DFS forms the basis of this Technical Report. The Constancia deposit is a large-scale porphyry deposit located 4100 metres above sea level (masl) in the Andes mountain range. Norsemont holds 100% rights to the mining concessions and surface rights covering the Constancia deposit, subject to a 3.5% net smelter return (NSR) royalty.

Page 19


The proposal is to develop a project comprising open pit mining and flotation of sulphide minerals, to produce commercial grade concentrates of copper and molybdenum. Silver and a small quantity of gold at payable levels will report to the copper concentrate. Annual production rates vary, but average 70 533 t/a copper metal and 54.5 t/a silver metal contained in the copper concentrate. Copper concentrate ramps up from 350 000 t/a in the first year to a peak of 450 000 t/a in Year 3. Production then drops to around 300 000 t/a until Year 10, after which it falls to 200 000 t/a and below until mine closure in Year 15. Molybdenum concentrate production ramps up from 2400 t/a in Year 1 to a peak of 4800 t/a in Year 3. It fluctuates between 2500 and 3000 t/a until another high is reached in Years 9 and 10, after which it drops to 2000 – 2500 t/a. The Project is largely self-contained, with mine, mill, maintenance facilities, administration and fully serviced accommodation camp located on the mine site. Supporting infrastructure includes grid supplied power from an upgraded supply point at Tintaya, 70 km away, and new transmission line from there to the mine. The public road to site will be upgraded to meet demands of extra traffic, particularly concentrate trucks and freight services. Raw water will be extracted from bores surrounding the open pit, and a tailings dam will be constructed within 5 km of the mine, on land owned freehold by Norsemont. The site has access by road to the port of Matarani via existing national roads through Arequipa, Yauri/Espinar, and Velille. Norsemont provided all the necessary drilling, sampling, analytical and geological data to GRD Minproc for Mineral Resource modelling and estimation purposes. Mine design, scheduling and costing for open pit operations were completed by GRD Minproc. Drill core samples were selected by GRD Minproc and sent to Lakefield SGS Laboratories in Chile for metallurgical testing. The results of this testwork were used by GRD Minproc to determine a process flowsheet and design criteria, and design the process plant. Geotechnical studies were completed for the process plant site by Knight Piésold Consulting (Knight Piésold). The same company carried out the design of the water and tailings management systems including the Tailings Management Facility (TMF), as well as the Potentially Acid Generating (PAG) Waste Rock Facility (WRF). Knight Piésold also provided geotechnical recommendations related to the pit walls while Saguaro Geoservices provided technical review and input on the pit walls. MWH Peru undertook hydrogeological investigations and modelling of the site, and provided pit dewatering information. Peruvian consultants CESEL and SIGT undertook design and costing for power supply and upgrading the access road to site, respectively. The accommodation camp for construction and operations personnel was based on estimates provided by experienced Peruvian suppliers.

Page 20


Capital cost estimate data for the plant was obtained from reputable equipment suppliers and Peruvian contractors, relying on technical specifications and material quantity takeoffs provided by GRD Minproc. Operating cost information and selected input parameters for the economic evaluation were obtained from market pricing and from Norsemont. GRD Minproc undertook pre-tax cashflow modeling, while post-tax financial analysis was completed by Norsemont. Note that all dollars referenced in this report are United States dollars. 1.2

GEOLOGY AND MINERAL RESOURCES

1.2.1

Geological setting, mineralisation and alteration

The Constancia deposit is a porphyry copper-molybdenum system which includes copper-bearing skarn mineralisation. Multiple phases of monzonite and monzonite porphyry have intruded a sequence of sandstones, mudstones and micritic limestone of Cretaceous age. The majority of the mineralisation is associated with potassic alteration and quartz veining, occurring as chalcopyrite-(bornite)-molybdenite-pyrite mineralisation in “A” and “B” type veinlets, and also replacing ferromagnesian minerals or filling fractures. Copper grades typically vary from 0.2% up to 4%, and are highest where fracture-filling style copper mineralisation is superimposed on earlier disseminated copper mineralisation. The high-grade hypogene copper mineralisation is hosted by a dense A-veinlet stockwork developed in an early porphyry phase. Pyrite/chalcopyrite ratio is typically low, being in the order of 1:1 to 2:1. Molybdenite commonly increases with depth, related to “B” veinlets. Bornite occurs sporadically especially at deeper levels, sometimes associated with some gold values. Propylitic alteration is peripheral to the potassic alteration and extends more than one kilometre from the porphyry intrusive contacts. The propylitic alteration mineral assemblage includes epidote-chloritecalcite-pyrite-rhodochrosite. Subordinate chalcopyrite is also present, filling fractures or replacing mafic minerals. Sphalerite-galena veinlets and veins are distributed as a halo to the copper-molybdenum mineralisation within the propylitic alteration halo, occurring at distances of up to 3 km away from the porphyry copper system. Phyllic alteration forms a pervasive carapace surrounding and sometimes overprinting potassic alteration. The phyllic alteration accompanies almost complete destruction of primary rock textures; the mineral assemblage includes sericite-quartz-pyrite, limited amounts of chalcopyrite and associated occasional “D” veins and veinlets. At the contact between the intrusives and limestones, magnetite ± garnet skarn develops, while a pyroxene–diopside (garnet–epidote) association is more common in calcareous sandstones and arkoses of the Chilloroya Formation. Skarn mineralisation is volumetrically much smaller, but grades are normally higher.

Page 21


Structural deformation has played a significant role in preparing and localising the hydrothermal alteration and copper-molybdenum-silver-gold mineralisation, including skarn formation. Major interand post-mineral fracture systems in the deposit area strike northeast, and include the Barite fault system. This is represented by a number of nearly parallel vein-faults carrying base metal sulphides and barite, which have been exploited by artisanal workings throughout the property. A second important system strikes north-south. It appears to be more recent than the Barite system, and controls part of the San José Pit mineralisation and most of the silicified breccias (sometimes mineralised) in the system. This is the same direction as that of the post-mineral dykes, and may have originated as tension gashes to the Barite direction. Oxide copper mineralisation occurs locally. While shallow, it is volumetrically small and, therefore, is not considered relevant to exploitation at this stage of Project development. Supergene enrichment occurs immediately beneath, and occasionally as remnants within, a leached cap. The highest copper grades in the Constancia porphyry are typically associated with this and with the skarn zone. Transitional (Mixed) mineralisation is present over a limited interval where the supergene and hypogene mineralisation co-exist. Two individual porphyry-style deposits are known within the project area, Constancia and San José, separated by some 350 m. The total mineralised zone extends about 1200 m in the north-south direction and 800 m in the east-west direction. Mineralisation occurs to surface at San José, but is deeper at Constancia. Glacial moraines cover the northern and eastern margins of the Constancia deposit: to the east these moraines cover potentially important extensions of copper mineralisation along broad east-west structural zones. Several additional exploration targets have been identified in the surrounding area, these being highlighted by Induced Polarisation (chargeability/resistivity) and ground magnetometry geophysical surveys. 1.2.2

Exploration

As of 18 June 2009, a total of 132 130.35 m (451 holes) have been drilled in the Constancia Project, including 7484.15 m drilled by Rio Tinto prior to 2005. The total also includes metallurgical, geotechnical and condemnation drilling programmes. Drilling comprises both diamond drilling and reverse circulation percussion drilling; diamond drilling constitutes 90% of the total. Exploration has been conducted to conventional industry standards, including surface and downhole surveying of drillholes, geological and geotechnical core logging, cutting and sampling of drill core, sample preparation and assaying. GRD Minproc has reviewed the methods used in the drill programs and considers them appropriate for a mineral resource estimate. A total of 1108 density measurements have been made for core from the Constancia-San José area. The density measurements were conducted by ALS Chemex, and are representative of the different rock and mineralisation domains.

Page 22


QAQC control for the assay data is based on inclusion of blank, standard and duplicate (1/4 core, coarse duplicate and pulp duplicate) samples with routine samples. A referee laboratory has been used to provide supporting analyses. Samples are securely stored before being loaded onto covered and secured trucks for transport to the laboratory in Lima. Chain of custody documents is maintained, with signatures of delivering and receiving parties and the names of persons accompanying the samples at all times. GRD Minproc is of the opinion that the Norsemont QAQC sampling protocol is rigorously set up and is continuously monitored to identify potential sampling and assaying problems. 1.2.3

Mineral resource estimation

Resource estimation for Constancia is based on integrated geological and assay interpretations of information recorded from diamond core logging and assaying. The database used was current as of 30 January 2009. Topographic and drillhole survey data, assay data and interpretative wireframes were supplied by Norsemont, including recently updated wireframes prepared by Atticus Associates, Lima using the Leapfrog (Version 2.1) software package. These wireframes comprised grade shells (0.2% Cu and 0.15% Zn), mineral zonation (hypogene, mixed, supergene, oxide and leached), lithological and alteration wireframes. The data was checked by GRD Minproc for internal consistency. Thereafter, assay data was desurveyed and coded by domain according to the lithology, mineral zonation and grade shell wireframes. Coding was checked rigorously against the wireframes. Assay data was composited into 2 m intervals. Statistical analysis of the composite data indicated differences between populations according to mineral zonation domains and 0.2% Cu grade shell, and these were maintained as hard boundaries for determination of top-cuts, geostatistical analysis and grade interpolation. Top-cuts were determined for Cu, Mo, Ag and Zn for each mineral zonation domain. Variographic analysis was undertaken for Cu, Mo, Ag, Au, Zn, Pb and S within each mineral domain inside the 0.2% Cu grade shell to determine directions of maximum grade continuity, and to model variogram parameters for grade interpolation. A fully coded 25x25x15 m (ExNxRL) block model was constructed representing all mineral zonation domains, lithological and grade shell subsets, by applying constraints using relevant surface and solid wireframes. Sub-celling to 2.5x5x3 m was allowed, to honour interpretative boundaries. Grade estimation was undertaken for Cu, Mo, Ag,Pb, Au, S and Zn by Ordinary Kriging using hard boundaries based on mineralisation type and further subdivided by inside/outside 0.2% Cu grade shell (and 0.15% Zn grade shell for Zn interpolation). A third layer of sub-domain occurs within the hypogene caused by the need to separate out the barren late-stage dykes as defined by the lithology domains.

Page 23


Kriging neighbourhood analysis was completed on the domain-coded Cu data in the Constancia zone to determine optimal parameters for grade estimation, i.e. block size, number of discretisation points, search ellipse dimensions, minimum and maximum sample numbers employed in a search plan. Inverse distance squared interpolation was also carried out for each element and domain, using the same search parameters as for the Ordinary Kriging estimate. This was used to aid in validation of the final resource model. Average density values were determined for each of the modelled lithological units, taking mineral zonation type into account, and applied to the block model. Resources were classified in line with the Canadian National Instrument 43-101, taking account of the quality and reliability of raw data, drillhole spacing, confidence in the geological interpretation, the number, spacing and orientation of intercepts through mineralised zones, and grade continuity information gained from observations and variography. GRD Minproc considers that there is sufficient drilling and sampling information, and that this information is of sufficient quality to classify the mineral resource in Measured, Indicated and Inferred categories over different parts of the deposit. The grade model was validated visually, statistically by comparison with declustered input (composite) data for each domain, and by comparison with the Inverse Distance Squared model generated from the same data using the same search parameters. 1.2.4

Mineral resources

The Mineral Resources are reported in Table 1.1, Table 1.2 and Table 1.3 for the combined Constancia and San José zones. The economic cut-off is an NSR value that varies depending on ore type and depth, and thus cannot be equated directly with a copper cut-off grade. Table 1.1 Constancia Project Global Mineral Resource Estimate (0.25% Cu Cut-off) Category

Cut off

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

MEASURED

0.25

119.2

0.47

0.014

3.73

0.05

INDICATED

0.25

195.3

0.48

0.010

4.17

0.06

MEAS+IND

0.25

314.5

0.47

0.012

4.00

0.05

INFERRED

0.25

28.5

0.45

0.009

4.75

0.07

Table 1.2 Constancia Project Global Mineral Resource Estimate (0.20% Cu Cut-off) Category

Cut off

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

MEASURED

0.20

138.3

0.44

0.013

3.54

0.04

INDICATED

0.20

254.2

0.42

0.010

3.81

0.05

MEAS+IND

0.20

392.5

0.42

0.011

3.72

0.05

INFERRED

0.20

48.8

0.35

0.008

3.82

0.06

Page 24

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 1.3 Constancia Project Global Mineral Resource Estimate (0.15% Cu Cut-off) Category

Cut off

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

MEASURED

0.15

146.8

0.42

0.013

3.46

0.04

INDICATED

0.15

376.2

0.34

0.008

3.24

0.05

MEAS+IND

0.15

523.0

0.36

0.009

3.30

0.04

INFERRED

0.15

144.6

0.23

0.006

2.53

0.04

Further reporting by domain and for different cut-off grades are included in Section 17. It is important to note the following when considering the grade and tonnage estimates: 

Mineral Resources that are not Ore Reserves do not have demonstrated economic viability.



Measured and Indicated Mineral Resources are that part of a Mineral Resource for which quantity and grade can be estimated with a level of confidence sufficient to allow the application of technical and economic parameters to support mine planning and evaluation of the economic viability of the deposit.



An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade can be estimated on the basis of geological evidence and limited sampling, and reasonably assumed but not verified continuity.

1.3

MINING

1.3.1

Introduction

The Constancia deposit is massive and relatively near-surface, making it amenable to open pit mining. It is, however, at high altitude (up to 4500 masl). Groundwater is present, and there is significant rainfall during the wet season (October to March). Rock strengths are moderate to hard. 1.3.2

Pit optimisation and pit design

Pit optimisation was undertaken on a regularised version of the resource model with 25x25x15 m blocks. Optimisation input parameters were based on preliminary information, including then current geotechnical information, specifically overall slope input (45° to 50°), and a copper price of $1.80/lb. Ultimate and staged pit designs were developed by GRD Minproc from optimisation shells. Four pit stages were selected for Constancia deposit and a single stage for San José. Pit access ramps are 30 m wide with 1:10 gradient and are designed to accommodate 220 t sized trucks including allowances for the construction of safety windrows and drainage. A 15 m single lane ramp has been designed for the last 60 vertical metres of pits and sub-pits. A minimum mining width of 50 m is maintained. The ultimate pit design is shown in Figure 1.2.

Page 25

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 1.2 Ultimate Pit – Plan View

1.3.3

Mineral reserve

The Constancia Mineral Reserve is the Measured and Indicated Resource contained in the open pit mine that can be processed at a profit and is scheduled for treatment in the DFS Life-of-Mine (LOM) plan. The Mineral Reserve estimate including Proven and Probable categories is summarised in Table 1.4. Since revenue is derived from four payable components (copper, molybdenum, silver plus minor payable gold) the reserve reporting cut-off is based on a Net Smelter Return (NSR) cut-off that is estimated using the metal prices and other treatment, recovery and concentrate realisation parameters as detailed elsewhere in the report. Table 1.4 Constancia Project Mineral Reserves Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

PROVEN

Category

161.8

0.45

0.012

3.68

0.05

PROBABLE

115.6

0.40

0.011

3.70

0.05

TOTAL

277.4

0.43

0.012

3.69

0.05

Page 26


1.3.4

Mine and process schedules

Mine and process scheduling was carried out on a monthly basis for the pre-strip (Year-1) and first year of production, quarterly for years 2 through 5 and annually thereafter. The schedule is based on an annual mining rate of 45 Mt. Figure 1.3 shows annual mining production by ore type. Figure 1.3 Annual Ore Production by Ore Type 35

High Zn Mt 30

Skarn 2 Mt Skarn 1 Mt

25

Supergene Mt Hypogene Mt

20

15

10

5

Yr 25

Yr 24

Yr 23

Yr 22

Yr 21

Yr 20

Yr 19

Yr 18

Yr 17

Yr 16

Yr 15

Yr 14

Yr 13

Yr 12

Yr 11

Yr 10

Yr 9

Yr 8

Yr 7

Yr 6

Yr 5

Yr 4

Yr 3

Yr 2

Yr 1

Yr‐1

0

Three principal ore types are present at Constancia, namely Supergene, Hypogene and Skarn. However testwork has shown that high zinc zones, particularly high zinc Skarn, tend to produce elevated zinc levels in resultant copper concentrates. Consequently, the mining blocks were re-coded according to Zn/Cu ratio for reporting. High zinc material will require blending with lower zinc ores. Testwork also indicates that metallurgical performance and copper concentrate quality are optimised if high zinc and Supergene ores are kept separate, whereas mixing of other ores is not an issue. The mine design and schedule allows for a degree of in-pit control of plant feed, while allowance has been made for stockpiling and rehandling to allow batch treatment of the different ore blends. 1.3.5

Mine fleet assessment

The following equipment has been selected to undertake large-scale bulk mining on 15 m benches:

Page 27




32 m3 capacity electric shovels. Electric shovels are preferred given the altitude of the Project.



220 t class haul trucks.



Crawler-mounted electric drills, capable of drilling 15 m benches on a single pass.



Support equipment including dozers, graders, excavator and water truck.

Fleet numbers were determined to achieve the required production of 45 Mt/a, based on haul profiles and load/haul cycles. Equipment numbers are included in Table 1.5. Table 1.5 Equipment Fleet, Capital and Hourly Operating Costs

Type

Equipment Class

Shovel 32 m3 Dump Truck 220 tonne FEL 18 m3 Track Dozers 391 kW Wheel Dozers 362 kW Graders 209 kW Water Truck 91 kl Integrated Tool Carrier 2.5 m3 Rock Breaker H130S Hammer Excavator General Duties 12m3 Cable Reeler Production Drill Diesel 279 mm Pre‐split Drill 127 mm Service Truck Tyre Handler Low Loader Light Vehicle Passenger Bus Lighting Plants 9m Hydraulic Mast

1.3.6

Fleet Units

2 13 1 2 1 2 2 1 1 1 1 3 1 1 1 1 14 2 8

Operating Operating Purchase Hours Costs Price Hrs/yr US$/hr US$M 6 701 6 701 5 585 5 046 4 840 5 606 3 723 3 723 2 234 3 723 1 489 4 906 5 957 3 723 2 102 491 1 862 1 862 3 723

283.99 267.28 345.81 129.11 183.38 114.18 105.66 44.29 54.62 188.65 118.75 276.64 159.68 20.00 32.86 103.54 9.45 11.20 8.17

Expected Life Hrs

16.29 100 000 3.43 65 000 3.74 50 000 0.94 30 000 0.82 30 000 0.67 30 000 1.54 60 000 0.23 40 000 0.42 35 000 2.79 60 000 0.86 30 000 1.54 60 000 0.70 25 000 0.17 25 000 0.15 30 000 0.63 20 000 0.03 10 000 0.27 15 000 0.03 25 000

Mine capital costs

Mine capital costs were developed based on supplier quotes, and include provision for replacement units (Table 1.7). Total costs are $164.93 M, including $119.51 M of initial capital. Table 1.6 outlines the major components of the mine capital costs estimated for the project. Mine capital costs do not include the infrastructure costs related to the mine including a $1.0 M provision to provide initial fuel, explosive and workshop facilities for the civil fleet. Those costs are included under general infrastructure facilities.

Page 28

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 1.6 Mine Capital Costs by Component

CAPITAL COSTS Units ITEMS Equipment Purchases Replacement Purchases & Overhauls Start up Spares Trailing Cables Replacement & Repair - Electric Shovel Cable Towers and Crossing Ramps, cable trays Tyres ( one set only) Crane rental for equiment commisioning Survey Equipment Shovel dipper Truck Tray Mining Software & Systems Heavy Equip Assembly ‐ Crane hire, transport etc Blasting Contractor Mob/Demob Dispatch and fleet management system Prestrip Total

$ $ $ $ $ $ $ $ $ $ $ $ $ $ $ $

Total 99 580 055 37 591 030 2 761 760 1 008 000 1 000 000 252 387 180 000 80 000 500 000 891 292 150 000 600 000 136 000 1 200 000 19 002 416 164 932 941

Page 29


Table 1.7 Mine Capital, Sustaining and Replacement Costs ($ M) Mine Area

Yr-1

Loading

36.73

Hauling

45.63

0.45

3.86

1.54

Drill & Blast Support

7.11

Other

7.19

Prestrip

19. 0

Total:

119.51

Yr 1

Yr 2

Yr 3

Yr 4

Yr 5

Yr 6

Yr 7

Yr 8

Yr 9

0.06

0.31

0.06

0.06

0.26

0.06

0.06

3.98

0.45

Yr 11

Yr 12

Yr 13

Yr 14

0.06

0.06

0.26

0.06

0.06

4.03 0.36

0.61

0.23

Total 42.04 70.44

0.35

0.61

Yr 15

23.91

2.79 0.34

Yr 10

0.07

5.82

3.62

17.56

0.73

10.08 19.0

0.34

5.45

0.31

0.51

0.42

4.29

0.67

0.06

3.98

24.55

4.41

0.26

0.06

0.06

0.07

164.93

Page 30


1.3.7

Mine operating cost

The mine equipment will be owned and operated by Norsemont. Third parties will manage explosive supply and diesel supply/storage/dispensing. Key operating cost drivers include diesel at $0.66/L and ANFO at $533/t. Powder factors are expected to range from 0.30 kg/t in oxide and NPAG waste to 0.40 kg/t in hypogene ore. Mine operating costs over life of mine are shown in Figure 1.4. Average cost is $1.19/t mined, including the pre-strip period. Figure 1.4 Mine Operating Costs over Time ($/t)

$4.00 $3.50 $3.00

Other

Labour

Support Equip

Blast

Drill

Hauling

Loading

$2.50 $2.00 $1.50 $1.00 $0.50

1.4 1.4.1

Yr15

Yr14

Yr13

Yr12

Yr11

Yr10

Yr9

Yr8

Yr7

Yr6

Yr5

Yr4

Yr3

Yr2

Yr1

Yr‐1

$0.00

GEOTECHNICAL INVESTIGATIONS Site investigations

Geotechnical investigations were carried out at the site in order to (1) map and characterise the geologic units and structures, (2) gain an understanding of the subsurface soil and rock types, their depths and engineering properties, (3) characterise underlying geotechnical and subsurface conditions to support THE hydrogeological program, (4) identify and characterise potential borrow material sources to evaluate their suitability for use as construction materials, and (5) obtain soil and rock samples for laboratory testing. The field work was conducted in phases between October 2007 and June 2009 and focused on areas where the main project structures will be located. These included the open pit, tailings management facility (TMF), Potential Acid Generating (PAG) Waste Rock Facility (WRF), and the process plant (Figure 1.5).

Page 31

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 1.5 Overall Site Plan

The geotechnical investigations included geologic site mapping and reconnaissance, excavation and sampling of test pits, Dynamic Probing Light (DPL) testing to determine the depths of soft, wet organic deposits in certain areas, geotechnical drilling and sampling at selected sites, Standard Penetration Testing (SPT) in the geotechnical drillholes, in-situ permeability testing in the geotechnical drillholes, piezometer installations and monitoring in the geotechnical drillholes and sample recovery and testing for index, compressibility, strength, and permeability properties. 1.4.2

Pit geotechnical design

In the Constancia pit area approximately 85% of the rock consists of intrusive rocks; 5% consists of sandstones and 10% of skarn. Structurally the Constancia pit area is controlled by four major systems expressed as regional faults and local faults, some of which follow regional trends. These structures

Page 32


influence the quality of the rock mass and therefore the slope stability of the pit walls, which influences the allowable slope angles of the walls. To assist in characterizing the rock mass and structures, a geotechnical database was developed using geotechnical site data obtained from exploration and geotechnical programs over the period of 2005 to 2009. The RMR classification data by core run in this database showed that 58% of the values corresponded to “poor” rock, 38% to “fair” and 4% to “good” rock. Pit slope stability analyses for the ultimate pit configuration were based on a geotechnical pit slope investigation program. Nine pit design sectors were established to group areas of the proposed pit having similar geometric, geological and rock mass quality characteristics. Methods used to develop the stability assessments for each sector included detailed kinematic stability, limit equilibrium analysis at bench face and inter-ramp scales, probabilistic analysis by the equilibrium limit method; and, in certain design sectors, stress analysis using finite element methods. The pit slope geometries for each design sector have been determined based on acceptance criteria according to each of these design methods and the expected maintenance operational cost. The maximum recommended angle is the flattest angle meeting either the catch bench or inter-ramp acceptance criteria. The minimum acceptable factors of safety (FoS) adopted for the Constancia pit walls were selected in accordance with the current engineering practice and Peruvian regulations. A value of 1.2 was adopted for static conditions; and 1.0 was adopted for earthquake loading using the peak ground acceleration for a 100-year return event. Overall, the probability of failure ranges between 8 and 27% in about 77% of the bench faces, which conforms to the chosen acceptance criteria (less than 35%). However, in the southwest and west pit sectors a probability of failure of 55% could be expected in about 23% of the bench face slopes. An evaluation of the operational cost needs to be developed considering this failure probability. The recommended overall slope in each sector was selected based on either catch bench integrity (catch bench design) or by global stability considerations. Stability analysis results indicate that a bench face angle of 65-70o or steeper is expected to be achievable at most places. The mining configuration considered for pit development consists of 30 m high double benches and berm widths of 11.5 m minimum. Recommended inter-ramp slope angles range from 48-54o depending on the design sector. Figure 1.6 presents the proposed design sectors and proposed inter-ramp and bench slopes. Pit optimisation and design performed by GRD Minproc as outlined in Section 1.3.2 was based on preliminary pit slope recommendations. After pit slope optimisation studies were performed and a specific pit wall configuration had been developed, the stability factors were checked according to the geologic conditions projected to occur in those specific pit wall configurations and locations. For design sectors VI and VII, it was found that slopes at the angles assumed in the optimisation studies held greater risk of instabilities. Slope flattening would mitigate the risk, but after assessment of the impacts, and due to timing considerations, the steeper preliminary pit slopes were used for the study. Additional slope management measures, as outlined in Section 18.1.3, were incorporated into the planned mining approach and the mining cost estimate, to offset the increased slope failure risk.

Page 33

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 1.6 Proposed Design Sectors, Inter-ramp and Bench Slopes

Page 34


1.4.3

Subsurface conditions

Subsurface conditions at the locations of the TMF, PAG WRF, and plant sites are closely related to the geologic units present as well as topographic features associated with erosion and glaciation. Bedrock at the site is mainly comprised of sedimentary rock of the Cretaceous period and intrusive rock of the Paleogene period. More recent soil formations are of the Quaternary period, and include organic bogs with some laccustrine deposits, alluvium and glacial till deposits. Under the site of the TMF bedrock consists of diorite and sandstone. The TMF site features two prominent south to north orientated valleys and in these valley bottoms the bedrock is overlain by bog deposits. The bogs are generally only a few meters deep but reach depths of up to 12 m in their centres. Current development plans for the TMF call for completely removing the bog deposits. It is estimated that a suitable foundation for the TMF embankment will be reached at depths of approximately 2 to 5 m below the existing ground surface and below the bottoms of the bogs. In the diorite however, the bedrock appears to be extremely weathered to depths greater than 20 m in some areas. Further investigations will be required to evaluate the need for removing the diorite to significant depths under the embankment. Bedrock under the site of the PAG WRF also consists of diorite and sandstone; however, a large portion of the site is overlain by consolidated glacial till deposits which vary in thickness from between 25 and 95 m. In the lowest lying areas of the site, bog and alluvial soils were encountered. Unsuitable foundation soils are currently planned for removal from beneath the critical slopes of the PAG WRF in order to maintain adequate stability of the slopes At the site of the process plant, foundation materials consist of sandstones, diorites and monzonites. Competent rock was found at depths between 7 and 10 m under the locations of the main structural components of the plant. Overlying the more competent bedrock, residual soils and very poor quality rock were found, which will most likely need to be removed during construction. 1.4.4

Construction materials

During the geotechnical investigations, potential sources for random/common fill, structural fill, drainage material, low permeable/core material, filter material wearing course and rip-rap were identified for construction materials. Current plans call for the non-PAG waste rock mined from the pit to be used for constructing a large portion of the TMF embankment, as well as for haul roads and access roads at the site. 1.4.5

Seismic conditions

The site is located in a zone that has experienced historical seismicity mostly related to the subducting Nazca plate passing under the South American plate. The province is located within Zone 2 (intermediate) of the 2003 Peruvian Seismic Resistance Code. Two potentially significant earthquakes have been identified as potential Maximum Design Earthquakes (MDEs) as follows: Magnitude 8.0 deep intraplate subduction earthquake, producing a mean plus one standard deviation PGA at the site of 0.38 g and Magnitude 7.0 shallow crustal earthquake artificially

Page 35


located 25 km directly below the site, producing a mean plus one standard deviation PGA at the site of 0.25 g. The natural slopes in the areas of proposed project structures appear stable and display no observed risk of mass displacement, sliding or landslides from historic events. 1.5

METALLURGICAL TESTWORK

An extensive laboratory testwork campaign was undertaken to evaluate metallurgical performance of the Constancia deposit at a DFS level of detail. In addition, a 25 t pilot campaign was conducted to provide concentrate for molybdenum flotation evaluation, settling, regrind and filtration tests. The sample collection and testwork were directed by GRD Minproc with assistance from Transmin Metallurgical Consultants (Lima). 1.5.1

Testwork sample selection

Representative samples were selected for metallurgical testwork, taking into account variations in lithology, mineralisation type, grade and three-dimensional location. To facilitate location and depth discrimination, the ore body was divided into “metblocks”, each metblock being a 100 m cube. Continuous core runs within the metblocks that met the sample selection criteria were then identified. From the available suitable core, a sample set that provided broad spatial representation and was relevant to the likely mine plan for the ore body was selected. 1.5.2

Comminution testwork

Comminution testwork was conducted on samples representing Supergene, Skarn and Hypogene ore types. The testwork campaign was conducted at SGS Lakefield Research in Chile during 2008, and included Bond abrasion, rod mill and ball mill work index, SMC and JK Drop Weight Index tests. Supergene material was classified as the least competent ore type at Constancia with an 80th percentile DWi of 3.7 kWh/m3. In terms of ore hardness for ball milling, Supergene ore is of moderate hardness with an 80th percentile BWi of 13.1 kWh/t. Hypogene ore is classified as the most competent ore type at Constancia with an 80th percentile DWi of 7.5 kWh/m3. In terms of ore hardness for ball milling, Hypogene ore is also the hardest with an 80th percentile BWi of 16.3 kWh/t. Skarn samples were analysed as two groups, medium-zinc Skarn (soft) and high-zinc Skarn (hard). The soft Skarn is located in the upper part of the orebody and the hard Skarn is deeper. There were eight samples representing medium-zinc Skarn (soft) and only three representing high-zinc Skarn (hard). Soft Skarn is classified as the least competent ore type at Constancia with an 80th percentile DWi of 3.8 kWh/m3. In terms of ore hardness for ball milling, Skarn is the softest with an 80th percentile BWi of 10.6 kWh/t. Hard Skarn is classified as competent as hypogene with an 80th percentile DWi of around 7.5 kWh/m3. In terms of ore hardness for ball milling, hard Skarn is slightly harder than soft Skarn with an 80th percentile BWi of around 11.5 kWh/t.

Page 36


1.5.3

Flotation testwork

The first flotation testwork campaign focused on identifying optimum processing parameters to produce a bulk copper / molybdenum concentrate. It evaluated the impact of grind size, reagents and pH on recovery, flotation kinetics and concentrate grades for the three main ore types. Variability and locked cycle tests were conducted and deportment of zinc and other impurities investigated. Supergene ore had a low metallurgical complexity with moderate variability. The locked cycle test produced concentrate grades of 28.6% to 30.2% Cu, at 86% recovery. Deleterious elements, including zinc, were all below penalty levels, however, zinc was present in sufficient quantity to have a material affect on transport and smelting charges. Hypogene ore had a low metallurgical complexity with moderately high variability. Copper recovery decreased at depth and recoveries are based on upper Hypogene ore and lower Hypogene ore, respectively. Upper Hypogene ore exhibited low variability, while lower Hypogene ore exhibited moderately high variability. The locked cycle test produced concentrate grades of 23.1% to 25.7% Cu, at 88% recovery. With the exception of zinc and lead, other elements were well below possible penalty limits. Zinc (1.62 % of concentrate) and lead (0.2 % of concentrate) were below expected penalty levels, but comprise a sufficient portion of the concentrate to have a material effect on transport and smelting charges. Skarn ore had a moderate metallurgical complexity with high variability. The locked cycle test produced concentrate grades of 18.9% to 24.7%Cu, at 85% recovery, depending on the zinc to copper ratio in the feed. Cadmium and bismuth were at values close to their penalty limits at 0.0325% Cd and 0.029% Bi, and will require monitoring during operation. Zinc at 9.9%-19.9% of concentrate weight is well above penalty limits. Depression tests showed that zinc sulphate could be used to lower the zinc content of the concentrate, with some loss in copper recovery. Expected elemental recoveries for the different ore lithologies are shown in Table 1.8

Page 37

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 1.8 Expected Elemental Recoveries Copper Concentrate

Cu

%

87.00%

After yr10

Hypogene

Cu

%

91.40%

Up to yr10

Zn

%

30.00%

Ag

%

80.00%

Au

%

60.00%

Pb

%

25.00%

Cu

%

89.00%

Zn

%

80.00%

Ag

%

80.00%

Au

%

60.00%

Pb

%

25.00%

Cu

%

89.00%

Zn

%

30.00%

Ag

%

80.00%

Au

%

60.00%

Pb

%

25.00%

Supergene

Skarn

Samples were also generated to evaluate the effect of blending and treating different ore lithologies within the concentrator. The blending of Supergene and Skarn ore produced a blend that had a high metallurgical complexity with high variability. Activation of sphalerite by soluble copper or sulpho-salts was suspected, leading to high levels of zinc content in the copper concentrate. Depression tests showed that sodium cyanide could be used to lower the zinc recovery, however, the loss in copper recovery was significant. Other ore blends showed no detrimental effects on grade or recovery. All ore types require a regrind of rougher concentrate, with P80 varying from 25 µm to 40 µm. Rougher concentrate mass recoveries varied from 10% to 20%, depending on ore type. Recovery of molybdenum was simulated using floatability component modelling. This indicated a recovery to molybdenum product of 55% at a grade of 40% Mo. The presence of talc resulted in lower grade molybdenum concentrates. 1.6

PROCESS DESCRIPTION AND PLANT DESIGN

A simplified process flowsheet is shown in Figure 1.7 and the plant layout in Figure 1.8.

Page 38

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 1.7 Constancia Process Flowsheet

Pebble Crushers Primary Cyclones

Primary Crusher SAG Mill

Ball Mills Coarse Ore Stockpile

Copper Roughers

Tailing Thickener

Regrind Cyclones

Regrind IsaMill Mo Flotation Circuit Copper Thickening and Filtration Molybdenum Thickening and Filtration

Page 39

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 1.8 Constancia Plant Layout

1.6.1

Throughput

Throughput rates vary depending on the ore type being treated, as shown in Table 1.9. In the first two years of operation supergene ore is dominant and the highest throughput rates are achieved. In Years 3 and 4, supergene, hypogene and skarn ore are mined in similar quantities. Hypogene ore becomes dominant through to the final years of operation and throughput rate drops progressively. Table 1.9 Theoretical Maximum Throughput For Each Ore Type Ore Type

t/h

t/d

Flotation Feed P80 (µm)

Hypogene

1990

47 760

106

Supergene

2973

71 352

108

Skarn (soft)

3158

75 792

70

Skarn (hard)

2520

60 480

50

Page 40


The mine plan and treatment strategy is based on campaigning separate ore types as much as possible. In most years all three ore types will be mined and this is reflected in lower actual throughputs. Table 1.10 shows the schedule of design throughput rates for the plant. Table 1.10 Plant Design Throughput Rate

1.6.2

Year

t/h

t/d

Mt/a

1

2700

64 800

21.6

2

2738

65 700

21.9

3

2688

64 500

21.5

4

2638

63 300

21.1

5

2450

58 800

19.6

6

2488

59 700

19.9

7

2300

55 200

18.4

8

2150

51 600

17.2

9

2138

51 300

17.1

10

2075

49 800

16.6

11

2013

48 300

16.1

12

2075

49 800

16.6

13

2125

51 000

17.0

14

2113

50 700

16.9

15

1988

47 700

15.9

Crushing

The primary gyratory crusher is fed by rear-dumping from two dump points by haul trucks, or fed by FEL from a stockpile. The crusher dump pocket is fitted with a drive-in ramp allowing FEL and bobcat access into the dump pocket for cleanout prior to maintenance work being undertaken, or for clearing blockages during normal operation. Dust suppression water is sprayed within the dump hopper in conjunction with tipping of each truck load. Crushed ore is conveyed to a 50 278 t open stockpile ahead of the concentrator plant. Three variable speed belt feeders reclaim ore from beneath the stockpile and discharge onto the SAG mill feed conveyor. Air, water and fire suppression services are run in the tunnel. Access and emergency exit tunnels are constructed beneath the stockpile. The main tunnel has feeders, the SAG mill conveyor and carries piping for air, water, fire services and electrical cables. 1.6.3

Grinding

The grinding circuit consists of a single line SABC circuit using a variable speed SAG Mill in closed circuit with pebble crushing and two fixed speed ball mills. The target grind size is a P80 of 106 µm. Steel balls of nominal 125 mm size are added to the SAG feed conveyor. The ball charge will be kept to a level of 12-15%, depending on the ore properties. The slurry in the mill exits through the discharge grate (with pebble ports -65 mm) and passes over the SAG mill discharge screen. Pebbles from the

Page 41


screen oversize are conveyed to the pebble crushing plant. SAG Mill discharge slurry which passes through the screen enters the SAG mill discharge hopper together with the discharge from both ball mills which gravitate through ball mill trommels. Dilution process water is added before slurry is pumped by the cyclone feed pump to the cyclone cluster for classification. 1.6.4

Copper flotation

Copper flotation feed is conditioned with slaked lime to ensure the circuit pH is maintained at its set value. A3302 collector, AF65 frother and ZnSO4 depressant are added to the conditioning tank, as and when required. On-stream analysis of copper and zinc in the flotation feed will be used to determine reagent requirements. The conditioned feed reports to two rougher flotation banks. Flotation concentrate is reground prior to three cleaning stages. The copper cleaner circuit consists of three stages of cleaning and one bank of cleaner scavenger flotation cells. On-stream analysis monitors the zinc and copper grades of the major concentrate and tailing streams to allow performance to be optimised. 1.6.5

Molybdenum flotation

The bulk copper/molybdenum concentrate reports to a thickener for removal of reagents that are present from copper flotation. The thickener underflow is pumped to a molybdenum rougher conditioning tank where NaHS is added to inhibit the flotation of copper minerals and sphalerite, with a light fuel oil promoter added to enhance the flotation of molybdenum. The molybdenum flotation circuit utilises covered, induced air flotation cells, with internal air recirculation. It consists of one roughing stage and seven cleaning stages. 1.6.6

Copper thickening and filtration

The tailing from the molybdenum flotation roughers and molybdenum cleaner scavengers reports to a copper concentrate thickener, via a static screen to remove any tramp material. Flocculant is added to enhance settling. The clarified thickener overflow reports to a molybdenum circuit process water tank and is used as filter cloth wash, flush water, copper thickener spray water and for general use in the molybdenum flotation area. The thickener underflow is removed at 60% solids via a peristaltic pump. The thickened, de-tramped slurry is stored in two agitated tanks. These provide a 24 hour surge capacity, allowing filter maintenance to be conducted without affecting mill throughput. The filter feed is pumped to three pressure filters and filter cake is dropped onto a conveyor and is conveyed to the copper concentrate stockpile. 1.6.7

Molybdenum thickening and filtration

The molybdenum concentrate gravitates to a thickener where it is thickened to 60% solids. The thickener overflow reports to the molybdenum process water tank from where is used as process water in the molybdenum flotation circuit. The thickened concentrate is pumped to a ferric chloride leach tank where copper, zinc and lead present in the concentrate is dissolved to reduce their content to less than 0.5%. The slurry is then pumped to a pressure filter to produce a filter cake of 25% moisture. The filtrate from the filter reports to the tailing thickener. Filter cake is transferred by bobcat to the bagging plant where it is bagged in 1 m3 bulk bags.

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1.6.8

Tailing thickening

The copper rougher tailing stream flows by gravity from the two rougher lines to a thickener, where it is combined with copper cleaner scavenger tailing and thickened to 52% solids. The thickened underflow is then pumped to the TMF. The thickener overflow gravitates to the main process water tank from where it is used in the grinding and copper roughing circuits. 1.6.9

Tailing water reclaim

Water is reclaimed from the TMF and pumped to a process water pond. From the process water pond it is pumped to the process water tank as required. 1.6.10

Concentrate storage and loadout

The copper concentrate is conveyed to an undercover storage shed. There is a live stockpile with seven days production capacity, and, in addition, a low grade concentrate stockpile with seven days production capacity. The concentrate is transferred from the stockpiles via FEL onto trucks for transport from the mine site to the terminal warehouse facility at the port facilities in Matarani. 1.6.11

Water services

Raw water is pumped from bores to the fire/raw water tank. The raw water tank provides water to the potable water treatment plant. The treated water is pumped to a potable water tank which discharges into the potable water reticulation system and the safety shower water network. Raw water is also piped to the molybdenum flotation area and also provides cooling water to lubrication areas and the gland water system. 1.6.12

Reagents

Slaked lime is pumped from a storage tank to the SAG mill, the copper rougher conditioning tank, the copper cleaner conditioning tank and the acidic water neutralisation area. Zinc sulphate is mixed and pumped to the copper rougher and cleaner conditioning tanks, as required. Solid NaHS is mixed and pumped to the molybdenum flotation circuit. Light fuel oil is used to assist flotation of molybdenum minerals in the molybdenum flotation circuit. Hydrochloric acid is used in conjunction with ferric chloride in the leach section of the molybdenum thickening/filtration area. A3302 collector is used for floating copper and molybdenum in the copper flotation circuit with addition to the primary cyclone feed sump, the copper rougher conditioning tank and the copper cleaner conditioning tank. Z-14 SIBX is a secondary collector in the copper rougher circuit. AF65 frother is used to provide a stable froth in the flotation cells. Flocculant is mixed in a dedicated plant and pumped to the copper/molybdenum feed thickener, the copper rougher tailings thickener, the copper concentrate thickener and the molybdenum concentrate thickener. 1.6.13

On stream analysis and laboratory

Two on-stream analysis systems have been included to allow continuous analysis of performance. The copper flotation circuit system will analyse Cu, Fe, Zn and Pb in the rougher feed, rougher concentrate,

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rougher tailing, cleaner scavenger concentrate and bulk copper/molybdenum concentrate. The molybdenum plant system will analyse Mo, Cu, Fe, Zn and Pb in the flotation feed, rougher concentrate, rougher tailing, cleaner scavenger concentrate and molybdenum concentrate before and after the ferric leach. The laboratory facility is located within the plant site. The laboratory is capable of processing samples for mine grade control, exploration, process plant metallurgical accounting, metallurgical optimisation and environmental control. In addition, the laboratory has the facility to undertake testwork to optimise grinding, flotation, leaching, etc. 1.7

WASTE MANAGEMENT

Major waste management facilities within the project area include the PAG WRF and the TMF. Unsuitable material and topsoil generated during construction site preparation activities will be disposed in dedicated structures or in the major waste management facility. Figure 1.5 presents the Constancia mine overall site plan and the location of these facilities. The major waste management facilities are described in the following sub-sections. 1.7.1

Waste rock facility (WRF)

A PAG WRF will be developed to store waste rock mined from the San José and Constancia Pits. Approximately 295 Mt of waste have been characterised as having the potential to generate acid and will be placed in the PAG WRF. Approximately 55 Mt of waste have been characterised as non-acid generating and will be used as construction material for the TMF embankment, haul roads, construction roads and access roads. The PAG WRF will be located immediately south of the Constancia pit. The PAG WRF has been designed with an ultimate storage capacity of 300 Mt, which will result in the maximum elevation of the facility at 4335 masl and a maximum vertical height of 200 m. Development of the PAG WRF is planned in five stages with the objective of reducing haul distances during the initial years of mining. Loading will start at the northern limits of the facility closest to the pit and will progress southwest towards the ultimate toe of the facility. The hydrogeologic studies and modelling show that the natural groundwater levels and gradients beneath the PAG WRF will provide hydraulic containment for any seepage, which will be directed to a single reporting point below the southwest toe of the facility. This containment precludes the need for a liner. The design does, however, include a robust underdrain system to collect the seepage from the base of the waste rock and direct it to the reporting point. Seepage from the PAG WRF and groundwater under the facility will be directed to a containment pond constructed downstream of the PAG WRF and retention pond. This pond will also contain surface runoff from the PAG WRF. A 28 m high earthfill embankment will provide approximately 600 000 m3 of water storage. The design consists of a cross-valley zoned embankment with a grouted curtain which spans nearly the entire length of the embankment and extends to depths of approximately 40 m into rock. Water stored in the pond will be used as process water for the mill after treatment. During infrequent, extreme wet periods, excess water may accumulate in the containment pond and be pumped to the upper surface of the PAG WRF for recirculation through the dump as a means of adding additional temporary storage. Water from the PAG WRF is not intended to be released; however, a

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valve controlled outlet pipe has been incorporated into the design to be used only in the event of an emergency in which water levels within the pond need to be rapidly lowered. A retention pond between the toe of the PAG WRF and containment pond will be constructed to contain any rocks falling down over the slope of the WRF and to provide for energy dissipation of the drainage flows. A 5 m high, flow-through, rockfill embankment will be constructed at the downstream end of the retention pond, which will provide an approximate capacity of 24 000 m3. 1.7.2

Tailings management facility (TMF)

The TMF will be developed behind an embankment dam crossing two broad, gently sloped, south to north valleys above the south side of the Chilloroya River as a site 5.2 km southwest of the mine and 3.7 km south of the process plant. The TMF has been designed with an overall storage capacity of 277 Mt (dry) of tailings assuming an average in storage dry density of 1.5 t/m3. The embankment will have an ultimate height of 130 m and length of 2300 m, although additional storage can be developed by raising the TMF above this elevation. The embankment will consist of a zoned earthfill structure that will be constructed in stages out of local borrow materials and selected non-PAG mine waste. During the first two years of operations the embankment will be constructed following a downstream configuration, but from Year 3 onward a modified centreline approach will be adopted. Tailings will be deposited from designated off-take points on a distribution pipeline along the upstream crest of the embankment. The points of active deposition will be rotated frequently to form a thin layered, drained and well consolidated beach that will slope away from the embankment towards the south side of the TMF basin. Initially, prior to beach development, the surface water pond will be in contact with the embankment in the east valley but it will be progressively displaced upward and to the south as the beach becomes established such that after the first two years of operations the beach is expected to have displaced the pond well away from the embankment. The surface water pond will vary in size throughout the life of mine depending on the season, precipitation, and operational requirements. Tailings deposited in the facility will consist of rougher tailings (RT) and cleaner scavenger tailings (CST) from the process. These streams will be combined at the plant at an approximate ratio of 4 to 1 RT to CST, prior to transportation and deposition into the TMF. Although the CST will contain significant sulphide minerals that make it potentially acid generating, the combined stream will have excess alkalinity from the mill such that its pH will initially be in the order of 8 to 9. Geochemical analyses indicate that an exposure period of six months to a year would be necessary for this alkalinity to be consumed, and the tailings deposition plan calls for a fresh layer of tailings to be placed over each previously deposited layer well within this period to reduce the potential for acidic conditions developing. The TMF includes a geomembrane liner over the base of the eastern valley and most of the western valley to provide geomembrane containment in areas where the surface water pond will be in contact with the base at any time over the life of mine. A 50 m wide tailings underdrain will be placed on top of

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the geomembrane against the upstream toe of the embankment to assist in depressing pore pressures in the tailings against the embankment and to minimise the head on this part of the geomembrane liner. The geomembrane liner has not been extended into the upper, southern reaches of the western valley since the surface water pond will never be located there. However, to assist in intercepting and collecting any small amounts of seepage that may pass beneath the western side of the embankment in the absence of the liner, an intercept trench and drain will be constructed under the embankment across this valley. Groundwater drains will be installed under the geomembrane to intercept groundwater seeps and keep them separate from the geomembrane. A separate foundation drain system will be installed under the embankment to collect water seeping through the embankment, which may include direct precipitation on the embankment and/or small amounts of seepage passing through it, as well as localised groundwater seeps in the embankment foundation. Water collected by the drain systems will be conveyed to sumps located immediately downstream of the embankment in each of the east and west valleys. Monitoring and control systems at the sumps will allow for automated water quality and flow rate determinations to be made. From the sumps, the water will either be released to the Chilloroya River or pumped back to the TMF based on the water quality. 1.7.3

Topsoil and unsuitable material stockpiles

Two main stockpiles, for storage of unsuitable materials and topsoil, are planned for the project. These structures include a combined stockpile (Unsuitable Material/Topsoil Stockpile No. 1), located downstream of the PAG WRF containment pond, which will provide storage capacity for bog, topsoil, and unsuitable materials. A second stockpile (Topsoil Stockpile No. 2) strictly for topsoil will be located directly north of the PAG WRF. The combined stockpile will be developed behind a cross-valley constructed embankment and will provide approximately 2.65 Mm3 of total storage capacity. The topsoil stockpile north of the PAG WRF will predominantly be used for storage of topsoil removed from the pits and WRF. Its capacity will be 0.4 Mm3. 1.8

INFRASTRUCTURE

1.8.1

Access roads

SIGT S.A. Ingenieros Consultores (SIGT), a Peruvian engineering consultant, undertook a study to investigate, design and provide capital costs for improvements to the existing access road required to support construction and operational phases of the Project. The existing access road runs 82.5 km from Yauri to the proposed Constancia Mine, and is unsealed in parts. Width varies from 4-6 m. The topography is typical of high plateaus along the Andes mountain range with small uphill and downhill sections ranging in elevation between 3900 and 4684 masl. Four bridges are present along the route between Yauri and Constancia.

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SIGT’s design provides for an unsealed all-weather road of 6 m width with 0.5 m shoulders and a maximum gradient of 9%. Bridges will be rated to live loads of 48 t, and have asphalt surfacing. The design makes maximum use of the existing alignment and is more than adequate to handle existing traffic plus the additional 50-60 daily truck movements associated with Constancia operations. Geotechnical, geological, hydrological and archeological investigations have been completed, and suitable construction materials identified. Bridge and drainage requirements have been assessed, taking account of extreme rainfall events (500 year return period for major bridges). Pavement works consist of the complete replacement/overlay of the existing pavement with a suitable quarry produced pavement material compliant with the documented design requirements. Drainage is improved through cross-drainage culvert replacement and longitudinal drainage. Signage and other safety improvements will be undertaken. The operating costs for the access road other than the section of road on the Constancia property will be assumed by the government authority. No major environmental constraints have been identified, but attention is required to avoid impacting archeological remnants or other historical buildings, and to minimise the impact of noise, dust and increased likelihood of accidents. The capital cost for the access road is estimated to be $19.00 M, excluding engineering design, construction management and IGV. 1.8.2

Water supply

Total raw water requirements have been estimated at 365 m3/h for process plant, mine and camp operations. Field investigations and groundwater modelling, together with surface water balance calculations, indicate that this will be met by ex-pit and in-pit dewatering operations. This is described in Section 1.9.1. The raw water is of sufficiently high quality to be used directly in the process plant for hose-down, gland water, cooling water and reagent make-up. Up to 11 m3/h of raw water will be diverted to the potable water treatment plant where it will be upgraded by reverse osmosis and ultraviolet treatment, making it fit for human consumption. Process water will be supplied from contaminated water captured by in-pit sumps, the PAG WRF and rain falling on the plant site. Additional requirements will be met by returning tailings dam supernatant water to the plant process water pond.

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1.8.3 1.8.3.1

Power supply Powerline

The design maximum power demand for the Project is 90 MW and the average continuous demand is estimated to be 75 MW. Cesel Ingenieros S.A. (Cesel) completed a power supply study, taking into account current and forecast power demand in the region. A study identified the preferred option is to initially secure supply at 138 kV from the existing Tintaya Substation, with transmission by means of a single circuit supported by lattice steel towers over a route length of 70 km to the Constancia mine site, designed for future operation at 220 kV. It is assumed that 220 kV supply will be available from Tintaya substation in 2012. The transmission line from Tintaya Substation to the Constancia mine site has been designed in accordance with relevant North American and European codes and standards, and in recognition of all environmental, geological, social and cultural considerations related to the land and airspace easement along the proposed route of the line. Topographical and geological surveys and environmental and socio-economic baseline studies were completed along the proposed route and at the Tintaya and Constancia substation sites. Earth resistivity measurements were conducted to obtain the necessary data for the design of the line grounding systems. The transmission line is designed for ultimate operation at 220 kV, will have a design capacity of 150 MW and traverse a route length of approximately 70 km. The transmission line will also include an OPGW or optical fibre composite overhead ground wire serving the multiple functions of earth bonding for the towers, lightning protection shielding for the power conductors and communications via the enclosed 24 optical fibres. The fibre optic communications will provide protection, relaying and control functions for the transmission line and data links from Constancia to the Peruvian communications network. 1.8.3.2

Constancia substation 138/220 kV

The Constancia substation switchyard has been designed with overhead busbar systems, switchgear, metering and protection equipment to control the ultimate installation of two 20 MVA primary transformers to provide the plant with 100% redundancy. Allowance has been made in the design for a bay to accommodate extension of the line or an alternative connection to the Peruvian grid. Initially, one 100 MVA primary transformer has been allowed in the estimate with all associated switchgear, metering and protection equipment to supply the estimated plant maximum demand of 90 MW or 95 MVA at the minimum required COES power factor of 0.95.

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Insulation levels and surge protection equipment reflect the elevation of Constancia Substation in excess of 4000 masl and the level of iso-keraunic activity in the region. 1.8.3.3

Tintaya substation 138 kV expansion

For the initial supply to Constancia at 138 kV, an additional switchyard bay at Tintaya substation has been allowed to accommodate the extension of the existing 138 kV double busbar, 45 MVAR of capacitive compensation equipment and all associated switchgear, transformer, metering and protection equipment for the transmission line. The additional bay proposed for Tintaya substation to serve the transmission line to Constancia will be designed for a capacity in excess of 200 MVA. 1.8.3.4

Power supply transfer from 138 kV to 220 kV

The transmission line and Constancia Substation have been designed for ultimate operation at 220 kV in year 2012. The main power transformer and associated metering and protection transformers have been specified with winding tappings to enable operation at either 138 kV or 220 kV. Following initial operation at 138 kV, the transfer to 220 kV in year 2012 will require a shut-down in order to changeover the power, metering and protection transformer tapping connections. The cost of the proposed 220 kV augmentation of Tintaya substation in 2012 has not been included in the estimate for the DFS. 1.8.3.5

Control and communications

Tintaya Substation is operated by Red de Energía del Perú (REP) from its regional control centre in Arequipa. The new transmission line from Tintaya to Constancia will be integrated into the REP SCADA system for purposes of monitoring, control, energy metering and load management. A local area network will form the platform for monitoring and control of the transmission line and Constancia substation and an operator workstation in the Constancia substation will provide access to monitoring and control functions, system protection, energy measurement and alarms. 1.8.3.6

Power supply capital cost estimate

The capital cost for power supply is estimated at $24.45 M, including the design and construction of the power line and the two substations. 1.8.3.7

Operation and maintenance

The power transmission system to the point of supply at the 22.9 kV terminals of the main transformer at Constancia Substation, will be constructed, owned, operated and maintained by REP. Based on information from OSINERGMIN, an annual cost of operation and maintenance (O&M) of 2% of the annual investment has been considered, producing an estimated figure of $588 527.

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1.8.3.8

Energy consumption and cost

On the basis of an overall average continuous demand of 75 MW at 138 kV and a peak/off-peak ratio of 8/16, the total annual energy consumption is estimated to be 615 GWH. Based on tariff information provided by OSINERGMIN, and an exchange rate of $1=3.2 Soles, this equates to an average cost of ¢4.89/KWH. 1.8.4

Internal roads

Internal roads will be constructed within the Constancia project site. Access between the plant, pit, PAG WRF, and TMF will be on the mine haul roads. Additional access roads will be constructed for connection between the former structures and secondary facilities. 1.8.5

Buildings

A complete inventory of buildings and related facilities has been developed, including: 

Heavy and light vehicle workshop and washbay: the mine office is included in this facility



Heavy vehicle fuel distribution bowser



Mine dispatch centre



Core shed



Process plant industrial buildings to include the milling, pebble crusher, molybdenum flotation, copper and molybdenum concentrate filtration buildings, and the product stockpile. Note, however, that the copper flotation circuit is not enclosed in a building.



Main security and gatehouse building



First aid and fire building



Substation building



Emergency diesel power station



Control rooms for primary crusher and the grinding building



Laboratory and laboratory store



Workshop to support plant operations



Warehouse – including an external fenced compound



Administration building and change-rooms



Kitchen and mess to provide for up to 432 personnel on day shift



Copper flotation blower building



Air compressor house



Reagents and packaging store

The majority of the buildings are of steel frame construction, fully clad, with full length overhead cranes provided for maintenance purposes in the plant buildings. Substations are custom design blockwork

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buildings, while several other buildings are modular pre-fabricated structure, e.g. laboratory, control rooms, fire station and first aid post, gatehouse etc. 1.8.6

Construction and accommodation camp

Construction and operational workforces are housed in a purpose-built accommodation camp located adjacent to the process plant. The camp is supplied and installed by a single construction contractor with another contractor operating the camp during the construction period. The management of the camp will be transferred to the Owner once the construction phase is complete. The camp is to be self-sufficient, with full sleeping, bathing, dining, laundering and recreational facilities. The camp will be constructed from modular units to minimise cost and expedite delivery. The camp comprises single, double and 4-person accommodation units, camp administration offices, internal access roads, potable water treatment, storage and distribution, medical requirements, a recreation hall, waste management, sewage treatment, laundry, landscaping, IT and communications infrastructure, and site security. The capital cost estimate is $30.27 M for the 1800 bed camp, based on budgetary quotes from specialist Peruvian contractors, and has an accuracy of ±15%. This equates to $16 819 per bed. The operational cost estimate is approximately $15 per person per day, again based on budgetary quotes from Peruvian contractors. 1.8.7

Concentrate transport and shipping

Concentrate will be transported by truck from site to the nearest port, Matarani, located 475 km by road from Constancia. Transportation will be undertaken by a specialist haulage contractor. The long term haulage contract would be competitively bid. Hopper-type trucks with a closed cover system will be used, each with a capacity of 35 t. Travel time per truck is estimated at two days, equating to a running fleet requirement of approximately 70 trucks in peak production years. The concentrate will be stored in Matarani in a warehouse owned and operated by the International South Terminal S.A. (TISUR), a private organisation that has held the port operations contract since 1999. The storage facility will be sized to equal the monthly production for the Constancia project (average 21 000 tonnes of concentrate). TISUR will also be responsible for shiploading services. Correspondence between TISUR and Norsemont indicates that port handling charges of $7.50/t (wet) will apply. 1.9

WATER MANAGEMENT

Water management within the project area is divided into process and non-process water. Process water is water that will be used in the plant or will be conveyed around the site via pipelines. On an annual basis, the process water circuit will be in a surplus condition and to avoid continuous and ongoing accruals of water the excess amounts will be released at a controlled rate each year during the wet season. Releases will be from the TMF into the Chilloroya River and will only occur upon the

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achievement of adequate quality. The wet season discharges will take place when the Chilloroya River has a large assimilative flow. The water will be treated if necessary before release. Non-process water is surface drainage and runoff water that will require management at the site. It will be comprised of two streams: contact and non-contact water. Contact water is water that will come into contact with facilities such as the pit, stockpiles, the plant site, PAG WRF, etc. This water will be collected in designated ponds for treatment, if necessary, prior to its release. To the maximum extent possible this water will be used in the process. Non-contact water will be directed around the facilities via diversion channels and discharged into natural drainages upon the achievement of sediment removal to acceptable levels. Hydrology and hydraulic analyses, including a detailed process water balance, have been completed to determine the types, locations and sizes of the water management structures as well as the times and quantities of discharges. The following sections discuss water management for the project. 1.9.1

Process water

Water for the mill will be supplied from the following sources: groundwater from dewatering wells located behind the pit walls, in pit surface water collected in the pit sumps, drainage from the PAG WRF that will be collected in the containment pond just below the PAG WRF, and reclaimed water from the surface water pond in the TMF. A temporary water reservoir will be available within the TMF in 2011, to collect and impound water to support the start of operations in early 2013. Raw water for sustaining mill operations will be obtained, in part, from pit dewatering wells installed behind the pit perimeter. Based on the current hydrogeologic model, a total of 18 dewatering wells will be installed in and around the pit. Water from these wells will be pumped to a collection box immediately upstream of the plant and then conveyed by gravity to a dedicated tank in the plant. Water from the PAG WRF and in-pit sumps is expected to be acidic and will be neutralised with lime at the process plant before use in the process. When the total volume of water available to the mill from the pit, TMF and PAG WRF exceeds the volume required, the excess amount will be removed from the pit groundwater component and a gravity flow pipeline will convey this from the pit dewatering wells system to the environment upon confirmation of adequate water quality. The priority will be to recycle as much water as possible from the PAG WRF, in pit sumps and TMF. During periods of excess water discharges from the TMF surface water pond, the water will be pumped to a TMF buffer pond located immediately north of, and below, the eastern side of the TMF embankment and then from this pond by gravity flow to the Chilloroya River. Such discharges, when required, will only take place during the wet season, which is between December and April. Results of the current water quality analyses and water balance modeling indicate that during these periods, when runoff flows into the TMF are higher, no water treatment will be necessary. The excess water removal system will be provided with instrumentation to monitor water levels and water quality and releases will only occur upon confirmation of adequate quality.

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1.9.2

Water balance

A site-wide process water balance model was developed for the project. This model evaluates the amount of stored water in the various components of the mine, as well as water inflows and outflows between the mine facilities on a monthly basis throughout the life of mine. The main objectives of the process water balance are to provide operating guidelines, design flows, and information inputs for the design of various water management facilities. The results also provided input to the environmental management plans for the project. Specific key objectives of the site wide water balance together with results are: 

Evaluate the range of normal operating volumes for the TMF surface water pond; this information is used to determine the pond configurations throughout the life of the facility. The predicted normal operating volumes for the TMF surface water pond vary between 1.8 and 3.3 Mm3.



Evaluate the maximum potential volume for the TMF pond above its normal operating volume to reflect conditions of extreme precipitation; this information is used to establish the minimum required freeboard heights for the TMF embankment at any time. The maximum operational volume for the TMF surface water pond may reach approximately 3.7 Mm3 after a 100 year/24 hr precipitation event.



Evaluate the amount of water to be discharged from the TMF surface water pond to the Chilloroya River via the TMF buffer pond; this information provides the design criteria for the discharge facilities as well as for the analysis of any downstream environmental impacts and mitigation plans. Under extreme precipitation conditions the amounts of water discharged via the TMF buffer pond may reach up to 3000 m3/hr. These discharges will occur only during the wet season and after adequate water quality has been confirmed.



Evaluate flows reporting to the PAG WRF containment pond to determine the required capacity of the containment pond and the pumping recirculation rates under conditions of extreme precipitation. The capacity of the containment pond is 600 000 m3 below a freeboard of 4 m, or 800 000 m3 at the pond crest. Pumping requirements for recirculation under extreme conditions may reach 1100 m3/hr.



Evaluate approximate flows from the in pit sumps and PAG WRF containment pond to determine the needs for treatment at the process plant, as well as the pumping requirements. Normal operating flows from the in pit sumps are expected to be 0 to 400 m3/hr and from the PAG WRF containment pond are expected to be 100 to 800 m3/hr, due to seasonal fluctuations. Maximum operating flows are expected to be between 1200 m3/hr and 1800 m3/hr respectively.



Evaluate the amounts of fresh water from the pit dewatering wells to the operating system. Operating flows expected from the dewatering wells are expected to be at least 330 m3/hr.

1.9.3

Non-process water

Contact water Non-process contact water includes water that has been in contact with areas that generate sediment loads and/or requires treatment before it is released. Diversion channels will be constructed to convey contact water to designated ponds or treatment areas. Contact water with sediment loads will be directed to sediment ponds before being released to natural drainages; these ponds include the pit

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sediment pond, the main sediment pond, the plant construction sediment pond:, and the crusher sediment pond. Contact water that will require additional treatment will be managed as follows: 

Plant contact water: during operations, water in contact with the plant area will be directed to a dedicated plant contact pond. From this pond, water will be directed for treatment before it is released to a natural drainage, or alternatively it will be directed to the mill process pond. Contact and non-contact water channels will be constructed around the plant to reduce the amount of water directed to the plant contact pond.



PAG WRF contact water: water in contact with the PAG WRF will be directed to the PAG WRF containment pond and will not be released to the environment. Excess water such as water from extreme precipitation events, will be recirculated onto the WRF. To reduce erosion and sediment loads, temporary diversion channels will be constructed as the PAG WRF is expanded.

Non-contact water Non-contact water will be directed around the mining facilities via non-contact water channels from and to natural drainages. Two non-contact water channels are currently contemplated, one upstream of the PAG WRF and the other upstream of the pit. All water channels will be lined for erosion protection with HDPE geomembrane (temporary channels) or rip-rap (permanent channels). The channels have been sized based on average surface runoff volumes and flows, as well as the 100 year storm event. Sufficient freeboard will be included in the permanent channels to contain flows associated with a 500 year storm event. 1.10

ENVIRONMENTAL AND SOCIAL CONSIDERATIONS

1.10.1

ESIA

The Ministry of Energy and Mines (MINEM) is the principal regulatory agency responsible for permitting mining projects in Peru. The project has been designed to consider all relevant legislation applicable to the development of mining projects in Peru including mines, roads, port and transmission lines. Additional legislation that has been considered includes legislation and regulations regarding archaeological areas of significance, endangered and protected species as well as community relations and public disclosure programs. Knight Piésold is undertaking an ESIA for the Constancia Project, which includes baseline studies which have now been completed. Results from the baseline studies indicate: 

Air quality within the project site and the surrounding communities is generally good.



Existing levels of noise and vibrations from static sources are within national standards.



The site is classified as an area with medium seismicity.



Soils are colluvial–alluvial and residual materials. In general the erosion potential of the soil in the area surrounding the project is low to medium.

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

Water quality: in general, results indicate that the water quality within the project area has neutral to alkaline pH and in some locations exceeds the national standards for iron, manganese, copper, lead, zinc and arsenic.



Flora: a total of 494 different species of plants and 10 vegetative formations were identified within the project area. While 18 species of flora are classified as endangered these are not located within the direct area of influence.



Terrestrial fauna: a total of 96 different species of birds, 19 species of mammals, four species of reptiles, and four species of amphibians were identified within the project area. Of these, five species of birds, three species of mammals and one amphibian are listed as protected species.



Aquatic life: three species of trout and two species of catfish (bagre and challhua) were found within these aquatic environments. Challhua was found only in wetland areas. One species of fish belongs to the IUCN Red List of Threatened Species.



Human interest environment: Landscape: some parts of the district are considered to contain medium values of visual quality, due to the presence of lakes and the dominant mountain landscapes in these sectors. Archaeological Heritage: a total of 46 archaeological sites were identified in the area of the future mine site. The process to obtain a certificate of non-existence of archaeological remains of significance from the INC (National Cultural Institute) has been initiated. Environmental liabilities from Previous and Current Mining Activities: five zones were identified where mining-related environmental liabilities exist on Norsemont mining concessions. Traffic: current levels of traffic will be assessed from a survey by SIGT that will identify vehicle types, the daily volume of traffic and the kinds of loads being transported along the existing road networks.

1.10.2

Stakeholder mapping

Stakeholder mapping has been undertaken, identifying the two principal stakeholder groups in the direct area of influence as the communities of Uchucarco and Chilloroya. There are three well-defined groups in each of the communities, namely: 

Community assembly: people from the community



Artisanal miners: from the community and possess land in the community (do not own title of the land).



Youth Groups: (in the case of Chilloroya, represented through the Youth Association), they have higher education than the rest of the local residents and the experience of having lived in other cities; however, they own no land and possess no economic capital.

The youth groups and miners do not have a solid organisational structure, but do have marked interests and represent a significant portion of the population. The association of artisanal miners in Chilloroya is one year old. Benefits from the artisanal mining activities in Chilloroya are not distributed evenly across the population, but are primarily experienced by those members of the population who possess the land where the illegal mining activities are occurring.

Page 55


Uchucarco has a greater number and variety of organisations inside its community structure; some of 1 which have duties that are independent from the Governing Board, these include: JAAS , the Committee of Water Users (“Comité de Regantes”), among others. In addition, Uchucarco is the most 2 active community in the District of Chamaca , and has more relationships with district and private institutions. In Chilloroya, unlike Uchucarco, the political strength of the Governing Board is weak and the different associations in the community are not well organised. Although there is a varied array of organisations, many are focused on specific tasks and have little political presence, as in the case of women’s and parents’ organisations. 1.10.3

Status of land ownership in the communities of Uchucarco and Chilloroya

The land administration system in Uchucarco and Chilloroya is consistent with the community system of land possession, under which communities have the right to decide upon their territories. These decisions must be made and undertaken by consensus, through a General Assembly of the community. Typically land is divided into plots among the dwellers (as in the case of Chilloroya and Uchucarco), and boundaries marked in the presence of members of the governing board. These plots of land are registered with the community. The dwellers do not have official title to the lands but can enjoy the benefits from the land (growing crops, grazing of livestock, construction of a house or dwellings) etc. The dweller is not permitted to make any transaction or exchange of the land. The project has purchased 4097 ha of private lands to date and has plans to purchase additional lands belonging to the two communities. As a result of the project development, approximately 35 families from the community of Chilloroya will need to be relocated. Norsemont is currently preparing a Resettlement Action Plan (RAP) in accordance with IFC performance standards on involuntary resettlement and Peruvian National legislation. 1.10.4

Impact identification and evaluation

Positive and negative impacts related to the project development phases will be identified and evaluated. Mitigation measures will be proposed and evaluated for their relative level of impact and significance. The process for the evaluation of impacts will be guided by national and international standards, using impact assessment matrices and predictive models. The evaluation of impacts will consider direct, indirect and cumulative impacts. Modelling of noise and vibration, air dispersion, water quality and visual impacts will be undertaken to evaluate the impacts on the local population. 1.10.5

Environmental management plan

As part of the ESIA, an Environmental Management Plan (EMP) will be developed to: 

Identify mitigation and management strategies

1

Water and Sanitation Administration Board (“Junta de Administración de Agua y Saneamiento”)

2

Uchuccarcco is one of the most densely populated districts of Chamaca, in addition to having a privileged economic

standing given the recovery of the mining activity (artisanal and big mining) and the development of livestock breeding.

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

Set objectives and targets



Define performance indicators



Document time frames to achieve targets



Allocate responsibilities and identify the necessary resources for the implementation of the plan



Establish mechanisms to monitor, evaluate and report on progress.

EMPs are important tools for ensuring that the management actions arising from the ESIA are clearly defined and implemented throughout all phases of the Project life cycle. 1.10.6

Resettlement Action Plan (RAP)

A Resettlement Action Plan (RAP) will be developed as part of the ESIA. The RAP will comply with international best practice for involuntary resettlement as promulgated by Equator Principles and IFC Performance Standards. The RAP will specifically address compensation procedures and measures for people undergoing physical and/or economic displacement as a result of project implementation, and will include: 

Census of project-affected people



Cut-off date



Compensation matrix



Framework for development of detailed resettlement actions

1.10.7

Community relations plan

The Community Relations Plan will be prepared taking into account the following requirements: 

Final determination of the Direct and Indirect Areas of Influence of the project, based on the outcomes of the Social Impact Analysis.



The needs in the construction and operation phases, as determined in the Social Impact Analysis.



The State’s requirements, as expressed in the Community Relations Guidelines of the MINEM and the Prior Commitment Act (DS-042-2003-EM).



The requirements of international financial institutions, taking into account as main references the Equator Principles, IFC Performance Standards, APELL for Mining, and, supplementary, social management standards.

Specific social programs will be designed for the mitigation and prevention of identified impacts. Generally, these programs address the following key topics: 

Communication and consultation



Participatory monitoring



Claims and dispute resolution



Local employment

Page 57




Social investment



Land acquisition and resettlement



Code of conduct for workers.

1.10.8

Health, safety and environment (HSE) management and monitoring plan

Norsemont will develop a comprehensive HSE management plan for the Constancia Project to: 

Ensure HSE compliance



Demonstrate that all hazards are appropriately managed



Achieve continuous improvement in HSE performance.

Periodic and regular monitoring constitute a principal component of the HSE plan for the construction, operation, closure and post closure phases of the project. The plan includes direct monitoring of air and water resources, and indirect monitoring of flora and fauna. The monitoring program will provide information for evaluating actual project impacts and the effectiveness of the mitigation measures in place. This will allow for dynamic adjustments to the mitigation plan as the project progresses. Six air quality monitoring stations have been proposed: four adjacent to the open pit, and one each in Uchucarco and Chilloroya. Strict measures to maintain air quality will be implemented. This will involve, for example, spraying water on access roads for dust control to ensure compliance with legislation regarding airborne particulates. Twelve surface water quality monitoring stations have been proposed in the area. The TMF, PAG WRF, topsoil and unsuitable material stockpiles and all associated ponds will be instrumented for performance monitoring. This will include pore pressures in the tailings and mine waste, as well as in the drain zones and embankment structural zone and foundations. Water flow rates and totalised volumes will be measured as will the water and tailings levels. Monitoring of slope movements and materials settlement will be made. Industrial and hazardous waste will be separated from common domestic waste. Domestic waste will be recycled whenever feasible. Norsemont wIll construct a sanitary landfill for the disposal of domestic wastes. Hazardous waste will be stored temporarily in secondary confinement areas prior to removal to designated facilities for resale, recycling or definitive storage, in accordance with Peruvian regulations. In all cases, storage facilities for fuel and chemical substances will be designed with secondary impermeable containment. These lined and bermed containment areas will be designed to hold 110% of the capacity of the largest tank to avoid spillage of contaminants. No underground storage facilities for fuel are planned.

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Sewage treatment facilities will be constructed servicing all project components. Operating procedures, including monitoring discharges, will comply with the corresponding Peruvian standards. Health and safety procedures will be developed in accordance with Peruvian legislation and will be strictly enforced. A restoration program will be developed to re-establish a landscape that is environmentally and aesthetically compatible with the surrounding countryside. All personnel and contractors will be required to comply with the standards and procedures contained in the ESIA for all Project stages. Internal and external audits will be performed periodically to verify compliance. 1.10.9

Closure plan

In accordance with Peruvian National Regulations for the mining sector, a conceptual closure plan will be developed for inclusion in the ESIA. The conceptual plan includes the principal impacts from project closure to the communities in the area of influence, and identifies measures to mitigate these impacts. Reclamation and closure of the project will be conducted in accordance with international best practices, the objective being to return mined lands to conditions capable of supporting prior land use or uses that are equal to or better than prior land use to the extent practical and feasible. In addition, longtem stability and safety issues will be addressed as a priority. The area to be disturbed and reclaimed encompasses approximately 796 ha. Reclamation and closure activities are to be conducted concurrently with mining operations, to the extent practical, to reduce the final reclamation and closure costs and minimise long term environmental liabilities. The key goals of reclamation and closure are to ensure the physical and chemical stability of the TMF and the WRF. The closure plan includes: 

Salvaging, stockpiling and ultimate replacement of topsoil and subsoils in the PAG WRF and process plant areas.



Reclamation of the open pit, including re-contouring, re-grading and re-vegetating of an approximately 30 m wide area surrounding the open pit. After closure the pit lake water will need to be treated in perpetuity, by lime treatment of water prior to discharge into the Chilloroya during the wet season.



WRF: closure of the PAG WRF will include covering the facility with an approximately 1.8 m cap comprising a low permeability layer, a drain layer and 0.3 m of topsoil/growth media. The soil cover will be scarified and seeded.



The upstream diversion channel of the PAG WRF will remain in place at closure. The current closure concept is to pump the seepage and run-off from the WRF to the pit for in-pit treatment, as discussed above.



TMF: the current closure plan is based around de-sulphurising (by pyrite flotation) plant tailings during the last one and half to two years of operations, to form inert tailings. These inert tailings

Page 59


will then be deposited over the existing tailings surface to form an impermeable cover. The pyrite float material will be stored in a lined area of the TSF. A spillway will be constructed in the southeast side of the TMF that has been designed to pass the PMP. A soil cover will be placed on top of the tailings beach and the facility and the dam face will be revegetated. 

Roads will be reclaimed by pushing safety berms down and over roads, removing culverts, backfilling ditches, and re-grading areas to re-establish the natural drainage system. After regrading is completed, road areas will be scarified and seeded.



Infrastructure will be removed from the project area when no longer needed. Concrete foundations will be buried in place, and scrap metal removed. The mill will be decontaminated. Associated yard areas will be ripped to eliminate compacted soils and regraded, after which previously disturbed areas will be scarified and seeded.



Production water wells will be abandoned in accordance with local regulations or transferred to support an approved post-mining land use. Monitoring wells will be abandoned once regulatory officials decide they are no longer necessary for monitoring purposes. Water lines, utility poles, power lines, fuel tanks, generators, transformers and other items remaining in the project area after mine operations cease will be removed from the site and disposed of properly unless they can be used by the communities, or sent to salvage. The non-hazardous sanitary land fill will be closed by placing an inert cap over the facility and removing any infrastructure (fences, platforms, etc.).

Annual reports will be prepared to document the closure and reclamation activities. Revegetation efforts will be monitored biannually by a range specialist to record vegetation success, monitor erosion, and modify reclamation plans if necessary. Groundwater wells and surface water sites will be sampled quarterly to record post-mining water quality. Closure and reclamation activities are anticipated to take place over a five year period. The total estimated cost, is approximately $38.23 M. 1.11

PROJECT IMPLEMENTATION PLAN

1.11.1

Approach and strategy

A project implementation plan has been developed as part of the Study. An Owners team will be formed to deliver the project through the engagement of an EPCM contractor and specialist engineering consultants, suppliers and Peruvian construction contractors. The Owners team will consist of specialist delivery personnel sourced locally and using expatriate resources where appropriate. The Owners team will develop policy and ensure its implementation and compliance through the consultancy, supplier and contractors’ systems in areas such as safety, environment, human and industrial relations and security. The EPCM contractor will be responsible for the delivery of the process plant and associated infrastructure area. The EPCM contractor will provide the underlying framework for systems and procedures for the Owner’s team.

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The Peruvian construction industry has sufficient capability and capacity to carry out the construction works. Foreign contractors would only be used in either high risk or specialist work areas, for example mill and drive construction, lining etc. Fixed price contracts are preferred; reimbursable contracts, where used, will be structured with incentive clauses to encourage performance. Both practices support a reduced level of performance management by the Owner’s team and the EPCM contractor. The Owners team will deliver the mining development work, and ultimately the mine operations. Tailings, waste rock, topsoil and water management facilities will be engineered by a single consultant to minimise interface management. Bulk earthworks across the site will be delivered by a civil construction fleet purchased and managed by the Owner’s team. This fleet will commence construction early in the project, with the majority of the fleet retained through to operations to continue with subsequent construction stages of the tailings management facilities. Construction of the balance of the tailings, waste, topsoil and water management works will be contracted to local specialist contractors. The access road, power supply and accommodation camp will be completed by specialist Peruvian construction contractors. 1.11.2

Quality assurance

A project Quality Plan will be developed for the Project. This Plan will form an integral part of the overall Project Execution Plan and detail how quality assurance will be achieved during engineering, procurement, construction and commissioning. 1.11.3

Project implementation schedule

A project implementation schedule is summarised in Figure 1.9. The schedule shows total project duration of approximately 39 months, including detailed design, procurement, construction and commissioning.

Page 61

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 1.9 Project Implementation Schedule Year Quarter

2009 Q3

2010 Q4

Q1

Q2

2011 Q3

Q4

Q1

Q2

2012 Q3

Q4

Q1

Q2

Q3

Q4

APROVALS Finance Approval ESIA Construction Permit MINE Planning Tender and Award Fleet Supply Assembly /Training Strip and First Ore CIVIL CONST. FLEET Planning Tender and Award Fleet Supply PLANT Design Long Lead Procurement Construction Commissioning TAILINGS STORAGE Design Mine Access Road and Camp Process Plant Pad Mine Area Pad Tailing Management Facility – Initial Stage Tailing Management Facility – Reservoir Fill Tailing Management Facility – Embankment Other Operational Facilities Waste Rock Dump ACCESS ROAD Detailed Design Tender and Award Construction ACCOMMODATION Design and Tender Procurement Complete 500 rooms Complete 1000 rooms Complete 1500 rooms Complete 1800 rooms

X X X

POWER SUPPLY Design Tender and Award Long Lead Procurement Construction

Page 62


Key dates for the project are as follows: 

Project Execution commencement – Q4 2009



Environmental approval – Q4 2010



Construction permit and commencement of construction – Q4 2010



Completion of Process Plant commissioning – Q4 2012.

Critical activities that need to commence upon Norsement Board approval are: 

Commence negotiations to secure a power supply agreement or option



Commence detailed design for the access road upgrade and investigate acceleration of its delivery



Commence the Front End Engineering Design which includes process design, flowsheet verification and optimisation, plant layout, and long lead item procurement



Commence detailed design and planning for bulk earthworks and purchase of Owners civil construction fleet



Commence recruitment of key Owners team members



Commence development of the accommodation camp contract and investigate availability of second hand camps



Investigate availability of long lead equipment, i.e. cancelled orders etc



Commence development of project systems. This includes OHS&E requirements and standards, equipment numbering, asset numbering, document numbering, cost control and reporting systems, document control, permit plan development and implementation, and procurement documentation and systems.

1.11.4

Civil construction fleet

A civil construction fleet, owned by Norsemont, will be utilised to construct bulk earthworks associated with the TMF, PAG WRF, water management structures and internal access roads. The basis of the fleet is conventional diesel-powered hydraulic excavators and off-road haul/dump trucks, together with supporting equipment such as FELs, dozers, graders, compactors, water trucks, service trucks, and the like (Table 1.11). Many of these items will be pre-purchased by Norsemont and will service both the mine and construction fleets. Pre-construction earthworks will take place over an approximate two year period with the TMF embankment construction continuing for an additional fourteen years. Most of the construction fleet will be sold after three years with the remaining equipment used on the construction of the TMF embankment and for plant field maintenance operations. The mine truck fleet will be used to provide the embankment material after the construction fleet has been demobilised. A separate contractor will be responsible for constructing the underdrains, geosynthetics, pipework, and small peripheral construction activities during the first three years, as well as the smaller earthworks in the latter years of the mine life.

Page 63


Aggregate material will be provided for the TMF and PAG WRF, and possibly surfacing rock for mine haul roads. This material will be obtained from quarries using drilling and blasting procedures. Table 1.11 Civil Construction Fleet Type

Equipment Class

Fleet Units

Operating Hours

Operating Costs

Unit Purchase Price

(Hrs/yr)

($/hr)

($M)

Hydraulic Excavator

382 kW

1

5304

149.61

0.827

Wheel Loader with Rock Bucket

597 kW

1

5304

135.18

1.952

LPG Dozer with VR Blade and Ripper

192 kW

1

5304

47.16

0.478

Dozer with CU Blade and Ripper

482 kW

1

5304

92.23

1.227

Soil Compactor

299 kW

2

5304

43.86

0.644

Vibratory Soil Compactor

130 kW

1

5304

43.86

0.113

Articulated Truck

30 T

1

5304

30.85

0.413

Dump Truck

60 T

8

5304

149.02

1.004

Light Duty Dump Truck

19T

1

2652

73.01

0.070

Fuel and Lubrication Truck

43 T

1

1326

8.12

0.585

Water Truck

60 T

2

5304

149.02

1.166

Service Truck

1

1326

8.72

0.175

Tire Truck

1

1326

10.77

0.175

1

5304

87.79

2.083

Motor Grader

397 kW

Utility Backhoe – 4WD

102 kW

1

1856

50.50

0.156

¾ ton Pickup Truck, 4WD

6

5304

5.27

0.027

Onan Light Plant

6

2122

1.68

0.027

Hydraulic Excavator

200 kW

2

5304

131.40

0.445

Carry Dozer with Ripper

698 kW

2

5304

118.85

1.753

Wheel Loader

260 kW

1

5304

39.02

0.395

Motor Grader

221 kW

1

5304

37.72

0.706

1.12

PROJECT OPERATIONAL PLAN

Site operations will generally be run by the Owners staff. Mining will be done by an Owner-operated mining fleet, and the plant operated and maintained by the Owner. Although the plant will have fully equipped maintenance workshops, where possible, maintenance will be performed off-site by suitable contractors (e.g. motor re-winds). Supply will be outsourced for: 

fuel and lime



transport of concentrates



storage and ship-loading of copper concentrate.

Page 64


The manning structure will be made up of management and administration staff, technical staff, operators, trades and general labour. The labour numbers will be adjusted to satisfy local requirements. It is estimated that total numbers employed on the site will approach 800. The employment philosophy is based on preferential hiring of local people to fill general labour and operating positions. Preference would be given to Peru Nationals to fill managerial and technical positions, although it is anticipated that some of these may have to be filled by expatriate staff initially. Training is considered to be a key activity before plant start-up, with the need to prepare the local workforce in areas of plant operation, maintenance, operating procedures and workplace health and safety. An Environmental and Social Management Plan will be implemented as part of the overall Operation Plan. This is currently under development by Norsemont, and forms part of the ESIA. 1.13

CAPITAL COST ESTIMATE

1.13.1

Initial project capital

Total initial Project capital cost is estimated to be $845.99 M, with an additional requirement of $147.64 M over the life of the operation (Table 1.12). The capital cost estimate has an accuracy of ±15% and is expressed in Q1 2009 US dollars. No provision has been made for escalation, taxes and project funding. Table 1.12 Constancia Project Capital Cost Estimate Description

Initial

Sustaining

Total Capital

Capital ($M)

Capital ($M)

($M)

Mining

120.51

45.42

165.93

Process plant and associated infrastructure

367.18

0

367.18

Waste and water infrastructure

84.17

46.74

130.90

Offsite infrastructure

87.65

0

87.65

access road

20.17

0

20.17

accommodation camp

43.02

0

43.02

HV power supply

24.45

0

24.45

Owners civil construction fleet

69.39

0

69.39

Project contingency

74.13

0

74.13

Owners costs

42.96

55.47

98.44

845.99

147.64

993.62

Total

GRD Minproc has estimated costs for mining, plant and on-site infrastructure. Knight Piésold estimated costs associated with the civil construction fleet, TMF, site-wide water management and waste rock disposal, while SIGT and Cesel provided capital costs for the access road and power supply, respectively. The cost of the accommodation camp was based on a budget quote from an experienced

Page 65


Peruvian contractor. Norsemont provided the Owners cost, which includes allowances for recruitment and manning, purchase and operation of the civil fleet (with inputs from Knight Piésold and GRD Minproc), insurance, marketing, land acquisition and taxes, etc. The plant and on-site infrastructure capital costs include provisions for engineering, procurement, construction and commissioning. Within the plant and associated infrastructure area, accuracy provisions have been applied within the cost estimate by discipline to allow for anticipated increases arising during final design, based on GRD Minproc experience. The EPCM estimate has been developed based on experience with other similar projects in South America. Mine capital costs are built up from vendor budget pricing and are inclusive of freight to site, site assembly, commissioning, and training of personnel. Mining equipment includes electric shovels, a support FEL, a fleet of 220 t class dump trucks plus track dozers, graders and other mine support equipment. A risk-based contingency estimate was developed collaboratively by GRD Minproc, Norsemont and Knight Piésold. At a P80 confidence level, the contingency on capital cost is $74.13 M. 1.13.2

Sustaining capital

Sustaining capital of $147.6 M allows primarily for replacement equipment for the mining fleet and the costs associated with expansions to the TMF dam over the project life. Additional sums are included for replacement of light vehicles, computer hardware and software, etc. 1.14

OPERATING COST ESTIMATES

The operating cost estimate for the entire project is summarised in Table 1.13. The estimate has an accuracy of ±15% and is expressed in Q1 2009 US dollars. No provision has been made for escalation. Table 1.13 Summary Operating Cost Estimate Area

LOM Average Annual Cost $

LOM Average $/t Milled

LOM Average $/lb Cu Metal

Mining

48 405 954

2.62

0.32


71 216 612

3.85

0.48

General and administration

11 144 200

0.60

0.07

Off-site costs (refining, smelting, transport)

53 533 883

2.89

0.36

Civil Construction Fleet

3 565 905

0.19

0.02

Royalty

7 980 470

0.43

0.05

195 847 024

10.59

1.31

Total

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On a LOM Project basis, the unit operating cost for the Constancia Project is estimated to be $10.59/t ore, or $1.31/lb of copper produced excluding by-product credits. After inclusion of credits for sale of molybdenum, silver and gold the unit cost reduces to $0.92/lb of payable copper. 1.14.1

Mine operating cost

The mining operating costs were determined by GRD Minproc, based on a first-principles build-up of load, haul, drill, blast, support equipment, labour and other costs (Table 1.14). Average cost (excluding pre-strip) is $1.18/t mined, but it varies over the operating life, as was seen in Figure 1.4. Table 1.14 Mine Operating Costs Area




Load

3 616 008

0.20

0.02

Haul

19 420 109

1.05

0.13

Drill

3 142 204

0.17

0.02

Blast

10 143 674

0.55

0.07

Support Equipment

7 770 704

0.42

0.05

Labour

3 518 724

0.19

0.02

Other

794 532

0.04

0.01

Total

48 405 955

2.62

0.32

Significant mining operating cost drivers include diesel price, truck tyre life, power cost for electric shovels and explosives costs. 1.14.2

Plant and infrastructure costs

Process plant and on-site infrastructure operating costs were estimated by GRD Minproc, with input for operating the TMF, waste and water management supplied by Knight Piésold. Costs were determined based on labour, mobile equipment fuel, reagents, power and grinding consumables requirements (Table 1.15).

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Table 1.15 Plant and Associated Infrastructure Operating Cost Estimate Area




Fuel & Miscellaneous

1 424 332

0.08

0.01

Labour

2 136 498

0.12

0.01

Maintenance materials

4,985,163

0.27

0.03

Reagents

9 258 160

0.50

0.06

Power

25 637 980

1.39

0.17

Grinding consumables

27 774 479

1.49

0.19

Total

71 216 612

3.85

0.48

The plant operating cost averages $3.85/t ore treated over the LOM. 1.14.3


The G&A cost estimate includes both site and Lima office costs, and is summarised in Table 1.16. Table 1.16 G&A Operating Cost Estimate Area




Labour

4 538 000

0.24

0.03

Administration

2 229 200

0.12

0.01

Other Costs

4 377 000

0.24

0.03

Total

11 144 200

0.60

0.07

The estimate includes provision for management, administration, community relations, environment and safety, medical, security, procurement, camp and general charges. Also included are the costs associated with transport of labour to and from site, insurance, communications, permits, fees and other government charges, directors fees, consultants, software, road, powerline and camp maintenance. The G&A operating cost averages $0.60/t ore treated over the LOM. 1.14.4

Off-site charges

Off-site costs include transport, port charges, smelting costs and penalties. These costs have been estimated by GRD Minproc or provided as input from Norsemont. The average LOM off-site cost is estimated at $2.89/t of ore treated, or $0.36/lb payable copper, as shown in Table 1.17.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 1.17 Off-site Operating Costs Area Shipping (inc Port Charges) Transport Smelting & Refining Penalties Total

1.14.5




11 777 454

0.64

0.08

9 100 760

0.49

0.06

31 584 991

1.71

0.21

1 070 678

0.06

0.01

53 533 883

2.89

0.36

Royalties

Royalties comprise a state royalty ranging from 1% to 3% of NSR depending on the NSR value, plus a 0.5% NSR royalty, capped at $10 M, for the Minera Livitaca and Katanga properties. Annual royalty payments average $7.98 M, equivalent to $0.43/t ore treated, or $0.05/lb copper produced (Table 1.18). Table 1.18 Royalty Costs Area State royalty Minera Livitaca & Katanga royalty Total

1.15




7 300 401

0.39

0.05

680 069

0.04

0.00

7 980 470

0.43

0.05

MARKETING, PRODUCT PRICING AND TREATMENT CHARGES

Norsemont has not yet entered into negotiations with potential purchasers of copper and molybdenum concentrates, but has reviewed market reports on copper pricing, treatment and refining charges, and penalties. For the purpose of the DFS, it is assumed that concentrates will be shipped to smelters in Asia for treatment. Norsemont has adopted a single average price of $2.00/lb Cu, $12.00/oz Ag, $800/oz Au and $13.00/lb Mo over the life of the Project for Base Case economic assessment. These prices are less than current spot prices. In common with metal prices, smelting and refining charges have varied widely over the past year, and prediction of future costs is difficult. Norsemont has adopted figures of $65/t and $0.07/lb. Standard metal refining deductions have been applied.

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1.16

PROJECT FINANCIAL ANALYSIS

1.16.1

Background

GRD Minproc and other consultants provided the capital costs, operating costs and production plan inputs, and prepared the pre-tax cashflow. Norsemont provided inputs regarding taxation, metal prices, royalties and off-site charges, and prepared the post-tax cashflow and financial analysis. In estimating ramp-up, it was assumed that design metal recovery would be achieved eight months after production started. Norsemont advised the basis for working capital was two month debtors. Operating cost inputs to the financial model are summarised in Table 1.19, metal price assumptions in Table 1.20 and the production schedule in Table 1.21. Table 1.19 Operating Cost Inputs to Financial Model Units Average Mining Costs

Processing Costs

Hypogene Supergene Skarn 1 Skarn 2 High Zn

$/t mined $/t ore processed

1.18 2.62

$/t ore processed $/t ore processed $/t ore processed $/t ore processed $/t ore processed

4.31 3.26 2.84 3.13 3.24

Mining Royalties

<$60M >$60M & <$120M >$120M Minera Livitaca & Katanga cap (max payment)

%NSR %NSR %NSR %NSR $M

1.00% 2.00% 3.00% 0.50% 10.00

Transport & Shipping Charges

Copper

Road Transport Port Charges Shipping Costs Insurance Transport & Shipping Losses

$/wmt $/wmt $/wmt $/wmt %

32.30 7.50 35.00 1.78 0.50%

Molybdenum

Road Transport Other Charges Insurance Transport & Shipping Losses

$/wmt $/wmt $/wmt %

77.57 5.86 1.78 0.00%

Copper

Treatment Charge Price Participation Upper Escalator Lower De-escalator Refining Charges Cu Ag Au Treatment Charge

$/dmt

65.00

Fixed

$M/a

Treatment & Refining Charges

State

Value

Molybdenum G&A

$/lb % $/lb % $/lb $/oz $/oz $/dmt

1.20 0.90 0.07 0.40 1.20 1,461.65 11.14

Page 70

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 1.20 Metal Price Assumptions in Cashflow Model Units Metals Prices

Copper Silver Gold Molybdenum

$/lb $/oz $/oz $/lb

Value 2.00 12.00 800.00 13.00

Page 71

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 1.21 Production Schedule for Cashflow Model Units Mine Production Ore

Value

2012 -1

2013 1

2014 2

2015 3

2016 4

2017 5

2018 6

2019 7

2020 8

2021 9

2022 10

2023 11

2024 12

2025 13

2026 14

2027 15

m dmt

326.79

0.82

21.95

22.97

23.14

24.59

25.17

24.62

16.89

22.69

23.21

18.79

19.34

24.96

30.57

22.65

4.43

Cu Zn Mo Ag Au Pb Sulphur

% % g/t g/t g/t % %

0.39% 0.11% 107.75 3.46 0.04 0.04% 2.40%

0.37% 0.07% 42.85 3.77 0.08 0.05% 0.53%

0.58% 0.10% 89.79 4.23 0.07 0.04% 2.28%

0.58% 0.11% 116.42 4.43 0.06 0.04% 3.01%

0.63% 0.15% 152.41 5.07 0.05 0.04% 3.40%

0.46% 0.10% 98.35 4.04 0.04 0.06% 2.42%

0.42% 0.10% 88.24 3.58 0.06 0.04% 2.84%

0.39% 0.14% 96.89 3.63 0.05 0.04% 2.86%

0.34% 0.19% 106.79 3.75 0.04 0.06% 2.55%

0.32% 0.12% 110.23 3.15 0.03 0.04% 2.65%

0.33% 0.11% 158.69 3.17 0.03 0.04% 2.29%

0.34% 0.12% 154.38 2.99 0.03 0.03% 2.07%

0.26% 0.05% 90.30 2.51 0.03 0.05% 1.60%

0.27% 0.09% 100.57 3.13 0.04 0.07% 2.05%

0.28% 0.09% 78.43 2.69 0.05 0.04% 2.10%

0.31% 0.06% 86.38 2.31 0.04 0.03% 1.70%

0.41% 0.03% 108.44 2.72 0.06 0.03% 0.86%

Waste TMM SR

m dmt m dmt

300.03 626.82

12.76 13.58

23.60 45.55

21.80 44.78

21.79 44.93

20.33 44.92

19.67 44.84

20.21 44.83

28.02 44.91

22.10 44.79

21.53 44.74

25.99 44.78

25.52 44.86

19.70 44.65

12.34 42.92

4.67 27.32

0.92

15.53

1.07

0.95

0.94

0.83

0.78

0.82

1.66

0.97

0.93

1.38

1.32

0.79

0.40

0.21

Stockpile

m dmt

1.17

8.91

7.04

12.76

23.32

43.27

17.74

16.27

21.76

27.87

30.02

33.24

41.56

55.10

Process Plant Feed

18.35 0.32% 0.18% 102.79 3.72 0.04 0.05% 11.53 2.56 0.92 3.34

17.20 0.37% 0.12% 125.13 3.44 0.04 0.04% 13.81 1.11 0.63 0.02 1.63

17.11 0.38% 0.12% 187.34 3.56 0.03 0.04% 13.98 1.04 0.53 0.00 1.56

16.64 0.36% 0.11% 164.90 3.16 0.03 0.03% 14.65 0.02 0.77 1.19

16.11 0.28% 0.04% 95.89 2.49 0.03 0.04% 15.58 0.14 0.38

16.64 0.31% 0.09% 120.18 3.47 0.05 0.07% 14.58 0.21 0.16 1.69

2031 19

2032 20

-

-

-

-

-

0.00% 0.00% 0.00% 0.00%

0.00% 0.00% 0.00% 0.00%

0.00% 0.00% 0.00% 0.00%

0.00% 0.00% 0.00% 0.00%

0.00% 0.00% 0.00% 0.00%

4.43

-

-

-

-

-

-

-

-

-

-

-

60.89

49.40

49.40

49.40

49.40

49.40

49.40

17.04 0.35% 0.10% 94.08 3.25 0.06 0.06% 13.88 0.13 1.10 1.93

16.86 0.36% 0.07% 96.09 2.52 0.05 0.03% 14.33 1.47 1.05

15.92 0.25% 0.04% 75.97 2.13 0.04 0.03% 15.92 -

0.00% 0.00% 0.00% -

0.00% 0.00% 0.00% -

0.00% 0.00% 0.00% -

0.00% 0.00% 0.00% -

0.00% 0.00% 0.00% -

277.39

-

21.64

21.89

21.45

21.05

19.64

% % g/t g/t g/t %

0.43% 0.11% 116.67 3.69 0.05 0.04%

0.00% 0.00% 0.00%

0.59% 0.10% 90.08 4.24 0.07 0.04%

0.60% 0.11% 119.81 4.50 0.06 0.04%

0.67% 0.15% 162.14 5.25 0.06 0.04%

0.50% 0.10% 108.16 4.35 0.05 0.06%

0.49% 0.10% 98.70 4.02 0.07 0.04%


m m m m m

163.55 71.03 14.59 4.73 23.49

-

0.72 18.18 0.04 1.07 1.62

4.40 13.36 0.89 1.50 1.74

5.24 10.18 3.58 0.55 1.90

6.54 11.28 1.36 0.87 1.00

9.58 6.69 0.88 0.70 1.79

19.86 0.44% 0.15% 107.54 3.96 0.06 0.04% 8.81 6.26 2.11 0.01 2.67

% % % % %

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

% % % % %

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

357,410

409,705

450,871

332,280

304,351

28.86% 3.55% 205.30 2.45 0.57%

28.77% 2.66% 192.40 1.83 0.49%

28.54% 2.72% 199.90 1.60 0.52%

28.57% 2.79% 220.27 1.77 0.91%

28.26% 2.64% 207.54 2.58 0.63%

287,649 27.18% 3.64% 218.82 2.30 0.75%

210,240 25.63% 5.11% 259.52 1.99 1.17%

211,969 27.02% 3.14% 223.20 1.75 0.82%

215,531 27.39% 2.99% 226.17 1.40 0.75%

196,298 28.12% 2.82% 214.12 1.38 0.57%

134,200 28.99% 1.48% 239.48 2.41 1.34%

163,779 27.58% 2.71% 281.94 2.79 1.83%

190,048 27.65% 2.85% 233.40 3.12 1.24%

183,100 28.63% 2.02% 185.69 2.58 0.62%

118,882 29.17% 1.59% 228.14 2.84 0.96%

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%


dmt dmt dmt dmt dmt

dmt % % g/t g/t %

3,766,311

Cu Zn Ag Au Pb Tonnes Mo

dmt

44,245

%

Payable Metal Cu Cu Ag Au

Mo

2030 18

m dmt

Concentrate Cu Tonnes

Mo

2029 17

Cu Zn Mo Ag Au Pb

Process Plant Recoveries Cu Hypogene Supergene Skarn 1 Skarn 2 High Zn Mo

t:t

2028 16

Mo

m m m m

lb's oz's oz's lb's

28.09% 2.93% 217.21 2.10 0.79%

2,237.79 22.55 0.13 39.02

0.00% 0.00% 0.00%

-

2,424

3,606

4,782

3,131

2,666

0.00%

40.00%

40.00%

40.00%

40.00%

40.00%

2,937 40.00%

2,594 40.00%

2,960 40.00%

4,406 40.00%

3,772 40.00%

2,124 40.00%

2,749 40.00%

2,204 40.00%

2,227 40.00%

1,663 40.00%

0.00%

0.00%

0.00%

0.00%

0.00%

-

218.36 2.00 0.02 2.14

249.53 2.13 0.01 3.18

272.37 2.45 0.01 4.22

200.90 2.02 0.01 2.76

181.99 1.73 0.02 2.35

165.20 1.74 0.01 2.59

113.58 1.54 0.01 2.29

121.00 1.31 0.01 2.61

124.77 1.35 0.00 3.89

116.77 1.16 0.00 3.33

82.34 0.90 0.01 1.87

95.51 1.32 0.01 2.42

111.11 1.24 0.01 1.94

110.98 0.91 0.01 1.96

73.39 0.75 0.01 1.47

-

-

-

-

-

Page 72


1.16.2

Summary

Post-tax analysis of predicted project cash flows for three metal price scenarios is summarised in Table 1.22. Table 1.22 Constancia Project After Tax Analysis Summary Post Tax

Commodity Price Scenarios

Analysis

Case 1

Case 2

Case 3

Cu Price ($/lb)

2.00

2.25

2.75

NPV (5%)

$496.8 M

$731.9 M

$1277.5 M

NPV (8%)

$303.6 M

$494.2 M

$931.8 M

IRR

15.5%

19.4%

26.9%

Payback

4 yrs

3 yrs

3 yrs

* NPV is quoted after taxes, royalties and profit sharing, and sunk costs.

Case 1 (Base Case): for NI-43-101 reporting purposes, Norsemont has elected to use the following long-term commodity prices: copper $2.00/lb, molybdenum $13.00/lb, silver $12.00/oz and gold $800/oz. Case 2: $2.25/lb copper represents the approximate mean analysts’ long-term copper price assumptions. Other metals remain constant. Case 3: $2.75/lb copper represents the mean copper price forward curve through to mid-2011, which historically has been the most accurate indicator of long-term copper prices. Other metals are based on recent metal prices of $18.00/lb Molybdenum, $14.00/oz silver and $950/oz gold. GRD Minproc provided the capital costs, operating costs and production plan inputs for the financial analysis and prepared the pre-tax cashflow, but expresses no opinion on the post-tax cashflow and financial analysis prepared by Norsemont. The financial analysis and discussion in the following sub-sections are based on the Base Case metal price assumptions outlined above. 1.16.3

Pre-tax analysis

The cashflow projections and financial evaluations assume that capital works commence Jan 2010 and that ore treatment commences January 2013 running to December 2027. Escalation and inflation have been excluded. The financial analysis assumes a 100% equity basis. Based on the assumptions set out above, the project pre-tax NPV at an 8% discount rate is $529.3 M and pre-tax IRR is 19.5%. The project would be expected to pay back the initial capital three years after commencement of production (pre-tax, undiscounted basis).

Page 73


The cash break-even copper price is $0.98/lb. The economic break-even copper price (the price at which NPV at 8% equals zero) is $1.55/lb. Total operating costs (including mining royalty, transportation, shipping, treatment and refining costs) are anticipated to average $1.31/lb of payable copper. After silver, gold and molybdenum credits of $0.40/lb of payable copper, the cash cost is estimated at $0.92/lb. Capital development costs are estimated to be $846 M, while deferred and sustaining capital, including closure costs is estimated to be $148 M. 1.16.3.1

Sensitivity analysis (pre-tax)

Figure 1.10 illustrates the sensitivity of the project economics to copper price, silver price, gold price, molybdenum price, mining costs, processing costs and capital development costs on a pre-tax basis. The Project is most sensitive to copper price, which accounts for approximately 84% of payable metal value. A 10% increase in copper price from the base case of $2.00/lb results in an increase in pre-tax NPV at 8% of $235 M, and conversely, a 10% decrease in copper price results in a decrease in NPV by the same amount. Figure 1.10 Pre-tax Cashflow Sensitivity Analysis (NPV – 8% discount rate)

1200

Copper Spot Price

1000

Silver Spot Price

NPV ($M)

800

Gold Spot Price

600

Moly Spot Price

400

Mining Cost

200

Processing Costs

0 ‐25%

‐20%

‐15%

‐10%

‐5%

0%

5%

10%

15%

20%

25%

Development Costs

‐200

1.16.4 1.16.4.1

Post-tax analysis Key cashflow assumptions (post-tax)

Norsemont procured the opinion of Peruvian tax specialist Picon & Asociados, to assist in determining the impact of tax on the project post-tax cashflows and financial evaluations. The key tax, depreciation and amortisation assumptions are summarised in Table 1.23.

Page 74


Table 1.23 Tax and Depreciation Rates Tax and Depreciation Assumptions Units

Value

Corporate Tax Rate

%

30.00%

Profit Distribution to Employees (pre-tax profits)

%

8.00%

Tax Depreciation Rates

Owned

Fixed Plant Buildings

%pa %pa

20% 5%

Leased

Movable Plant Buildings

%pa %pa

50% 20%

Pre-operating Costs 100% deductible in one year

1.16.4.2

Summary of results (post-tax)

Based on the assumptions set out above, the project post-tax NPV at 8% is $303.6 M and post-tax IRR is 15.5%. The project is expected to pay back the initial capital after four years of production. 1.16.4.3

Sensitivity analysis (post-tax)

Table 1.24 illustrates the sensitivity of the project economics to copper price, silver price, gold price, molybdenum price, mining costs, processing costs and capital development costs on a post-tax basis.

Page 75

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 1.24 Post-tax Project Sensitivity Analysis Post-tax NPV@8% Post-tax IRR $M

% Change

Copper

Silver

Gold

Molybdenum

Mining Costs

Processing Costs

Development Costs

Base = $ -20% -10% 0% 10% 20%

2.00 /lb

Base = $ -20% -10% 0% 10% 20%

12.00 /oz

Base = $ -20% -10% 0% 10% 20%

800.00 /oz

Base = $ -20% -10% 0% 10% 20%

13.00 /lb

Undisc Payback years

2 150 304 456 608

8.0% 11.9% 15.5% 18.7% 21.6%

5.00 4.00 4.00 3.00 3.00

286 295 304 313 322

15.1% 15.3% 15.5% 15.7% 15.9%

4.00 4.00 4.00 4.00 4.00

297 300 304 307 311

15.3% 15.4% 15.5% 15.5% 15.6%

4.00 4.00 4.00 4.00 4.00

272 288 304 320 336

14.8% 15.1% 15.5% 15.8% 16.1%

4.00 4.00 4.00 4.00 4.00

Base = $ variable -20% -10% 0% 10% 20%

351 327 304 280 257

16.4% 15.9% 15.5% 15.0% 14.5%

4.00 4.00 4.00 4.00 4.00

Base = $ variable -20% -10% 0% 10% 20%

372 338 304 269 235

16.8% 16.2% 15.5% 14.7% 14.0%

3.00 4.00 4.00 4.00 4.00

Base = $ -20% -10% 0% 10% 20%

417 360 304 247 189

20.2% 17.6% 15.5% 13.6% 12.0%

3.00 3.00 4.00 4.00 4.00

845.99 M

Page 76


As with the pre-tax analysis, the post-tax NPV and IRR are most sensitive to copper price (Figure 1.11). A 10% change in copper price results in a change in post-tax NPV at 8% of ±$153 M. Figure 1.11 Post-tax NPV Sensitivity (8% discount rate)

800

Copper Spot Price

700 Silver Spot Price

600

NPV ($M)

500

Gold Spot Price

400 Moly Spot Price

300 200

Mining Cost

100 Processing Costs

0 ‐100

‐25%

‐20%

‐15%

‐10%

‐5%

0%

‐200

1.17

CONCLUSIONS

1.17.1

Project overview

5%

10%

15%

20%

25%

Development Costs

A DFS has been completed for the Constancia Project, covering all disciplines, i.e. resource modelling, mining, metallurgical testwork, process design, plant and infrastructure design, project implementation, environmental studies, and capital and operating cost estimation to ±15%. A post-tax cashflow model has been prepared by Norsemont, which indicates a Base Case NPV of $300 M and an IRR of 15.3%. However, these values do not take into account the financing costs required to develop the Project, which will now be investigated by Norsemont. Project economics are most sensitive to the long-term copper price: a constant price of $2.00/lb has been assumed for the Base Case cashflow model. Other important economic variables are the total capital cost, treatment charges, the cost of diesel and electricity. 1.17.2

Risks and opportunities

Risk analysis identified the following key issues: 

Power supply - there is potential that all available capacity at the current supply point of Tintaya substation is secured by other parties. Should this occur, Norsemont would have to fund the additional capital cost of obtaining a supply from the more distant Cotaruse Substation.

Page 77




Power Price - the price of electricity paid by Norsemont during operations will be subject to market conditions, hence competition for power may result in the cost of power being higher that the current assumption. Negotiations with the various organisations that generate and transmit power may allow more confidence on power pricing and availability. However, it is understood that firm commitments by government organisations can only be provided closer to the time of supply.



Escalation - the estimate is based on pricing lower than that experienced in late 2008 at the peak of the market. Should commodity prices increase and market competition intensify then it is likely that the project will be exposed to positive escalation.



Political - the upcoming elections will coincide with the execution of the project, this has potential to affect both the project delivery and the project outcome through social conflict and civil unrest.



Pit geotechnical - geotechnical parameters are key criteria in the mine design, in particular the assessment for overall pit slope stability. Additional operational costs would be incurred if the pit slope stability estimates prove to be too optimistic.



Norsemont as Developer/Operator - the implementation strategy is based on Norsemont undertaking the overall project management role for the execution phase of the project, for constructing the bulk earthworks using its construction fleet, and the owner-operator role for the operations phase of the project. Significant project and construction management capabilities will be required by the Owner for this and particularly to ensure the timely completion of the bulk earthworks. This approach risks causing delay to the project schedule and prolongation claims from other contractors with liability falling on Norsemont. Cost increases arising from the need to employ additional staff to provide project management are a further risk.



To avoid delays to the construction schedule the purchase of community lands and the relocation action plan must be implemented at the beginning of 2010. If these activities are delayed then there is a high likelihood that the construction schedule will not be met.

Opportunities exist to improve the Project in the following areas: 

Detailed planning for the bulk earthworks and purchase of the Owners civil construction fleet would provide greater certainty on the capital and operational costs of the fleet.



Investigation of the availability of second hand accommodation camps has the potential to save cost.



The investigation of the availability of new long lead equipment, i.e. cancelled orders etc., may provide schedule opportunities for the project. Currently these long lead items are critical path items, hence a reduction in time of the delivery of key equipment will provide schedule assurance and possibly an overall schedule reduction.



Early execution of Front End Engineering Design of the process plant would provide greater degree of scope definition and more definitive costing.



Detailed design of the improvement works for the access road upgrade will increase the level of confidence around the scope of works. In addition, accelerated completion of the access road upgrade will reduce transportation and travel risks and reduce travel time and delays.

Page 78


1.18

RECOMMENDATIONS

1.18.1

Mineral resources

No further work required until a decision has been made to proceed to development. 1.18.2

Mining and mineral reserves

The DFS pit designs include a quadrant where the probability of bench scale failure is 44%. The incremental costs associated with flattening the pit slope in this area should be assessed relative to the corresponding reduced probability of bench scale failures. The fuel (diesel) and explosive unit prices used in the study should be confirmed by following up on the budget pricing submissions already requested from local suppliers. The mining and civil construction fleets should be evaluated to determine any synergies that might improve project value. Electric shovels and diesel drills have been assumed as the basis for the study. Further work should be performed to confirm the best approach considering capital and operating costs, mine operability and productivity to confirm the most suitable mine equipment solution. 1.18.3

Geotechnical and hydrogeological studies

Selected geotechnical investigations are recommended to provide support for detailed engineering. These include: 

Additional field geotechnical investigations within the southern portion of the PAG WRF (limited investigations have been possible to date, due to access restrictions).



Investigations to evaluate the need for removing the extremely weathered diorite to significant depths beneath the TMF embankment.



Additional geotechnical characterisation of the tailings and waste rock materials.



Additional investigations in potential borrow areas to further characterise potential construction materials and refine the quantity estimates.



Within the mine pit, two additional drillholes in Sector VI and one in Sector VII are necessary to investigate the potential that locally observed geological faults may extend through this area.



Risk analyses should be carried out to investigate the probability and consequences of a rock mass slide occurring over the access ramp, thereby removing access to the pit in the Sector VI area.

With regards to hydrogeological studies, further investigations should be completed within the TMF and PAG WRF to confirm the criteria used in the design of these facilities. Within the PAG WRF, the objective will be to confirm the containment capability of the existing regime. An investigation is underway to assess the potential influence of pit dewatering on the lakes and karstic area situated north of the pit. The investigation would include drilling, hydraulic testing and piezometer

Page 79


installations along the Yanak fault and within the lakes catchment area to assess the existing ground water system and the degree of hydraulic continuity and connection between the Yanak fault and lake basins. Due to the risks associated with the availability of groundwater during drought conditions and uncertain water quality, a groundwater exploration program is recommended. Targets include: the Yanak fault south of the pit area and structures paralleling the Chilloroya valley below the glacial valley infill material. The installation of additional groundwater monitoring and sampling wells is recommended for the following locations: (1) within Quebrada Telaracaca to monitor potential seepages from the PAG WRF via fault zones; (2) upgradient and downgradient of the PAG WRF within Quebrada Huayllachane; and (3) downgradient of the proposed plant site. 1.18.4

Process testwork and plant design

Additional testwork is recommended to investigate: 

The potential to recover zinc concentrate as a saleable by-product



Silver recovery to copper concentrates



The potential to reduce the talc/amphibole content of the molybdenum concentrate

1.18.5

Environmental and permitting

Norsemont intends to submit the ESIA in December 2009. ESIA approval is expected in September 2010. Approval for construction is expected October 2010.

Page 80


2.

INTRODUCTION

2.1

BACKGROUND

Norsemont is developing the Constancia Cu-Mo-Ag Project in Southern Peru, 100 km south of Cusco. A 43-101 compliant Technical Report was previously prepared by GRD Minproc in May 2008 to support an updated mineral resource and provide preliminary metallurgical testwork information. Following this, GRD Minproc was appointed by Norsemont to undertake and manage a Definitive Feasibility Study (DFS) which has included a further resource update, comprehensive metallurgical testwork, mine design, plant and infrastructure design and development of capital and operating costs. The DFS forms the basis of this Technical Report. The Constancia deposit is a large-scale porphyry copper orebody located 4100 m above sea level (masl) in the Andes mountains. The proposal is to develop a Project comprising open pit mining and flotation of sulphide minerals, to produce commercial grade concentrates of copper and molybdenum. Silver and a small quantity of gold at payable levels would report to the copper concentrate. The Project is largely self-contained, with mine, mill, maintenance facilities, administration and fully serviced accommodation camp located on the mine site. Supporting infrastructure includes a power line to bring power from an upgraded supply point on the national grid at Tintaya, 70 km away. The public road to site will be upgraded to meet demands of extra traffic, particularly concentrate trucks and freight services. Raw water will be extracted from bores surrounding the open pit, and a tailings dam will be constructed within 5 km of the mine, on land owned freehold by Norsemont. The trigger for preparation of the Technical Report was a press release disclosure of the Mineral Resource estimate made by Norsemont on 28 September 2009. This report has been prepared in accordance with form 43-101F (the “Technical Report”) of the Canadian Securities Administrators National Instrument 43-101 (NI 43-101). 2.2

SCOPE OF WORK

GRD Minproc’s scopes of work for the DFS and current NI 43-101 Technical Report were: 

Review and verify Norsemont’s digital models of geology and copper domains



Develop a new resource model to include all new exploration drilling obtained subsequent to the May 2008 NI 43-101 Technical Report



Report resources in accordance with the requirements of NI43-101



Undertake a comprehensive metallurgical test programme on representative core samples of the deposits



Prepare a comprehensive plant design including on-site and off-site infrastructure, and prepare associated capital cost estimates



Prepare an estimate of mining costs and plant operating cash costs, including post-production (i.e. off-site) costs and provide a pre-tax economic evaluation of the Project.

Norsemont engaged the following consultants in addition to GRD Minproc:

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

Knight Piésold responsible for site geotechnical investigations, and design and costing of the waste rock and tailings storage facilities, and water management systems.



SIGT S.A., a Columbian consulting company which provides services in design and supervision of roads and highways, supervised the design and costing for the access road to the project.



Cesel Ingenieros S.A. undertook design and costing for the Constancia project HT power line and sub-stations.



Jorge Picon, financial consultant to Norsemont, was responsible for the preparation of the post-tax financial model.



Ground Water International undertook hydrogeological studies

2.3

SOURCES OF INFORMATION

Norsemont provided all the necessary drilling, sampling, analytical and geological data to GRD Minproc for mineral resource modelling and estimation purposes. The mineral resource estimate reported here relies on a comprehensive drillhole database that included historical data for the site. The drilling programme is discussed in detail elsewhere in this report. Drill core was selected, in accordance with a sampling plan, from the client’s core shed in Cusco. Core samples were sent to Lakefield SGS Laboratories in Chile for metallurgical testing. The results of this testwork have been relied upon in development of the process flowsheet and plant design. Cost estimate data was obtained from reputable equipment suppliers and Peru-based contractors relying on technical specifications and material quantity take-offs provided by GRD Minproc. Operating cost information and selected input parameters for the economic evaluation were obtained from market pricing and from Norsemont. 2.4

SITE INSPECTIONS

GRD Minproc staff visited the site on several occasions. Lynn Widenbar, consultant geologist for GRD Minproc, has inspected the site, viewed exploration drill core and analysed the geological database at the client’s office in Lima. Ross Oliver (Manager Mining), Craig Cutriss (Principal Process Engineer), Nigel Ricketts (Process Manager), Peter Shaw (Lead Mechanical Engineer) and Rick McCracken (Lead Mechanical Designer) visited site in 2008. Mr David Miranda and Mr. Luis Caldera of GRD Minproc Chile, supervised the laboratory test programme and made frequent visits to the laboratory throughout the test programme. In addition, Mr Miranda inspected the drill core in Cusco and the core samples that were transported to Santiago. Dr Greg Harbort of GRD Minproc Brisbane visited the laboratory in Santiago and supervised a pilot scale test program using a 25 tonne bulk composite sample of Constancia drill core.

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Robert Cummings, Principal Geological Engineer for Saguaro Geoservices, subconsultant to Knight Piésold and Co. (USA), visited the site March 12-14, 2008 and September 26-30, 2008. Thomas F. Kerr, President, Knight Piésold and Co. (USA), Senior Geotechnical Engineer, visited the property on several occasions in 2008 and 2009. Gilberto Dominguez, Knight Piésold and Co. (USA), Executive Project Manager, visited the site in February 2009. Olimpio Angeles, Senior Geologist, Knight Piésold Consultores S.A. visited the site on several occasions between 2006 and 2009. Rubén Vargas, Knight Piésold Consultores S.A., Geotechnical Engineer visited the site on several occasions between 2008 and 2009. Lorena Viale (Knight Piésold) Project Manager for the Environmental Impact Assessment visited the site on numerous occasions and has participated in workshops and presentations to communities living within the area of influence. Carol Harrison (Project Manager) and Lucia Avila (Project Coordinator) of Social Capital Group, are responsible for the Social Impact Assessment. They, too, have visited the site on numerous occasions and have participated in workshops and presentations to communities living within the area of influence. 2.5

CONTRIBUTORS TO THE REPORT

The major contributors to this report are listed in the Table 2.1.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 2.1 Contributors to Constancia DFS Technical Report Discipline Overall Compilation of

Section

Responsible Party

Qualified Person

Other Expert

All

GRD Minproc

Dan Greig

Sections 5 to 13, 15

Norsemont

Gaston Loyola

Section 4

Norsemont

Gaston Loyola

Technical Report Background, geology, history of exploration and sampling Mineral concessions and permits Data quality

Section 14

GRD Minproc

Lynn Widenbar

Resources

Section 17

GRD Minproc

Lynn Widenbar

Mineral reserve and Mining

Sections 17.14 and

GRD Minproc

Ross Oliver Robert Cummings

Olimpio Angeles

Tom Kerr

Gilberto

18.1 Geological engineering (pit

Sections 18.1.3 and

Saguaro

slopes)

18.2.2

Geoservices

Site geotechnical

Section 18.2,

Knight Piésold

investigations, waste

excluding the above

Dominguez,

characterisation, waste rock

pit geotechnical

Ruben Vargas

and water management

sections; plus Sections 18.6 and 18.7

Hydrogeological

Section 18.3

MWH Peru

David Evans and Jean Cho

Metallurgy

Section 16

GRD Minproc

Greg Harbort

Plant engineering

Section 18.4

GRD Minproc

Craig Cuttriss

Capital cost estimate (plant

Section 18.13

GRD Minproc

Craig Cuttriss

Chris Gilmour

GRD Minproc/

Greg Harbort

Sean Spraggett

Tom Kerr

Gilberto

and plant infrastructure) Operating cost estimate

Section 18.14

Norsemont Tailings management facility

Section 18.7

Knight Piésold

and waste dumps

Dominguez and Ruben Vargas

Power supply

Section 18.5.4

Cesel

Pablo Lozano and Mario Rojas

Access road

Section 18.5.1

SIGT

Luis Yafac

Environmental and Social

Sections 18.10 and

Knight Piésold

Lorena Viale

18.12

Norsemont

Carol Fries

Social Capital Group

Carol Harrison

Section 18.8

Norsemont

Sean Spraggett

Financial cashflow model

Section 18.16.1 to

GRD Minproc

Ross Oliver

(pre-tax)

18.16.3

Financial cashflow model

Section 18.16.4

Picon

Jorge Picon

Section 18.16

Norsemont

Sean Spraggett

Road transport and port operations

(post-tax) Financial analysis

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2.6

DISCLOSURE OF INTEREST

GRD Minproc is not an associate or affiliate of Norsemont, or of any associated company. GRD Minproc’s fee for this Technical Report is not dependent in whole or part on any prior or future engagement or understanding resulting from the conclusions of this report. The fee is in accordance with standard industry fees for work of this nature.

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3.

RELIANCE ON OTHER EXPERTS

In preparing this report, GRD Minproc has relied on input from Norsemont and a number of wellqualified independent consulting groups as recorded in Section 2.5, particularly regarding the hydrological/hydrogeological and geotechnical investigations and recommendations, off-site infrastructure engineering, environmental and legal matters. Further, GRD Minproc has relied on results of metallurgical testwork undertaken by a well-credentialled laboratory as a basis for its process flowsheet design. The authors of this Technical Report have relied on input from a number of Experts who would not be considered Qualified Persons under NI43-101, but who have the necessary qualifications and experience to provide input and opinions incorporated into the Report. These include information regarding: 

Status of Mining Concessions



Land ownership and permitting requirements to obtain surface rights, construction and exploitation permits and licences for plant and infrastructure



Environmental studies



Access road upgrade and costs



Post tax financial analysis

Specific information provided by Norsemont and its consultants includes: 

Metal prices



Treatment charges



Transport costs (road, port and shipping)



Access road upgrade costs, through Norsemont’s sub-contractor SIGT



Power cost, including power transmission, through Norsemont’s sub-contractor CESEL



Diesel cost, delivered, including storage and dispensing



Exchange rates



Information concerning operations such as organisational structure, labour conditions (sourcing, salaries and on-costs, transport), camp services and maintenance.

A simplified project cashflow model was prepared by GRD Minproc on a pre-tax, 100% equity basis. Norsemont has provided information regarding taxation, royalties and employee profit sharing, and has undertaken post-tax financial analysis, which was completed by Picon Associates a local firm specializing in Peruvian taxation laws. GRD Minproc does not claim to be expert in financial matters, and has relied on Norsemont with respect to the above.

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4.

PROPERTY DESCRIPTION AND LOCATION

4.1

GENERAL LOCATION

The Constancia Project is located in the southeastern Andes of Peru, in the Chamaca and Livitaca Districts, Province of Chumbivilcas, Department of Cusco on map sheet 29-S Livitaca (Figure 4.1). The property is approximately 600 km southeast of Lima at elevations of 4000 to 4500 masl. Road access to the property is from either Arequipa (7 hours by road) or Cusco (6.5 hours by road). Geographic coordinates at the centre of the property are longitude 71°47’ west and latitude 14°27’ south. Figure 4.1 Project Location

4.2

PERUVIAN MINING LAW

The General Mining Law of Peru defines and regulates different categories of mining activities, from sampling and prospecting to commercialisation, exploitation, and processing. Mining concessions are granted using UTM coordinates to define areas generally ranging from 100 ha to 1000 ha in size. Mining titles are irrevocable and perpetual, as long as the titleholder maintains payment of the “Derecho Vigencia” fees up to date (Ministerio de Energia y Minas, 1998). No royalties or other production-based monetary obligations are imposed on holders of mining concessions; instead, a holder must pay an annual maintenance fee of $3/ha (for metallic mineral concessions) for each concession actually

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acquired, or for a pending application (petitorio), by June 30th of each year. The concession holder must sustain a minimum level of annual commercial production of greater than $100/ha in gross sales within six years of the grant of the concession (‘’Minimum Required Production’’). If the concession has not been put into production within that period, then the concession holder must make an additional payment in the form of a penalty of $6/ha for the seventh through eleventh year of the grant of the concession, rising to $20/ha thereafter. The concession holder shall be exempted from the penalty if the investment made during the previous year was 10 times the penalty. The holder of a mining concession is entitled to all the protection available to all holders of private property rights under the Peruvian Constitution, the Civil Code, and other applicable laws. A Peruvian mining concession is a property-related right; distinct and independent from the ownership of land on which it is located, even when both belong to the same person. The rights granted by a mining concession are defensible against third parties, are transferable and chargeable, and, in general, may be the subject of any transaction or contract. To be enforceable, any and all transactions and contracts pertaining to a mining concession must be entered into a public deed and registered with the Public Mining Registry. Conversely, the holder of a mining concession must develop and operate his/her concession in a progressive manner, in compliance with applicable safety and environmental regulations and with all necessary steps to avoid third-party damages. The concession holder must permit access to those mining authorities responsible for assessing that the concession holder is meeting all obligations. 4.3

CONSTANCIA MINING CONCESSIONS

Figure 4.2 shows the Constancia and San José mineralised zones on the property in relation to the concession boundaries. Most of the known mineralisation is located in the claims Katanga J, Katanga Q, Katanga K, and Peta 7, although small mineralised outcrops are common throughout the area. Norsemont is the current registered titleholder of the Constancia concessions which comprise 22516.04 ha in thirty-six mineral concessions that are in good standing. All of the concessions have been titled and have also been recorded with the Public Registry.

Page 88

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 4.2 Concession Boundaries

Page 89


Table 4.1 shows the names and sizes of the pertinent concessions, concession dates and the company that holds the title. All the properties are located on the Livitaca (29-S) map sheet. Table 4.1 Constancia Concessions Concession Name

Code

Concession

Concession

Granted

Holder

Hectares

Peta 5

05006089X01

28-11-1989

Livitaca

934.52

Katanga J

05004406X01

29-03-1990

Mitsui

400.00

Katanga Q

05005529X01

09-05-1990

Katanga

150.01

Katanga K

05004407X01

16-07-1990

Mitsui

300.00

Peta 6

05006090X01

29-10-1996

Katanga

Santiago 4

010083495

23-12-1996

Mitsui

34.16

Santiago 3

010083695

25-03-1997

Mitsui

700.58

Santiago 5

010083295

30-04-1997

Mitsui

602.12

Katanga V

010248497

31-10-1997

Mitsui

100.00

Katanga T

010248397

15-11-1997

Mitsui

100.00

Santiago Apostol I

010229294

31-03-1998

Mitsui

424.49

Peta 17

0506198AX01

13-12-1999

Katanga

49.05

Peta 7

05006198X01

13-12-1999

Katanga

351.70

1000.00

Since 14 December 2006 two new claims were added to the pre-existing Katanga “Unidad Economica Administrativa” (UEA), bringing a total of 13 claims (6785.2831 Ha.) to the UEA. The UEA Katanga Este (Code N° 01-00029-83-U) was created by Rio Tinto and transferred to Norsemont Peru S.A.C according to the Resolucion Jefatural N° 5404-2006-INACC/J. Project concessions are currently in good standing and concession fees have been paid for the calendar year 2009. Norsemont confirms that it has renewed the Semi Detailed EIA in January 2009 for a period of 14 months. On November 23, 2007, 11 Mining applications were submitted to Ingemmet to cover new areas with a total of 7700 ha. In addition, the Company also entered into an auction at the Ministry of Energy and Mines for an additional 12 claims in 2007 and received notification in 2008 that the claims were granted to the Company. The mining applications were granted to Norsemont as shown in Table 4.2.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 4.2

Concession Name

Code

Constancia Concession Grants Concession Granted Resolution

Area (Ha)

Constancia 5

01-00253-07

01/03/2007

D.M. Titulado D.L. 708

1000

Constancia 6

01-00254-07

01/03/2007


1000

Constancia 7

01-00255-07

01/03/2007


1000

Constancia 8

01-00256-07

01/03/2007


900

Constancia 9

01-00257-07

01/03/2007


1000

Constancia 10

01-00258-07

01/03/2007


100

Constancia 11

01-00259-07

01/03/2007


1000

Constancia 12

01-00260-07

01/03/2007


1000

Constancia 13

01-00261-07

01/03/2007


1000

Constancia 14

01-00262-07

01/03/2007


1000

Constancia 15

01-00263-07

01/03/2007


1000

Constancia 16

01-00264-07

01/03/2007


300

Constancia 17

01-06147-07

05/05/2008

001061-2008-INGEMMET/PCD/PM

700

Constancia 18

01-06148-07

21/04/2008


400

Constancia 19

01-06149-07

12/05/2008


700

Constancia 20

01-06150-07

12/05/2008

001234-2008-INGEMMET/PDC/PM

600

Constancia 21

01-06151-07

12/05/2008


700

Constancia 22

01-06152-07

12/05/2008


600

Constancia 23

01-06153-07

18/04/2008


1000

Constancia 24

01-06154-07

17/04/2008


800

Constancia 25

01-06155-07

17/04/2008


800

Constancia 26

01-06156-07

18/04/2008


600

4.4

MINERAL RIGHTS OWNERSHIP

4.4.1

Rio Tinto purchase

On 29 November, 2007, Norsemont exercised its exclusive option to acquire an initial 51% of the Constancia copper project from Rio Tinto Mining and Exploration Ltd (Rio Tinto). Combined with the 30% interest that Norsemont previously acquired from Mitsui Mining and Smelting, this acquisition brought Norsemont’s interest in the Constancia Project to 81% at that time. In order to exercise the initial option, granted in 2005, Norsemont was obligated to completed work expenditures at Constancia of $7.8 M, make $5 M of payments and issue 1 250 000 common shares (or their cash equivalent) to Rio Tinto. All three obligations have been met. Table 4.3 presents the cash payment, share issuance and work expenditure schedule. Under the terms of the same Rio Tinto option agreement, Rio Tinto had 60 days within which to decide whether to claw back a 17% undivided interest in Constancia. This could only be done if, in the

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reasonable opinion of Rio Tinto, Constancia’s global resource estimate is not less than 4 Mt (8.8 billion pounds) of contained copper. In the event that Rio Tinto were to claw back a 17% interest, they would have had to pay Norsemont 300% of the Company’s net cash payments, work expenditures and share issuances related to the project. Rio Tinto advised Norsemont on 18 January 2008 of its intention not to exercise its claw-back, allowing Norsemont to acquire the remaining 19% interest for $8 M. This payment was made on 26 March 2008. Table 4.3 Option Exercise Schedule Option exercise schedule

Cash ($)

Exploration expenditures ($)

Shares

On signing of LOI

10 000 (1)

-

-

45 days following LOI

90 000 (1)

-

-

150 000 (1)

-

-

April 20, 2005 June 30, 2005

-

250 000 (2)

October 2005

200 000 (1)

500 000

-

April 2006

250 000 (1)

500 000

-

June 30, 2006

-

-

250 000 (2)

October 2006

300 000 (1)

500 000

-

April 20, 2007

400 000 (1)

-

-

June 30, 2007

-

-

250 000 (3)

October 2007

500 000 (1)

1 300 000

-

April 2008

500 000 (1)

1 000 000

-

June 30, 2008

-

-

250 000 (2)

October 2008

750 000 (1)

-

1 500 000

-

October 2009

1 850 000 (1)

2 500 000

250 000 (2)

Total

5 000 000

7 800 000 (4)

1 250 000

(1) Paid (2) Issued (3) Paid $365,169 in lieu of shares (4) Incurred. As of July 2009, the Company has expended in excess of $73 000 297 of qualifying expenditures for exploration.

Upon commencement of commercial production, the Company is required to make the following additional payments to Rio Tinto: 

$250 000 should the Company have between 34% and 50.9% interest in the Constancia Property, or



$500 000 should the Company have a majority interest in the Constancia Property of 51% or greater

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4.4.2

Mitsui Mining and smelting purchase

On 1st November 2007, Norsemont announced that it had signed an agreement to purchase 30% of the Constancia Project from Mitsui Mining and Smelting (Mitsui). The Mitsui agreement called for a $9.8 M payment over 20 months on the following schedule: Initial payment on signature of the public deed of the agreements:

$100 000

First payment on or before January 31, 2008:

$700 000

Second payment on or before June 30, 2008:

$2 000 000

Third payment on or before December 31, 2008:

$3 000 000

Final payment on or before June 30, 2009:

$4 000 000

Total payments of $9.8 M have now been made, bringing Norsemont’s interest in the Constancia property to 100%. 4.5

SURFACE RIGHTS

Most of the project area is located within private property owned by Norsemont, though adjacent parts of the deposit and infrastructure are within third party land. For rights purposes, Norsemont previously purchased the Fortunia property that covers most of the main resource area (see Figure 4.3). Other areas of interest are being investigated to assess their value to the project. Figure 4.3 displays the private land holdings by Norsemont, which are also summarised in Table 4.4. Table 4.4 Private Lands Summary Name

Area

Land Registered

Titled

Fortunia

974.71

yes

yes

Cristina Velazco

423.38

Yes

yes

Morocota

885.35

yes

yes

Arizona

371.81

Pending

pending

Antonio Velazco 1

943.00

yes

yes

Antonio Velazco Lot B

499.64

Pending

pending

Total

4,097.89

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 4.3 Surface Rights

Additional purchase agreements for private surface land have been signed on the following properties: 

Arizona / Morocota On 31 January 2009, Norsemont signed a purchase agreement for the acquisition of the surface rights which covers the mining concessions Santiago 3, Peta 17, Katanga K and Katanga J. The private lands, totalling 1257.16 ha, are known as Moroccota (885.35 ha) and Arizona (371.81 ha). The previous owners, Enrique Antonio Velasco Torrelio and Giovanna Abuhadba Sayhua, agreed to the sale of the properties for the total amount of $620 000 as a final payment for the transaction. The company provided a certified cheque for the amount of $250 000 at the contract signature, and paid $220 000 at the contract registration in Sicuani, Cusco Public Register on 28 April, 2008, leaving a balance of $150 000 to be paid when Moroccota is registered in the name of Norsemont.



Cristina Velasco On 19 March 2009, Miriam Rozas on behalf of Cristina Velasco; signed the purchase agreement for the Cristina Velasco property (326.09 ha) for the total amount of $200 000. The company advanced a payment of $60 000 at the contract signature, leaving a pending balance of $140 000 to be paid once the property is registered in the name of Norsemont.

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On 21 January 2009, Norsemont signed a modification to the purchase agreement, increasing the value by $50 000 to $250 000. This modification was approved after certification that the land covered an additional 97.29 hectares, i.e. a total of 423.38 ha. The final payment of $190 000 was processed on 31 March 2009, and the property is now registered in the name of Norsemont in the public registers of Sicuani – Cusco. 

San Antonio On 23 March, 2009 Norsemont signed a purchase agreement for the San Antonio property (1442.64 ha) for a total amount of $2 175 000. The property is divided into two portions consisting of a 943 ha parcel that is titled and registered and a 499.64 ha parcel that is registered and in the process of being titled. Norsemont made full payment of 1 421 716 for the 943 ha land parcel and holds 100% title. The company advanced a payment of $112 992.60 at the contract signing for the 499.64 ha land parcel, having a pending balance of $640 291.40 which will be paid when the property is registered in the name of Norsemont.

Other areas of interest are being investigated to assess their value to the project. 4.6

ENVIRONMENTAL REGULATIONS

The General Mining Law of Peru is the primary body of law with regard to mining activities in Peru. The regulation surrounding environmental obligations of mining activities is governed by the Environmental Protection Law (D.S. 016-93 EM). The Law is administered by the Ministry of Energy and Mines (MEM). Depending on the phase of the project the MEM can require a mining company to prepare an Environmental Evaluation (EA), an Environmental Impact Assessment (EIA), a Program for Environmental Management and Adjustment (PAMA), and a Closure Plan. Mining companies are also subject to annual environmental audits. According to Peruvian regulations (S.D. 038-98-EM) the environmental requirements for mining exploration programs are divided into classifications Category 1, 2 and 3. Category 1 is for general exploration activities and requires no authorisation or fees. Classification 2 includes drilling of less than 20 drillholes within a 10 ha area. An application in the form of a sworn declaration must be submitted. Category 3 pertains to mining exploration programs with more than 20 boreholes, exploration areas greater than 10 ha, or construction of more than 50 m of tunnels. Submission and acceptance of an EA (Evaluación Ambiental) is required for approval of Category 3 activities. The MEM has a period of 25 days to review and approve or disapprove the EA; the EA is considered approved if the MEM does not respond within that period. A mining company that has completed its exploration stage must submit an ESIA (Environmental Social Impact Assessment) when applying for a new mining operation or filing for a processing concession, to increase the size of its existing processing operations by more than 50%, or to execute any other mining project. An ESIA is subject to public workshops and audiences with the local communities prior to its approval. A mining company planning to enter into production must also prepare and submit a Closure Plan (Plan de Cierre) prior to the start of its construction phase. The Closure Plan must outline what measures will

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be taken to protect the environment over the short, medium and long term following closure of its operations. The General Mining Law of Peru has in place a system of sanctions or financial penalties that can be levied against a mining company not in compliance with the environmental regulations. The Constancia Project has been designed to consider all relevant legislation applicable to the development of mining projects in Peru, including mines, roads, ports and transmission lines. Additional legislation that has been considered includes legislation and regulations regarding archaeological areas of significance, endangered and protected species, and community relations and public disclosure programs. The ESIA was developed in accordance with the following legislation and standards: 

Peruvian Political Constitution, 1993



Environmental General Law, 2005



Private Investment Growth Law, 1991



Environmental Impact for Works and Activities Evaluation Law, 1997



Sustainable Utilisation of Natural Resources Organic Law, 1997



Environmental Impact Assessment National System Law, 2001



Water Resources Law, 2009



Water Quality National Standards, 2008



Air Quality National Standards, 2008



Noise Quality National Standards, 2003



Health Law, 1997



Environmental Management National System Law, 2004



Cultural Heritage Protection Law, 1985



Solid Waste Law (2004) and modifying (2008)



Forestry and Wildlife Protection Law, 2001



Mines Closure Law, 2003



Public Participation Regulations, 2008



Regulation for Environmental Protection in Mining - Mineral Processing Activities, 1993



Maximum Permissible Levels for Liquid Effluents from Mining – Mineral Processing Activities, 1996



World Bank Environmental and Communities Standards.

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5.

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1

ACCESSIBILITY

The project is accessible from Lima by air, via either Arequipa or Cusco, and then by road over paved and gravel highways. The routes and approximate distances and driving times are shown in Table 5.1. Table 5.1 Travel Distance and Time to Constancia From

Leg

Arequipa

Canahuasi

80

1.25

Imata

63

0.75

103

2.50

68

2.00

Yauri/Espinar Uchucarcco turnoff Project site

Cusco

Time (hours)

20

0.50

Total

334

7.00

Sicuani (paved road)

140

2.00

Descanso

40

1.00

Yauri/Espinar

45

0.75

Uchuccarco turnoff

68

2.00

Project site

20

0.50

313

6.25

Total

5.2

Distance (km)

CLIMATE AND VEGETATION

Climate at the Constancia site can be classified as humid and seasonably cool with well-defined, rainy, and drier seasons. The majority of the precipitation typically occurs over the period of October to April, the summer months. Orographic (elevation) and physiographic effects also have influences on the climate. A design set of climatological data has been developed based on collected site data and data from stations located near the site. Due to the short length of climatological record at the on-site station (18 months), a synthetic precipitation record was developed for the site. A regression analyses was used by correlating the precipitation data from the on-site station with nearby climatological stations located in Santo Tomas, Yauri, Pisac, Livitaca and Imata, which have longer periods of record. Estimates for the design 24 hour storm events were calculated using a theoretical Extreme Type I (Gumbel) probability distribution on precipitation data from the Livitaca station, which is the closest station to the Constancia site and is located within the same drainage basin. Storm events observed at the Livitaca station were also the largest of the nearby stations.

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The following summarises the climatological and hydrologic data adopted for the design: 

Annual Precipitation values: Annual average precipitation – 1172 mm Annual maximum precipitation – 1887 mm Annual minimum precipitation – 485 mm Average wet season (October to April) precipitation – 1078 mm Average dry season (May to September) precipitation – 94 mm



24 hour Storm Event Estimates: 100 year/24 hour storm precipitation – 92 mm PMP/24 hour storm precipitation – 263 mm



Run-off and Infiltration: Average annual runoff – 172 mm Average annual infiltration – 468 mm



Air Temperatures: Daily average maximum – between 13 and 16 °C Daily average minimum – between 1 and 12 °C



Average Annual Evaporations: Potential annual evaporation – 1685 mm Annual evaporation from existing ground – 532 mm Evaporation from dry tailing – 475 mm Evaporation from wet tailing – 1518 mm

Vegetation is sparse and comprises limited brush on the flanks of the valleys. Exploration and mining operations are possible throughout the year. 5.3

LOCAL RESOURCES AND INFRASTRUCTURE

The main economic activity in the area is agriculture and cattle farming. A labour force with basic mining knowledge is present in Espinar. Food and basic supplies can be obtained in Espinar. Arequipa and Cusco are the nearest major centres, and both are over 300 km from the Project site by road. Several small permanently flowing streams, which are adequate for exploration requirements, are present in the area. More significant water sources include the Apurimac and Chilloroya rivers. Infrastructure includes several poorly maintained unsealed roads. All bridges on the access routes have low load capacity.

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Public electricity is available in Uchucarcco and, only recently, in the Chilloroya community. This power is supplied from the main grid via Chamaca. Grid electricity is available at Tintaya, 70 km from the project. Power supply is 138 kV, but there are plans to upgrade this to 220 kV by 2012. A public, but privately managed, landing strip is located at Espinar, about two hours from the project. The air strip is paved and maintained by Xstrata, and Norsemont has authorisation to use the strip. Small commercial and/or charter flight can use the air strip for daytime operations. 5.4

GEOMORPHOLOGY

The Katanga-Constancia district has moderate relief with shallow glacial valleys and low rolling (400-500 m) hills. Two erosion surfaces can be distinguished; an older one overlain by the Tacaza volcanics, and a younger surface that forms the present landscape. This erosion surface has incipient dendritic drainage indicating the beginning of a more intense erosion cycle. The geomorphology of the area suggests that the porphyry intrusives were exposed to an extensive period of weathering associated with an unusually deep oxidation profile compared to other porphyry deposits in the vicinity of similar age. There are abundant indications of glacial activity corresponding to the last glaciation period. This glacial erosion formed the present U-shaped valleys and resulted in the deposition of abundant moraines of up to tens of meters in thickness. Holocene glacial deposits are located on the margins of U-shaped valleys as lateral moraines, or in the centre of the drainage as terminal moraines. Recent alluvial deposits are not common, and are located only at the margins of the larger rivers.

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6.

HISTORY

Katanga was the main mine in the area and is located approximately 3 km northwest of Constancia, outside the Norsemont property boundaries. Copper and gold were exploited at Katanga from early last century to the early 1990s. The deposit consists of narrow skarn bodies developed in the contact between marbles and monzonite stocks, with Cu, Ag and Au mineralisation in hypogene sulphides. Mitsui and Minera Katanga operated the mine at different times between late 1970s to early 1990s. The San José prospect (now part of the Constancia prospect) was explored by Mitsui during the 1980s with a focus on locating more high-grade ore to add to the nearby Katanga mine operation. Exploration consisted of detailed mapping, soil sampling (1949 samples), rock chip sampling (1138 samples), ground magnetic and IP surveys with several drilling campaigns, mainly located in the western and southern sides of the prospect. Mitsui completed 24 drillholes (4190.5 m) and Minera Katanga completed 24 shallow, close-spaced drillholes at San José (1239.8 m). In 1995, reconnaissance prospecting identified evidence for porphyry-style mineralisation exposed over an area of 1.4 km x 0.7 km, open in several directions, with some Cu enrichment below a widespread leach cap developed in both porphyry and skarn. Negotiations with other parties to acquire an interest in the property were unsuccessful at this time. In May 2003, the area was revisited by Rio Tinto and the presence of a leached cap and potential for a significant copper porphyry deposit was confirmed. Negotiations with Mitsui, Minera Livitaca and Minera Katanga resulted in an agreement being signed on October 2003 with the underlying owners. The agreements included a Joint venture (JV) Option between Rio Tinto and Mitsui, and Purchase Option Agreements with Minera Livitaca and Minera Katanga to acquire 100% interests in their property. Rio Tinto commenced exploration in December 2003. The Rio Tinto exploration activities consisted of geological mapping, soil and rock chip sampling, and surface geophysics (magnetics and IP). Rio Tinto completed 24 diamond drillholes for a total of 7484.15 m. Based on this drilling, and using 50x50x10 m blocks with inverse distance squared (250x250x20 m) interpolation, Rio Tinto estimated a resource of 189 million tonnes of 0.68% copper and 200 ppm molybdenum, using a 0.5% copper cut-off. Although considered relevant, this resource can only be considered historical in nature as it does not comply with the guidelines of National Instrument 43-101. In late 2004, Rio Tinto sought partners for the Constancia prospect because the property did not meet their threshold which is nominally a project with potential to produce 200 000 tonnes of copper annually for a minimum of 20 years. Eventually, Norsemont entered into negotiations with Rio Tinto and these negotiations led to an agreement in early 2005. The first Norsemont geologists visited the property in June of 2005.

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7.

GEOLOGICAL SETTING

7.1

DISTRICT GEOLOGY

The oldest rocks in the area (Figure 7.1) correspond to a sequence of white, red, violet or grey mediumgrained sandstones with intercalations of reddish mudstones of the Lower Cretaceous Chilloroya Formation (also referred to as the Murco Formation). The Arcurquina Formation discordantly overlies the Chilloroya Formation and correlates with the Upper Cretaceous Ferrobamba Formation. These rocks are exposed in a north-south elongate area, 15 km long by 5 km wide, comprising a sequence of limestones, calc-arenites and lenses of conglomerates. These sedimentary formations have been intruded by plutonic rocks belonging to the AndahuaylasYauri Batholith of Oligocene age. The batholith varies from dioritic to granodioritic in composition, with plagioclase and orthoclase feldspar, quartz, hornblende, biotite, apatite, zircon and sphene being the main rock-forming minerals. Small mantos, veins and lenses of massive magnetite skarn are common in the area, and are probably related to the batholith emplacement. Several monzonitic stocks, dykes or laccoliths intrude and cross-cut all the lithologies mentioned above. Where these rocks have intruded limestones, it is common to find mineralised skarns, some of which contain Cu-Au-Ag mineralisation such as those at the Katanga mine. Some of the stocks have characteristics typical of porphyry copper deposits such as at Constancia.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 7.1 Simplified Geology of the Andahuaylas-Yauri Area

7.2

PROPERTY GEOLOGY

The Constancia porphyry copper prospect is located on the eastern margin of the Andahuaylas-Yauri Batholith, approximately 3 km southeast of the old Katanga mine (Figure 7.2). Surface geologic mapping by Norsemont between 2005 and 2008 has been greatly aided by the new drill roads and platforms for the 317 drill sites constructed by Norsemont to date. Interpretation of core drilling results has been very useful in improving the surface mapping.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 7.2 Geological Map of the Constancia Deposit

7.2.1

Stratigraphy

The oldest stratigraphic unit recognised on the prospect comprises clastic sediments corresponding to the Chilloroya Formation, consisting of a sequence of white, red, violet and grey medium-grained sandstones with intercalations of reddish mudstones. Immediately overlying this basal unit are massive, grey micritic limestones with minor intercalations of shales, which outcrop sporadically around the prospect and near the contacts with monzonite, sometimes occurring as roof pendants. This unit corresponds to the Arcurquina Formation (locally known as Ferrobamba Formation). When in contact with intrusive rocks, these alter to marble or pyroxene diopside-garnet-magnetite-epidote skarn, with or without sulphides. The limestones and skarns dip gently south-east, away from the principal monzonite in the southern part of the Constancia mineralised zone. The overall thickness of the sedimentary package is unknown. The determination of the base of this limestone unit may be of importance at Constancia, as this potential contact seems to correlate with the favourable skarn horizon at the Tintaya Mine and

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elsewhere in the region. This contact may be found at depth south and east of the presently drilled area, if not intruded by the monzonite. The underlying clastic sediments of the Murco Formation, and possibly the upper parts of the further underlying Hualhuani Formation (locally known as Soraya Formation), which lithologically consist of sandstones and siltstones, with occasional calcareous and quartzite horizons, are known to host copper mineralisation in several copper systems within the belt. At Mitsui’s Quechua deposit near Tintaya, and also at the Antillas porphyry copper deposit near Antabamba (north of Constancia), these rocks are the main host for the copper mineralisation. The same style of mineralisation occurs at Haquira (10 km south of Las Bambas), where Murco sediments are the host for most of the copper oxide resource of the system. At Constancia, these types of clastic sediments, especially the Murco or Chilloroya Formation which is the host for the copper oxide mineralisation at Haquira, have been identified in the southern sector of the property, around the Chilloroya village, where recent surface exploration has identified evidence of porphyry-related copper-gold-molybdenum mineralisation associated with the sediments and altered porphyry rocks. Common to these clastic sediments is their iron-stained colouration, which comes from oxidation of former disseminated pyrite. Glacial moraines cover the northern and eastern margins of the Constancia deposit. To the east these moraines entirely cover potentially important extensions of copper mineralisation along broad east-west structural zones. 7.2.2

Intrusions

Multiple phases of monzonite and monzonite porphyry characterise much of the surface area of the prospect, as well as dominating the rock types observed in the drilling to date. At this stage of the understanding of the deposit, at least four main phases of intrusion are recognised, with the second oldest being associated with the main mineralisation event. From oldest to youngest they are: Diorite: while not part of the intrusive event associated with mineralisation, the Andahuaylas-Yauri Batholith forms the ‘Intrusive Basement’ to the Constancia deposit. Monzonite Porphyry 1 (MP1): this unit outcrops as a large stock on the Constancia hill, extending west to San José. It hosts most of the porphyry-related mineralisation. It is characterised by abundant (40-50%) plagioclase phenocrysts up to 3 mm long. Hornblende constitutes 5-7%, with elongated crystals up to 6 mm long. The matrix is pinkish, with smaller magnetite and orthoclase crystals. The upper parts of this stock are mostly leached, making its recognition difficult. Micro Monzonite Porphyry (MMP): characterised by a fine-grained texture with plagioclase crystals (60-70%) up to 2 mm long. Biotite (1-3%) and magnetite (<1%) are visible in the light gray matrix. Within this body, wide (1-7 mm), rounded alteration patches occur, simulating xenoliths. This body outcrops as a stock in the south, but is more widespread west of the deposit, including the San José zone. Some dykes of MMP have seen cutting the MP1 porphyry.

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Quartz Monzonite Porphyry (QMP): this occurs mostly in the Constancia zone, as wide, NS to NNW trending dykes with dark, fine–grained chilled boundaries, misclassified sometimes as andesites in previous logging. Abundant (40-60%) plagioclase with tabular, well-formed hornblende crystals (7%) occur as phenocrysts in a greenish matrix. The most distinctive characteristic is a small amount (1-2%) of quartz ‘eyes’ three to four millimetres wide. These dykes are up to 60 m wide, and, based on sections, most are vertical. There is no mineralisation associated with the dykes. This unit appears to be offset by a dextral fault in the vicinity of drillholes CO-05-024 and CO-05-036. Monzonite Porphyry 2 (MP2): this dyke-like monzonite porphyry outcrops mostly to the west of Constancia, between the Constancia and San José zones. It is characterised by abundant (60-70%) plagioclase phenocrysts up to 4 mm long. These crystals are normally sub-rounded in a whitish matrix with little magnetite and mafics (hb>bio) for about 5-8%. The monzonite porphyry occurs as dykes up to 150 m wide, which strike north-south, possibly with a steep easterly dip. Andesite (AAN): dark-gray, aphanitic rock (greenish from chloritisation), with plagioclase and hornblende phenocrysts, and with magnetite about 1%. This occurs mainly as narrow dyke-like bodies, some of them close to the contacts with quartz monzonite porphyries. Additionally several brecciated units and minor felsic dykes occur in the area. 7.2.3 7.2.3.1

Alteration Potassic alteration

The potassic alteration assemblage is characterised by secondary potassium-feldspar, and by variable amounts of hydrothermal biotite replacing earlier ferromagnesian minerals and rock matrix. Quartz veining is common, especially “A” and “B” type veinlets. Intensity of alteration is variable, ranging from weak to strong. Hydrothermal magnetite is also present as disseminations and associated with A type veinlets in deeper sections. Anhydrite veinlets are also common. Within the potassic alteration zone, chalcopyrite-(bornite)-molybdenite-pyrite mineralisation is present in A and B veinlets, and also replacing ferromagnesian minerals or filling fractures. Copper grades in general vary from 0.2%-4%, and are highest where fracture-filling style copper mineralisation is superimposed on earlier disseminated copper mineralisation. The high-grade (hypogene) copper mineralisation is hosted by a dense A-veinlet stockwork developed in an early porphyry phase. Pyrite/chalcopyrite ratio is typically low and in the order of 1:1 to 2:1. Molybdenite also commonly increases with depth, related to “B” veinlets. Bornite occurs sporadically especially at deeper levels, sometimes associated with gold values. 7.2.3.2

Propylitic alteration

Propylitic alteration is transitional to the potassic alteration and extends more than one kilometre from the porphyry intrusive contacts. The propylitic alteration mineral assemblage includes epidote-chloritecalcite-pyrite-rhodochrosite. Subordinate chalcopyrite is also present, filling fractures or replacing mafic minerals. Sphalerite-galena veinlets and veins are distributed as a halo to the copper-molybdenum mineralisation within the propylitic alteration halo, occurring at distances of up to 3 km away from the porphyry copper system.

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7.2.3.3

Phyllic alteration

Phyllic alteration forms a pervasive carapace surrounding and sometimes overprinting potassic alteration. This alteration accompanies almost complete destruction of primary rock textures. The phyllic alteration mineral assemblage includes sericite-quartz-pyrite, limited amounts of chalcopyrite and associated occasional “D” veins and veinlets. 7.2.4

Structural geology

As with most porphyry copper complexes, structural activity at Constancia has played the most significant role in preparing and localising the hydrothermal alteration and accompanying coppermolybdenum-silver-gold mineralisation, including skarn formation. Major inter- and post-mineral fracture systems in the deposit area strike northeast. One of the best known is the ‘Barite Fault’, and reference is made to the ‘Barite’ direction. The Barite system is formed by a number of nearly parallel vein-faults carrying base metal sulphides and barite. These veins have been exploited by artisanal workings throughout the property. This system is clearly visible in the ground magnetic maps, controlling major features including topographic and shear fabric changes in the San José zone. A second important system strikes north-south. It seems to be more recent than the Barite system, controlling part of the San José deposit and most of the silicified breccias (some of them mineralised) in the system. This is the same direction as that of the post-mineral dykes, and may have originated as tension gashes to the Barite direction. Latest in the sequence is the NNW-SSE oriented fault system, with the Yanak Fault as the main example. These faults generate wide areas of gouge and milled rock, some of which show high hydraulic gradients. 7.2.4.1

Barite fault

The Barite Fault is a late, northeast-southwest striking structure between the Constancia and San José zones. The fault zone is generally 5-10 m in width, and characterised by broken and brecciated monzonite-barite-quartz-copper oxides and galena. In detail, the vein shows an en echelon structure, with tight northeast sections followed by east-northeast and even east-west tension gashes which contain better grade mineralisation. Numerous exploration pits and tunnels are located along this structure for more than 1000 m. Parallel veins occur, showing similar strike and mineralisation. Owing to the scarcity of outcrops and the changing strike, the Barite Fault is defined as a zone with single defined faults as opposed to a shear zone, where the individual faults/veins have not been defined. The Barite Fault has late post-hypogene copper movement that post-dates and limits the Constancia zone on its north-western side. The fault system may also control the San José zone on its south side. It is possible that the San José and Constancia zones may be offset equivalents along a left-lateral Barite Fault. 7.2.4.2

San José pit wall fault

This is a fault that bounds the eastern side of the San José pit. It strikes north-south, but meanders in its northern extension, where it may be associated with the Yanaccaca silicification zones.

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7.2.4.3

Yanak Fault

This fault belongs to the latest structural event in the area. The structure strikes NNW-SSE and is located between the Constancia and San José deposits. It has been recognised for about 3 km, and has developed a zone of influence up to 50 m wide where gouge and milled rock are present. This type of structure commonly shows high water conductivity.

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8.

DEPOSIT TYPES

The Constancia deposit is a porphyry Cu-Mo-Ag system which includes copper-bearing skarn mineralisation. This type of mineralisation is common in the Yauri-Andahuaylas metallogenic belt where several porphyry Cu-Mo-Au prospects have been described but not exploited. The principal porphyry deposits in the belt include Antapaccay (720 Mt @ 0.54% Cu, Xstrata plc Half Yearly Report 2009), Quechua (~260 Mt @ 0.61% Cu, Mitsui Kinzoku press release, Nov 7, 2007 ) and Los Chancas (~200 Mt @ 1% Cu, 0.08% Mo, 0.12 g/t Au). Several other porphyry prospects are also being explored in the district. Historically, the belt was better known for copper skarn deposits such as Tintaya (130 Mt @ 1.7% Cu, 0.26 g/t Au) and Las Bambas (copper skarn-porphyry) where Xstrata recently reported resources of 1100 MT @ 0.77% Cu (Xstrata plc Half Yearly Report 2009)..

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9.

MINERALISATION AND ALTERATION

Five distinct mineral associations are found within the Constancia Project area, namely: 1)

Hypogene, porphyry-style mineralisation including disseminated, quartz-vein stockwork and fracture-controlled chalcopyrite-molybdenite mineralisation in the intrusive

2)

Hypogene chalcopyrite, rare bornite, galena and sphalerite mineralisation in skarns

3)

Supergene digenite-covellite-chalcocite (rare native copper) mainly hosted by intrusive, lying below a leached cap

4)

Transitional (Mixed) including secondary copper sulphides/chalcopyrite in the monzonite (overlap of 1 and 3, above)

5)

Oxide copper mineralisation.

Of these, the hypogene mineralisation (Type 1) constitutes the bulk of the deposit, extending to well below the 3900 m level. Type 2, skarn, is volumetrically smaller, but grades are normally higher, and mineralisation occurs at or near the surface. At the contact between the intrusives and limestones, magnetite ± garnet skarn develops, while the pyroxene–diopside (garnet–epidote) association is more common in calcareous sandstones and arkoses of the Chilloroya Formation. Supergene enrichment (Type 3) occurs immediately beneath, and occasionally as remnants within, the leached cap. The highest copper grades in the Constancia porphyry are typically associated with this and with the skarn zone. Transitional (Mixed) zone (Type 4) corresponds to the zone where the supergene and hypogene mineralisation mix, e.g. both supergene and hypogene sulphides co-exist. Oxide copper mineralisation (Type 5) occurs locally. While shallow, it is volumetrically small and, therefore, is not considered relevant to exploitation at this stage of development. Two areas of porphyry-style mineralisation are known within the project area, Constancia and San José. At Constancia, mineralisation is deeper than that observed at San José which occurs at surface. The mineralised zone extends about 1200 m in the north-south direction and 800 m in the east-west direction. Several new exploration areas were identified in the surroundings of Constancia, other than “Yanaccaca” and “Old Adits” already mentioned in the December 2007, NI 43-101 Technical Report. The new areas include the Pampacancha, Eastern Uchucarco and South Chilloroya, all of them highlighted by the recent Induced Polarisation (chargeability/resistivity) and ground magnetometry geophysical surveys carried out on the property during 2008. Some exploratory holes were drilled at the Pampacancha area during August-September 2008. 9.1

CONSTANCIA

The Constancia porphyry has been evaluated by 317 diamond drillholes, some to depths of more than 600 m, with a maximum of 675.80 m (CO-08-133). The majority of the mineralisation occurs as

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disseminations, quartz vein stockworks and fracture filling of sulphides, mainly associated with the “Monzonite Porphyry 1 (MP1). When mineralised, MP1 shows either extensive quartz-sericite or potassic alteration. Copper mineralisation is best developed in the central part of the deposit within the monzonite porphyry and is open at depth below the level 3800 m (e.g. CO-06-083, CO-07-105, CO-08-233, CO-08-278 and CO-08-288). Oxidation and leaching are intense, with almost no fresh sulphides occurring at surface. A leached cap occurs to variable depths, up to a maximum of 100 m where fracturing is more intense, or in rocks with intense stockwork development. Oxidation decreases towards the margins of the prospect and in the magnetite skarn, reaching only a few tens of meters depth. The main iron-oxide minerals are jarosite and goethite with lesser amounts of hematite after supergene chalcocite. Typical copper grades in the leached cap are in the order of 100-200 ppm, values considered as strongly geochemically anomalous. Molybdenum and gold values in the leached cap are generally similar to hypogene grades. 9.2

SAN JOSÉ

The San José zone lies some 350 m west-northwest of the western limit of the Constancia deposit and is separated from it by the Barite Fault Zone. Sixty drillholes have been completed in the area; most of them intended to identify the extension and depth of the mineralisation. Major controls for the area appear to be two faults that bound it on the east (San José Fault) and west. These two faults strike generally north-south, but would merge into one at the Yanaccaca zone. Mineralisation in this area corresponds mostly to hypogene type assemblages, with skarn mineralisation playing a secondary role.

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10.

EXPLORATION

Exploration is currently underway in the project area, and is focused on areas/targets described in the following sub-sections. 10.1

SURFACE MAPPING AND SAMPLING

During the first months of 2007, geological mapping at 1:1000 scale was conducted at and surrounding the Constancia deposit, covering an area of approximately 450 hectares. Owing to the scarcity of outcrops, most of this mapping was made along roadcuts, and combined with information from some trenches. First rock intercepts from drillholes were also used to help define the different rock types at surface. Locations were based on GPS measurements. From April to October 2008, surface mapping was focused on the areas where geophysical anomalies were identified at depth as well as the surface reconnaissance of the projected waste dump and plant facilities. By October 2008, additional 2200 hectares were mapped at 1:2000 and 1:5000 scale including the collection of +900 rock samples and 41 stream sediment samples, the latter in the eastern sector of the Constancia deposit where scattered gold occurrences were observed in an area of 4.5 km by 0.8 km, now known as the Pampacancha prospect. Surface mapping resumed on mid February 2009 to the south of the Pampacancha prospect (SE of the Constancia deposit), and the Arizona and San Antonio areas (the latter recently acquired by Norsemont), where the tailings storage facility (TMF) is planned to be built. By 31 March 2009, an additional 1500 ha had been mapped and about 350 rock samples collected. From April to June 2009 mapping was carried out in the Chilloroya South area, where several evidences of porphyry-related copper-gold-molybdenum were found. This area is still under evaluation. Summarizing, from year 2007 to June 2009, about 5540 ha were mapped in the Constancia project at several scales, including 1:1000, 1:2000 and 1:5000; this represents some 25% of the total Norsemont mining concessions in the area. Additionally, 1890 rock samples and 41 stream sediments samples were collected during this period. 10.2

GEOPHYSICS

An in-house interpretation of the geophysical data along with interpretation of available surface mapping and rock and stream sediment geochemistry helped identify several targets within the project area. Currently, the most important ones are the anomalies associated with the Pampacancha prospect, the chargeability-magnetic anomalies observed in the Chilloroya South prospect and the chargeability anomalies located in Uchucarco, at 3.8 km northeast of the Constancia porphyry.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 10.1 Constancia Project – Exploration Targets

10.2.1.1

Pampacancha prospect

The Pampacancha prospect is located 3 kilometres SE of the Constancia porphyry. The prospect was identified past May 2008 after a stream sediment survey revealed a 27 sq-km, Au-Ag-Cu anomalous area, which was subsequently corroborated by mapping and rock sampling conducted on the area.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 10.2 Pampacancha Geology Map

Pampacancha is also coincident with a striking NE structural break portrayed by the ground magnetometry survey, along with several chargeability anomalies that occur at depth. The prospect comprises scattered outcrops of limestone and minor fine grained clastic sediments intruded by dioritic and lesser monzonite intrusive which generated magnetite and lesser garnet-calc silicate skarns at their contacts. On surface high grade Au-Ag-(Cu) mineralisation associated with veins, shear zones and limestone replacements occur in an area of about 6 km2. The longest structure can be projected up to one kilometre in length. Gold and silver returned values up to 39 g/t Au and 38 oz/t Ag.

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Figure 10.3 Pampacancha Prospect – Exploratory Drilling

Exploratory drilling carried out in Pampacancha South during August-September 2008 defined a high grade skarn and porphyry exploration target with preliminary dimensions of 1,000 by 400 m. Preliminary results indicate at least two zones of skarn mineralisation over this 1,000 m strike length that requires follow-up drilling. Four holes intercepted significant mineralisation (PR-08-008, 009, 011 and 012). The best intercept is in hole PR-08-008 which reported 43.50 m @ 1.70% Cu, 0.1% Mo, 10.35 g/t Ag and 0.76 g/t Au between 112.50 and 156.00 m. Among a number of narrow copper-gold and gold only intercepts encountered in this reconnaissance drilling program, some of the most relevant are those reported in hole PR-08-003, located 950 m NE from hole PR-08-010, where the last 19.5 m (from 235.50 to 255.00 m) averaged 0.78 g/t Au, with the last sample (1.5 m length) assaying 0.81 g/t; and hole PR-08-005, located about 1000 m to the west of hole PR-08-010, which intersected 15 m of highgrade Cu-Mo-Au skarn averaging 0.68% Cu, 0.048% Mo and 0.46g/t Au.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 10.4 Pampacancha Prospect – Main Drilling Intercepts

10.2.1.2

Chilloroya South prospect

This highly prospective area is located 5 km south of Constancia porphyry. Evidence of porphyryrelated copper-gold-molybdenum mineralisation and copper-bearing skarns occur in an area of about 3.5x3.5 km coincident with several composite chargeability and magnetic anomalies at depth. Strong evidence of porphyry-related copper-gold-molybdenum mineralisation occurs at the southern sector in an area of about 2.5x2 km. At the western sector a series of EW-oriented quartz-limonite brecciated structures hosted by feldspathic sandstones occur in an area of about 500x500 m. On the hill crest oxidation and leaching of former sulphides have been strong, leaving limonite crusts and gossanous areas where former sulphides were massive. About 50 m down slope there are green and black copper shows on fractures in feldspathic sandstones, visible at only few tens of centimetres below surface.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 10.5 Chilloroya South – Geological Map

Of 152 rock samples taken from this area, 52% returned values from 0.1 up to 7.84 g/t Au, 85% were anomalous in copper, returning from 70 ppm up to a maximum of 1.33% Cu and 40% were anomalous in molybdenum, with values in the order of 10 up to 446 ppm Mo. This area is coincident with a magnetic high and an 800x900 m chargeability anomaly, the latter extending westward to an area covered by post-mineral quaternary deposits. One kilometre to the south, there is another hill containing several SE-oriented quartz-tourmalinelimonite brecciated structures which occur in an area of 750x340 m. Evidence of multistage brecciation events has been seen in outcrops, including massive quartz-tourmaline replacements, quartztourmaline-limonite (after sulphides) matrix-supported breccias and massive limonite (sulphides) brecciated structures. A total of 48 samples were collected directly from the quartz-tourmaline breccias, where 79% were anomalous in gold, from 0.1 g/t Au up to 5.32 g/t Au; 65% anomalous in copper, from 71 ppm Cu up to 693 ppm Cu and 53% anomalous in molybdenum, from 8 ppm Mo to 75 ppm Mo. This area is coincident with a large, 1000 x 900 m chargeability anomaly at depth, partially coincident with a magnetic high. Extending below the post-mineral quaternary alluvium westward of the edge of the brecciated hills, two composite, NS-oriented, 1200 x 800 m chargeability anomalies occur.

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Approximately 2 km to the northeast of the quartz-tourmaline breccias, a 2 m wide, SSW-oriented, quartz-sericite-altered felsic porphyry is emplaced discordantly within fine-grained hornfelsed siltstones. The felsic porphyry is strongly altered, showing randomly-oriented quartz veinlets and carrying pyrite, chalcopyrite and bornite. Three samples taken from the porphyry returned up to 2.1% Cu, 32 ppm Mo and 265 ppb Au. The porphyry is thought to be associated with a 500 x 700 m chargeability anomaly located right to the east of the porphyry outcrop. Several copper occurrences hosted by the Chilloroya sediments associated with shear zones and veinlike structures occur also in the area. At a distance of 500 m southwest of the felsic porphyry there is an open cut exposing green copper oxides with randomly-oriented chalcopyrite veinlets; mineralisation is not well exposed on surface, but starts to be clearly defined just 0.50 m below surface. Five samples taken from this area returned up to 1.49% Cu, 2.88 g/t Au and 521 ppm Mo. The evidence of mineralisation reported for Chilloroya South strongly suggests the presence of a large copper-gold-molybdenum system at depth. Excellent potential exists for the discovery of additional mineralisation of this style and/or porphyry copper-gold-molybdenum mineralisation in other phases of the porphyry bodies. 10.2.1.3

Uchucarco chargeability anomaly – skarn target

This anomaly is located 3.2 km northwest of the Constancia porphyry, consisting of a magnetic anomaly at least 1 km wide and 0.8 km long, coincident with a chargeability anomaly of 0.7 x 0.4 km which remains open below 250 m depth (Figure 10.1). 10.3

EXPLORATORY DRILLING

During year 2008 exploratory drilling was concentrated at the Pampacancha prospect, 3 km southeast of the Constancia porphyry, where 20 holes were drilled totalling 5735.25 m. Some porphyry-related Cu-Mo-Au skarn intercepts were registered in holes PR-08-008, 010, 011 and 012. Additional exploratory drilling is warranted to evaluate this highly prospective area. Preliminary exploration at the Yanaccaca area (about 900 m north of San José) started in early 2007. Yanaccaca is a large (approximately 1000 x400 m) magnetic anomaly with a north-south to N10E orientation, coincident with the border of a similarly oriented resistivity low IP anomaly. Hole CO-07109 intercepted a zone of 38.7 m with an average of 1.6% Cu at a 0.5% Cu cut-off. The host to the mineralisation is a magnetite-bearing skarn. A second hole drilled from a similar location, but directed at 90o from hole 109 failed to intercept the mineralised zone, as it was probably drilled parallel to the main structure. Two additional holes were drilled in this area after November 2008 (holes CO-08-291 and CO-08-293), both failing to intercept significant copper mineralisation. The area is still not well defined and additional exploration is warranted to better define the skarn mineralisation at depth. A number of exploration holes were drilled in the zone between Constancia and San José to assess the continuity of the mineralisation between both areas. This drilling indicates that mineralisation continues between these two zones at depth.

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Drilling on the southwestern side of the Constancia resource indicates a continuation of the skarn mineralisation. Hole CO-08-229 returned two high-grade skarn interceptions, 58.50 m @ 1.53% Cu and 54.50 m @ 0.95% Cu. Additional drilling is planned to test the mineral potential of the existing and newly identified anomalous areas and targets, the latter defined after the evaluation of the recently produced geophysical data and information from surface mapping and rock geochemistry.

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11.

DRILLING

As of 18 June 2009, a total of 132 130.35 m (451 holes) have been drilled in the Constancia Project (this includes 7484.15 m drilled by Rio Tinto prior to 2005), in six drilling programs (infill, condemnation, metallurgical, geotechnical, hydrogeological and exploration). Drilling comprises both diamond drilling (core recovery) and reverse circulation (chip recovery). Diamond drilling constitutes 90% of the total metreage. Diamond drilling constituted all the drilling until May 2008 when a reverse circulation rig started on the project. Since 2005, GEOTEC has been drilling on the property and during year 2008 they had six rigs on site (four Longyear 44 and two UDR650 rigs). AK Drilling started operating in May 2008, using a Foremost Prospector 4x4 reverse circulation rig. MCA was the contractor for the geotechnical drilling between May 2008 and February 2009, operating one Atlas Copco CS-1000 rig. Since March 31 2009, two UDR 200-DLS of AK Drilling continued geotechnical drilling in the TMF facility until completion of this work in April 2009. Finally, since May 2009, a Barber Foremost DR24 rig of AK Drilling began drilling of four large diameter water holes along the Chilloroya River and the surroundings of the Constancia pit. Table 11.1 Drilling Programmes by Year (in metres drilled) Company

PQ

HQ

Rio Tinto (2003-2004)

6 965.15

NOM 2005

9 799.05

NOM 2006

19 682.25

NOM 2007 NOM 2008

3 380.75

NOM 2009 (to June 6th) GRAND TOTAL

3 380.75

NQ

RC

519.00

HOLES

TOTAL

24

7 484.15

41

9 799.05

722.15

66

20 404.40

23 863.75

5 197.35

77

29 061.10

39 497.45

7 379.60

12 792.70

219

63 050.50

1 808.50

113.65

409.00

24

2 331.15

101 616.15

13 931.75

13 201.70

451

132 130.35

General characteristics of the drilling programs carried out in Constancia are described below. Infill Drilling ( prefix CO): started prior to 2005 with the holes drilled by Rio Tinto and ended after December 2008, totaling 109 206.75 m distributed in 317 holes (including 7484.15 m, 24 holes, drilled by Rio Tinto). Information gathered from this program is used for resource estimation of the Constancia deposit. All the metreage was diamond drilling. Condemnation Drilling (CR): started in May 2008 and ended in September 2008. Drilling was carried out at the surroundings of the pit outline and sterilisation of the plant facility locations, including the testing of some chargeability anomalies south of the Constancia pit outline. Drilling totaled 8751.70 m in 54 holes. All of this drilling was carried out with a reverse circulation rig.

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Metallurgical Drilling (CM): started in September 2008 and completed by mid-November 2008. Twenty-one holes were drilled with PQ line totaling 3380.75 m. This drilling allowed the collection of approximately 35 t of samples for metallurgical tests conducted in Chile. Geotechnical Drilling (CG): from May 2008 to April 2009 drilling was completed in the surroundings of the pit, the waste dump and the TMF facilities, totaling 3716.75 m in 26 holes. Hydrogeological Drilling (CH): from January 2009 to May 2009 drilling was completed in the waste dump and TMF facility areas, totalling 928.15 m in 9 holes. Additionally, as of June 18th 4 large diameter holes (water holes) were completed along the Chilloroya River and at the southern part of the Constancia pit, totalling 409 m. Exploration (PO, PR): during 2008, exploration drilling was fully dedicated to the Pampacancha prospect. A total of 20 holes were drilled during August-September 2008, 15 being reverse circulation (4041.00 m) and 5 core drilling (1696.25 m) totaling 5737.25 m. Table 11.2 Drilling Programmes (to 6 June 2009) Program

Company

Constancia Infill (CO)

Norsemont/Rio Tinto

Metreage

Holes

109 206.75

317

Completed

Condemnation (CR)

Status

Norsemont

8 751.70

54

Completed

Metallurgical (CG)

Norsemont

3 380.75

21

Completed

Geotechnical (CG)

Norsemont

3 716.75

26

Completed

Hydrogeological (CH)

Norsemont

928.15

9

Completed

Hydrogeological (CH)

Norsemont

409.00

4

Completed

Norsemont

5 737.25

20

First

132 130.35

451

Barber Rig (water holes) Exploration (PR,PO)

Phase

Completed TOTAL

11.1.1

Collar location

Drillhole locations are initially defined using hand-held GPS. The drillhole collars and elevations are then surveyed by a surveyor every three months or as necessary. The instruments used accord with the date of the survey, with Total Station, and with differential GPS being used mostly. All measurements are tied to the National Grid. UTM coordinates based on the Provisional South America 1956 (PSAD56) datum are used throughout the project.

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11.1.2

Rig setup

The borehole azimuth is set up by marking front and back sights with a Brunton compass. After the rig is set up, the azimuth is checked by the geologist. Inclination is measured with the inclinometer incorporated into the rig. Comparisons between the planned and downhole survey measurements (done between 5 to 20 m below the surface to allow for the casing) have shown variations under 0.5o, with only one case being greater than 1o. Since November 2007, the initial inclination has been measured with a more precise digital inclinometer, and the front and back sights are left in the field to be measured by the surveyors at a later date. Once the rig is positioned, the senior geologist on site approves the location and set-up. During the drilling, a Norsemont-appointed drilling supervisor checks the compliance with the drilling procedures. The same supervisor is responsible for the correct measurements of downhole surveys at the end of the drilling. 11.1.3

Downhole survey

Downhole deviation surveying (showing both dip and azimuth) has been completed at reasonable intervals down the hole for most of the drillholes. Instruments that have been used for this task are the Eastman Single Shot, Flexit and Maxibor. It is clear that measurements in the southern end, where magnetite is more common, should be repeated with a non-magnetic system. In the case of the photographic records (Eastman), these are read by the senior geologist, and entered into the database. Measurements not within range are flagged as invalid in the database. . The Maxibor and Flexit data is delivered both electronically and as in a printout. Flexit data contains measurements of the magnetic field intensity which is used to validate the data. All data is input into the database, and anomalous values are flagged as invalid. Flexit data has been collected at 30 m intervals since January 2007. Prior to this date, the measurements were made at 50 m intervals, proceeding upwards from the base of the hole. A final measurement is made just below the surface casing (5-20 m below surface). If, for any reason, the rods have to be removed prior to completion of the hole, then the hole is surveyed at that time as a safety measure.

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12.

SAMPLING METHODS AND APPROACH

Sampling methodology and approach for Rio Tinto drilling and Norsemont 2005 drilling was detailed in the McCrae (2006) Technical Report. Sampling methodology and approach for Norsemont 2006 to March 2007 drilling was detailed in the Snowden (2007) Technical Report. GRD Minproc has reviewed the methods used in previous drill programs and considers them appropriate for a Mineral Resource Estimate. 12.1

DRILLHOLE SAMPLING METHODS NORSEMONT 2007-2008

12.1.1

Sample collection

Core trays are delivered to the core facility at the camp site where they are digitally photographed and marker blocks and depths checked. A technician records geotechnical measurements. Gross geological intervals are logged. Preliminary 2 m sample breaks are marked, with the geologist deciding on smaller intervals to coincide with geological features where necessary. A minimum sample length of 0.3 m has been specified. Sample numbers are assigned, and then detailed lithological and mineralogical logging takes place. Cut lines are marked on the core by the geologist. For heavily fractured core, a steel bar is used as a divider. The same side of the core is always taken for the sample. Samples for density measurement in each major rock unit are extracted at this stage, at approximately 50 m intervals. Samples are transferred to technicians for splitting into two halves by diamond saw. QAQC control for the project requires the provision of field duplicates. These are obtained by cutting the required samples into quarters. All samples (one half core, two quarter core) are left in the box for the samplers to collect. This takes place in a separate work area; split core is then returned for placement in canvas bags, tagging (inside and out) and weighing. Blank, standard and duplicate samples are inserted, and sample bags are then packed into larger bags for transport. The remaining half core is separately stored, with security guards on site 24 hours a day. Samples are securely stored before being loaded onto covered and secured trucks. Samples were transported on a regular basis to the ALS Chemex shipping point at Arequipa until June 2008, and since July 2008 to date samples have been sent to the SGS shipping point (also in Arequipa) where they are transferred to authorised trucks for shipment to the laboratory in Lima. Chain of custody documents with signatures of delivering and receiving parties and the names of persons accompany the samples at all times. Sampling information, including date, hole, sample interval, sample ID, sample type, QAQC type and sampler is entered on site into Excel spreadsheets, which are ultimately transferred to the database.

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12.1.2

Drillhole logging

The core is geologically and geotechnically logged by site-based geologists, using standard terminology and procedures, which were put in place in October 2006. Previous drillholes (including Rio Tinto holes) have been re-logged and re-coded. Core logging takes place in a sheltered and comfortable core shed, with sufficient space to lay out core trays for several holes. Core logging includes: 

Drillhole summary information, e.g. project details, hole number, logged by, collar coordinates, azimuth and dip, downhole survey information, geological summary and other relevant comments.



A geotechnical log with interval, recovery, RQD, fracture type and fill, frequency, shape, roughness, intensity, width, angle and details of lost core.



Lithology, alteration, mineralisation and structure – type, association, intensity and proportion are recorded for each category.



Sampling details with sample interval and number, and QAQC sampling information.

Data is logged on paper and then entered by hand on site into standardised Excel spreadsheets. Transcription checks are made by Norsemont on one in every 10 records. Data is finally uploaded into a custom-designed Access database in Lima, where further validation and integrity checks are carried out. Only one senior geologist has access to and can change the primary database. Paper logs are periodically photo-copied; originals are retained on site, with copies being sent to the Norsemont office in Lima. 12.1.3

Density measurements

A total of 1247 density measurements have been made for core from the Constancia-San José area. The density measurements are conducted by ALS Chemex and are representative of the different rock and mineralisation domains recognised to date. 12.1.4

Sample preparation, analysis and security

Sample preparation and assaying for Norsemont was carried out by ALS Chemex in Peru until June 2008, after which SGS del Peru continued with this work. Both laboratories are Peru-based registered laboratories, conforming to ISO 9000 and ISO 9001 standards, respectively. Samples are prepared and analysed in Lima following standard procedures. Samples are routinely analysed for gold (Fire Assay, AAS finish, 30 g charge) and 41 elements by ICP (HNO3-HClO4-HF-HCl digestion, HCl Leach). Samples above detection limit are analysed by AAS. All samples with copper values above 0.2% are analysed by a Sequential Copper Method. Analytical data is delivered electronically by the laboratory in Norsemont’s format, and is input directly into the company database.

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12.2

CORE SAMPLING

Individual samples respect geological/mineralogical/structural boundaries, and are generally collected at 2 m intervals. Minimum core length has been established at 30 cm, and the maximum is 3 m. The shortest sample recorded is 0.35 m (CO-06-094) and the longest is 4 m (CO-05-035). Three core saws are used, working two 12 hours shifts when required. The saw facilities are spacious and allow easy handling of the core boxes, which are stored in order before starting the shift. Each saw has a table besides it where the box containing the core to be cut is deposited. Each piece of core is cut along the centre line and, in cases where the core is disaggregated or is easily broken it is wrapped in plastic adhesive tape to facilitate handling. The cut core is then moved to the other side of the room. QAQC control for the project requires the provision of field duplicates. These are obtained by cutting the required samples into quarters. All samples (one half core, two quarter core) are left in the box for the samplers to collect. When a number of boxes have been cut they are re-arranged by number and taken out of the saw shed for sampling. Samplers have been instructed always to take the same side of the core samples. The left half is placed in the pre-numbered canvas bag. Long core pieces are broken to fit in the bag. The sample number is written with black, waterproof marker on both sides of the bag, and one of the sample tags is placed in the sample bag. A written record is kept indicating the drillhole, and the start and end metreage for every shift. 12.3

QAQC PROCEDURES

The Norsemont QAQC procedures include the use of field duplicates, certified standards and blanks. The sample numbers are selected randomly before sampling the core. The sample tickets are appropriately marked and the sample tags ripped off so as to avoid any confusion. All control samples are collected and inserted during the core sampling process. 12.3.1

Field duplicates

Field duplicates are obtained by splitting half core samples, obtaining two quarter core sub-samples, one quarter representing the original sample and the other quarter representing the duplicate sample. These samples are collected to assess the homogeneity of the mineralisation and sampling precision. Field duplicates are inserted by Norsemont at the proportion of 1 in 20. 12.3.2

Blanks

Blanks are introduced at a 1:20 frequency. These samples have been provided by a third laboratory (SGS del Peru) and have been produced from barren material, where particle size is +10 mesh. Originally (up to drillhole CO-06-062, mid-2006) blanks were obtained on-site from country rocks. However these blanks were not homogeneous and at least one of them had high zinc values, making them unsuitable for quality control.

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12.3.3

Standards

Standards (or certified reference materials, “CRM”) are samples with established grades, prepared under special conditions by certified commercial laboratories. The standards for Norsemont have been prepared by SGS in Lima from coarse rejects of previous drilling from Constancia. Norsemont uses four standard levels, characterised by Cu grades of approximately 2000 ppm, 5000 ppm, 7500 ppm and 25 000 ppm. Three generations of standards have been used in the project. During the first phase, Ag and Zn were not reported for the standards. The recommended values are summarised in Table 12.1. Table 12.1 Recommended Values for Standards Period

Standard

Cu_ppm

Mo_ppm

Ag_ppm

Zn_ppm

GQ601193

2452.4

+/-

122.8

86

+/-

5.5

May

GQ601194

5165.7

+/-

316.25

319

+/-

22.4

2006

GQ601195

7383.6

+/-

332.8

65

+/-

8.5

GQ601196

12 500

+/-

550

159

+/-

8.5

MV600011

2052

+/-

122

74

+/-

5.5

2.3

+/-

0.8

1231

+/-

70

November

MV600013

5052

+/-

229.5

127

+/-

10

3.5

+/-

1.1

929

+/-

81

2006

MV600014

7503

+/-

215.5

81

+/-

7

4.8

+/-

0.8

1321

+/-

87

MV600015

24 476

+/-

899

120

+/-

11

6.4

+/-

1.7

138

+/-

20

MV700038

1888

+/-

73

101

+/-

8

5.6

+/-

0.6

280

+/-

10

MV700039

4858

+/-

228

70

+/-

8

3.6

+/-

0.5

778

+/-

27

MV700040

7617

+/-

301

58

+/-

9

3

+/-

0.3

495

+/-

20

MV700041

24 917

+/-

762

117

+/-

14

14.9

+/-

0.9

2242

+/-

GEO-1611

37

+/-

3

9

+/-

2

0.8

+/-

0.4

135

2007-8

62 14

The standards are inserted sequentially into the sample stream at an overall proportion of 1 in 20 (i.e. 1 in 80 for each individual standard). The grades of these samples remain “blind” to the analytical laboratory. The primary (ALS Chemex) and secondary (ACT) laboratories are not involved in the preparation of the standards (SGS Laboratory). 12.3.4

Other QAQC samples



Pulp duplicates are second splits of the pulps. These samples are routinely analysed by the laboratory. While these samples are good indicators of the assay reproducibility (precision) they are not blind for the laboratory, and their value is therefore diminished. These are inserted at a proportion of 1 in 20.



Pulp blanks (internal laboratory samples), are samples of pulverised barren material. These samples provide a check on contamination during assaying. The pulp blanks are inserted immediately after highly mineralised samples. These are inserted at a proportion of 1 in 20.

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

Check samples (same pulp, externally analysed) are second splits of pulps resubmitted to an external certified laboratory under a different sample number. These samples are used to estimate the assay accuracy, together with the standards. These are submitted at a proportion of 1 in 20.



Coarse duplicates (or preparation duplicates) are splits of some samples taken after the first crushing and splitting step; these samples provide information about the sub-sampling variance. These are inserted at a proportion of 1 in 20.

12.3.5

Referee laboratory

ACME has been selected as the referee laboratory for the Constancia drilling program. Pulps and coarse rejects have been sent for check analysis. Standards and blanks are inserted in these check batches as in a standard batch. 12.4

STATEMENT ON SAMPLE PREPARATION AND ANALYSIS

GRD Minproc is of the opinion that Norsemont’s site-based sample preparation procedures are of industry standard. Similarly, the chain of custody and security procedures is considered to conform to industry best practice. GRD Minproc is of the opinion that the Norsemont QAQC sampling protocol is rigorously set up and is continuously monitored to identify potential sampling and assaying problems.

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13.

SAMPLE PREPARATION, ANALYSES AND SECURITY

Sample preparation, analysis and security are detailed as part of the foregoing Section 12.

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14.

DATA VERIFICATION

14.1

NORSEMONT INTERNAL DATA VERIFICATION

14.1.1

Collar location

Drillhole locations are initially defined using hand-held GPS. The hole collars and elevations are later surveyed by a surveyor every three months or as necessary. Collar azimuth and dip set-up are checked and approved by Norsemont geologists; front and back sights are left in place to be checked by surveyors at a later date. 14.1.2

Downhole survey

During drilling, a Norsemont-appointed drilling supervisor checks compliance with the drilling procedures. The same supervisor is responsible for the correct measurements of downhole surveys at the end of the drilling. 14.1.3

QAQC data verification

Analytical results are assessed by Norsemont staff for deviations from the norms; values from duplicate samples are compared, and quality control charts are drawn for standards and blanks. The decision points are as follows: 

For duplicates, if more than three values have variations greater than 30% the batch is rejected



For standards, three values more than 2 standard deviations, or two values if with the same sign, or a constant deviation from the average value are sufficient to reject the batch



For blanks, three or more values above the detection limit, or two consecutive values above the standard deviation are sufficient to reject the batch



Internal laboratory standards are also reported by the main laboratory, and are compared to the assumed values before data entry. The same decision parameters are used for detecting invalid batches.

No major differences have been detected in the analysis issued by the main laboratory. In case of any errors, a complete re-analysis of the batch is requested. 14.1.4

Database generation and validation

All lithologic, alteration, geotechnical and mineralisation data is logged on paper logs that are later entered in spreadsheets from where they are imported into the database. The paper logs have a similar structure to the spreadsheets to facilitate data entry (repetitive information, like logging date and geologist is recorded once per sheet, at the top of the page). The data entry spreadsheets have a number of in-built checks to facilitate the correctness of the data. For example, there are provisions that prevent incorrect or mistyped codes in certain columns. Ranges are checked for consistency, as well as for formatting issues. After the data is delivered to the senior geologist, batch tests are again carried out on the data.

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Before data upload, 10% of the entry rows are checked by comparing the written logs with the typed data. This check is done by a geologist in the field (not the data entry clerk). Upload into the database is through specially written programs that also check the data for consistency. Assay data is delivered by the main laboratory in digital form. Checks for extreme values are made by the senior geologist before data upload. Special programs are used to upload analytical data to the database, checking the data for completeness and formatting. The data are also cross-checked with the sample and collar data previously input into the database. Collar positions are checked visually on plans for correctness in the data entry. Downhole surveys are checked by examining coarse changes in the variables and also graphically in special purpose charts. Check runs are at regular intervals to check consistency of the drilling data. End-of-hole measurements, gaps in sampling and/or logging are revised and corrected when detected. 14.2

GRD MINPROC DATA VERIFICATION

14.2.1

Drilling

GRD Minproc reviewed drill and core handling procedures whilst on site, and concluded that the drilling contractor (Geotec) is carrying out drilling according to industry standards. 14.2.2

Sampling

GRD Minproc reviewed the complete sampling process on site, from the drill rig through to the final bagging of samples for dispatch to the laboratory. GRD Minproc considers that sampling process is carried out to acceptable industry standards. 14.2.3

QAQC data verification

14.2.3.1 Blanks A total of 1727 blank results were reviewed for the time period covering the data used in the resource model. Twenty samples returned >0.02% Cu, with only three of these (or 0.17%of the blanks) reporting above 0.1% Cu. The sequence of Cu blanks is illustrated in Figure 14.1.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 14.1 Cu Blanks (%)

Of 1727 blanks assayed for Mo, 64 returned greater than 10 ppm and 2 above 20 ppm. The sequence of Mo blanks for the Norsemont drilling is illustrated in Figure 14.2. Figure 14.2 Mo Blanks (ppm)

Of 1727 blanks assayed for Ag, 6 returned greater than 1 ppm. The sequence of Ag blanks for the Norsemont drilling is illustrated in Figure 14.3.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 14.3 Ag Blanks (ppm)

Of 1727 blanks assayed for Au, 19 returned greater than 1 ppb. The sequence of Au blanks for the Norsemont drilling is illustrated in Figure 14.4. Figure 14.4 Au Blanks (ppb)

Of 1727 blanks assayed for Pb, 42 returned greater than 10 ppm and nine greater than 20 ppm. The sequence of Pb blanks for the Norsemont drilling is illustrated in Figure 14.5.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 14.5 Pb Blanks (ppb)

Of 1727 blanks assayed for Zn, 13 returned greater than 0.005%. The sequence of Zn blanks for the Norsemont drilling is illustrated in Figure 14.6. Figure 14.6 Zn Blanks (%)

14.2.3.2 Standards Norsemont Standards A total of 1065 standard assays were available and have been reviewed. Apart from a few examples from early stages of the drilling program, most standards fall within two standard deviations of the recommended value. One of the standards used in the early drilling programs (GQ601195) shows a bias in Cu and Mo, and two of the later standards (MV700039 and MV00038) show a slight bias in Cu and Mo.

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Some examples of recent standards are illustrated in Figure 14.7, Figure 14.8, Figure 14.9 and Figure 14.10. Figure 14.7 Cu Standard MV700041

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 14.8 Cu Standard MV700040

Figure 14.9 Cu Standard MV700039

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 14.10 Cu Standard MV00038

14.2.3.3 Field duplicates Field duplicate ICP data (below 1% Cu) have a correlation coefficient of 0.93. A correlation plot is illustrated in Figure 14.11. Figure 14.11 Cu ICP Field Duplicate Correlation Plot

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Field duplicate data for high grade Cu (over 1% Cu, assayed by AAS62) have a correlation coefficient of 1.00. A correlation plot is illustrated in Figure 14.12. Figure 14.12 High Grade Cu Field Duplicate Correlation Plot

Field duplicate data for Mo ppm have a correlation coefficient of 0.89. A correlation plot is illustrated in Figure 14.13. Figure 14.13 Mo Field Duplicate Correlation Plot

Field duplicate data for Ag ppm have a correlation coefficient of 0.88. A correlation plot is illustrated in Figure 14.14.

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Figure 14.14 Ag Field Duplicate Correlation Plot

Field duplicate data for Zn ppm have a correlation coefficient of 0.80. A correlation plot is illustrated in Figure 14.15. Figure 14.15 Zn Field Duplicate Correlation Plot

In general, with the additional data acquired as part of the 2007-2008 drilling programs, correlation coefficients have improved.

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Pulp duplicates for Cu, Mo, Ag and Zn were reviewed and all had correlation coefficients of 0.99 or 1.00, as should be expected from internal laboratory checks. Coarse duplicate data was not available for analysis. 14.2.4

Collar location

GRD Minproc visited several drillhole platforms whilst on site, but was unable to view drill rig set-up procedures. Recent rig alignment practices use more accurate digital inclinometers, and set-ups are left in place for later checking by surveyors. All Norsemont holes used in the resource estimate have now been independently surveyed by Differential GPS. Permanent concrete monuments are now in place for all holes, and a number of these were viewed during the site visit, including the very first hole drilled by Norsemont at Constancia (Figure 14.16). Figure 14.16 CO-0501 Collar Monument

14.2.5

Downhole survey

GRD Minproc was unable to view downhole surveying on site. The detailed database of all methods of downhole surveying applied to holes at Constancia was, however, reviewed, and it was confirmed that the most appropriate survey data had been used in the resource estimate. Downhole traces of desurveyed data were viewed in section, plan and 3-D in both Micromine and Datamine Studio software. GRD Minproc considers that these traces appear realistic. The downhole survey data itself was also validated to ensure there were no breaks or sharp changes in orientation. GRD Minproc concludes that the downhole survey data is of an appropriate standard for use in the current resource estimation.

Page 138


14.2.6

Sample database integrity

GRD Minproc checked the database against ten randomly selected drillholes for transcription errors during data entry; none were found. A random selection of six Microsoft Excel spreadsheets received from the laboratory was compared to the database to check for data transfer/merging problems; none were found. GRD Minproc performed a series of validation steps during the import of data from the CSV and Access database formats provided. This data was imported into both Micromine and Datamine software, and routine checks were carried out as follows: 

Surveyed collar RL location compared with RL projected from topography digital terrain model



Existence of 0 RL azimuth and dip readings



Correct sequence of downhole survey data



Correct sequence of from-to increments in interval files



No missing intervals in interval files



Assignment of nulls/zeroes as appropriate to negative or character (<) assay values



Checks for errors during the desurvey process.

GRD Minproc considers the final data sets available for input to resource estimation are as error-free as is practicable to verify. 14.2.7

Independent samples

No independent samples were collected and submitted by GRD Minproc. 14.2.8

Core recovery considerations

Average core recovery is 95.9% (Figure 14.17), with the vast majority of recoveries greater than 90%. To ensure there were no issues with recovery within mineralised zones, this was checked also (Figure 14.18); average core recovery in mineralised zones is 96.3%, with a large majority of recoveries greater than 90%.

Page 139

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 14.17 Core Recovery

Figure 14.18 Core Recovery Distribution in Mineralised Samples

Page 140


15.

ADJACENT PROPERTIES

A number of mineral occurrences are known to exist in the general Project area, but outside the Norsemont properties. Intrusive rocks ranging from diorite to monzonite in composition, belonging to the Abancay Batholith of Eocene-Oligocene age, occur extensively to the west, southwest and north of the Constancia porphyry. To the east, their continuity is truncated by the occurrence of MioceneQuaternary volcanic rocks which define a clear NE-oriented lithological contact. Large remnants of Cretaceous rocks composed of limestones of Ferrobamba Formation and clastics of Chilloroya and Soraya Formations occur immediately north and south of the Constancia porphyry, respectively. Among the main mineral occurrences are the old Katanga mine, located about 5 km northwest of the Constancia porphyry, identified by several small Cu (Ag-Au)-bearing skarns developed at the contact between limestones and intrusives of monzonitic composition. These were exploited extensively during the past century, before the mine was shut down in the early 1990s. Monte Rojo and Sorapata are other occurrences in the area. They are located about 8 km northnorthwest of the Constancia porphyry. Both prospects are Cu-bearing skarns bodies developed at the contact between small intrusive plugs and limestones of the Ferrobamba Formation. In addition, there are a number of small Cu (Au-Ag-Zn) skarn and vein-like structures dispersed throughout the area, some of them having been mined in the past as artisanal to small-scale operations. These include the Aurea mine (12 km northwest of Constancia), where copper oxides were exploited from vein-like and small skarn bodies up to 4 m wide, developed at the contact between granodiorites and Ferrobamba limestones.

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16.

MINERAL PROCESSING AND METALLURGICAL TESTING

16.1

PRE-DFS TEST WORK

16.1.1

Iditec-Chile

Preliminary metallurgical tests using drill core rejects were performed by Iditec-Chile in 2004. Initial work indicated that there was insufficient leachable copper for a viable copper leaching option. Rougher flotation tests, which were conducted on single samples from two drillholes (KA006 and CO008) achieved recoveries as shown in Table 16.1, suggesting that flotation recovery of valuable minerals was a technically viable option (Lytle, 2004). Table 16.1 Iditec-Chile Flotation Recoveries Drillhole

Metal Recoveries (%) Copper

Gold

Molybdenum

Silver

KA006

80

74

37

67

CO008

90

83

71

63

16.1.2

Plenge Laboratories

During 2007 and 2008, Plenge Laboratories in Lima, Peru conducted a test program to provide information for preliminary flowsheet development. This included Phase 1 and Phase 2 flotation tests, comminution tests through SGS Chile, mineralogical analysis and Phase 1 and Phase 2 flotation variability tests by SGS Canada. 16.1.2.1

Plenge Phase 1 flotation test work

The objective of the Plenge Phase 1 test work was to obtain the metallurgical response of mineralised material from the Constancia deposit to flotation and to provide preliminary design information for a Scoping Study. This involved head assay, specific gravity and natural pH determination, preliminary flotation reagent screening, and the effect of primary grinding, regrinding and pH on flotation performance. Optimum conditions were then used in locked cycle tests to approximate plant metallurgical results. The core sample sets selected for metallurgical testing were fresh quarter core samples collected from diamond drillholes drilled during 2006. The samples were considered representative of the mineralisation zones identified for the Constancia deposit and were designated as either Hypogene, Supergene or Skarn. Three master composites were produced. The optimum flotation conditions for the Hypogene composite were a primary grind d80 of 150 µm, a regrind d80 of less than 37 µm, using A-3302 as a primary collector and Z-14 as a secondary collector, at a pH of 9.0, with 16 minutes batch flotation rougher residence time and three stages of cleaning. These conditions, when used in locked cycle tests, produced the results as shown in Table 16.2. Copper and silver recoveries were acceptable for unoptimised tests, although molybdenum recoveries were below industry accepted benchmarks. The combined zinc plus lead content of the concentrate

Page 142


was 4%. Rougher concentrate mass recovery to the cleaner circuit of 20% of flotation feed was approximately double the industry average. Table 16.2 Plenge Laboratories Phase 1 Locked Cycle Test Results – Hypogene Ore Concentrate

Metal Recoveries (%)

Test

Grade

Copper

Silver

Molybdenum

%Cu

6448 - 25

89.4

78.2

35.2

29.55

6448 - 26

86

64.1

10.5

32.04

The optimum flotation conditions for the Supergene composite were a primary grind d80 of 150 µm, a regrind d80 of less than 37 µm, using A-3302 as a primary collector and Z-14 as a secondary collector, at a pH of 9.0 with 16 minutes batch flotation rougher residence time and three stages of cleaning. These conditions, when used in locked cycle tests, produced the results as shown in Table 16.3. Copper and silver recoveries were acceptable for unoptimised tests, although molybdenum recoveries were at the lower end of industry accepted benchmarks. The combined zinc plus lead content of the concentrate was not a concern, being below 1%. Rougher concentrate mass recovery to the cleaner circuit was considered high. Table 16.3 Plenge Laboratories Phase 1 Supergene Locked Cycle Test Results Concentrate


Test

Grade

Copper

Silver

Molybdenum

%Cu

25

87.3

88.5

52.0

31.37

26

88.7

91.3

45.1

27.93

The optimum flotation conditions for the Skarn composite were a primary grind d80 of 106 µm, a regrind d80 of less than 37 µm, using A-3302 as a primary collector and Z-14 as a secondary collector, at a pH of 9.0, with 16 minutes batch flotation rougher residence time and three stages of cleaning. These conditions, when used in a locked cycle test, produced the results as shown in Table 16.4. Copper and silver recoveries were acceptable for unoptimised tests, although molybdenum recoveries were well below industry accepted benchmarks. The combined zinc plus lead content of the concentrate was excessive at >5%. Rougher concentrate mass recovery to the cleaner circuit was approximately double the industry average. Table 16.4 Plenge Laboratories Phase 1 Locked Cycle Test Results – Skarn Ore Test 6449 - 13

Concentrate


Grade

Copper

Silver

Molybdenum

%Cu

88.2

51.6

14.1

26.62

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The Phase 1 Plenge Laboratories test work indicated that the Constancia Project could employ a conventional Cu-Mo porphyry flotation circuit. 16.1.2.2

Plenge Phase 2 flotation test work

The objective of the Plenge Phase 2 test work was to provide additional preliminary design information for the DFS by optimising the flotation response of copper and molybdenum, and evaluating zinc depression. This involved extensive flotation reagent screening, assessing the effect of regrinding on flotation performance and evaluating the use of zinc depressants. Optimum conditions were then used in locked cycle tests to approximate plant metallurgical results. The optimum flotation conditions for selected locked cycle tests are shown in Table 16.5, with flotation performance detailed in Table 16.6. The Phase 2 tests produced comparative concentrate grades and recoveries to those of Phase 1 for Hypogene and Supergene ores. The various molybdenum collectors trialled failed to achieve any significant improvement in molybdenum recovery to the bulk copper-molybdenum concentrate. For Skarn ore treatment the concentrate grade achieved was low due to dilution by zinc and lead, with the use of sodium cyanide proving the most effective zinc depressant. The Skarn ore investigation found that zinc recovery by flotation from the copper tailing may be warranted. Table 16.5 Plenge Laboratories Phase 2 Locked Cycle Test Results (July 4, 2008) Ore Type

Condition Primary Grind d80 (µm) Rougher pH

Supergene

Hypogene

Skarn

150

150

150

9

9

9

Rougher Res. Time (min)

12

12

12

Regrind d80 (µm)

25

27

37

Cleaner pH

11

10.5

7.7 / 10.5

Cleaning Stages

4

3

4

Lime (g/t)

910

910

735

Diesel (g/t)

25

25

25

NaCN (g/t)

50

40

Zinc Sulphate (g/t)

90 650

A-3477 (g/t)

30

30

30

Z-14

18

30

18

Dow 250

30

30

30

MIBC

42

24

30

Page 144


Table 16.6 Plenge Laboratories Phase 2 Selected Locked Cycle Test Results Concentrate


Test Supergene No. 17

Grade

Copper

Silver

Molybdenum

%Cu

83.6

59.3

18.0

27.55

Hypogene No. 26

88.1

63.7

38.9

27.79

Skarn No. 24

85.7

53.3

25.6

21.72

16.1.3

SGS Chile test work

16.1.3.1

Leaching

Twenty-one drillhole core assay reject samples were subjected to bottle roll leach testing at SGS Chile. Results confirmed that leaching recoveries were low compared to flotation, and that leaching/SX-EW was not a viable copper recovery route for the Constancia Project. 16.1.3.2

Comminution

Preliminary comminution tests were also conducted at SGS Chile, with the results shown in Table 16.7. The Hypogene ore had a significantly higher Bond Rod and Ball Mill Work Index compared to both Supergene and Skarn samples. As such, the Hypogene ore was classified as medium-hard while both the Skarn and Supergene ores were classified as medium-soft ores. The Minnovex SPI index showed that Hypogene ore would have the highest resistance to SAG milling with the Skarn sample SPI index somewhat lower and Supergene ore lower still. The Minnovex Crushing index (Ci, derived from the SPI test) is considered by GRD Minproc to have no direct relationship with ore crushing properties and, therefore, was not used for assessment of crushing properties. The Bond Abrasion index for Hypogene ore is significantly higher than either Supergene or Skarn ores, which will result in higher ball consumption and liner wear when processing this ore. Table 16.7 Historical SGS Chile Comminution Test Work Ore Type Test

Hypogene

Supergene

Skarn

Bond Rod Mill Wi

kWh/t

14.3

10.9

11.2

Bond Ball Mill Wi

kWh/t

15.8

12.9

11.2

75.2

39.8

64.5

Test screen size

150 µm

Minnovex SPI

min

Minnovex crushing index, Ci

26

29

15

Bond Abrasion index Ai

0.173

0.1162

0.0922

Specific gravity

2.75

2.75

3.66

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16.1.4

SGS Canada flotation variability tests

Twenty samples of Constancia RC rejects and drill core were supplied to SGS Canada for flotation variability testing and mineralogical analysis. Two sets of conditions were used for rougher flotation variability testing as shown in Table 16.8. Table 16.9 shows the results from the tests on Supergene ore, Table 16.10 shows the results from the Hypogene ore, and Table 16.11 shows the results from Skarn ore tests. Table 16.8 SGS Canada Rougher Flotation Variability Testing Conditions Condition

Grind d80 (µm)

SIBX (g/t)

A3302 (g/t)

pH

Residence

Frother

Time (min) 1

150

10

20

9

16

As required

2

150

40

30

9

16 / 19

As required

Table 16.9 SGS Canada Rougher Flotation Variability Testing Results – Supergene Ore Sample

Condition

Feed Grade

Concentrate

Rougher Copper

(% Cu)

Grade (% Cu)

Recovery (%)

NSC06

1

1.21

4.5

83.4

NSC07

1

1.38

7.3

71.8

NSC08

1

0.52

3.4

92.9

NSC09

1

0.78

4.9

93.4

NSC10

1

0.43

1.5

87.4

NSC11

1

0.42

2.7

81.4

NSC12

1

0.43

1.9

87.9

NSC13

1

0.80

4.4

85.3

NSC14

1

0.52

2.0

70.5

NSC15

1

0.50

3.6

71.0

Average

1

0.7

3.62

82.5

NSC06

2

1.21

4.2

90.9

NSC07

2

1.38

3.6

85.6

NSC10

2

0.43

1.6

90.4

NSC11

2

0.42

3.1

92.1

NSC12

2

0.43

1.5

93.2

NSC13

2

0.80

3.7

91.6

NSC15

2

0.50

3.6

95.1

Average

2

0.74

3.04

91.2

Supergene ore samples varied in head grade from 0.42% Cu to 1.38% Cu. Supergene treatment with the lower collector addition rates produced low copper recoveries, ranging from 70.5% to 93.4%, to achieve an average of only 82.5%. The increase in collector addition in the second series of tests produced a significant increase in rougher copper recovery to 91.1%. In both cases the rougher

Page 146


concentrate grade showed marked variation, independent of both feed grade and recovery. appears due to variation in the operator’s flotation technique.

This

Table 16.10 SGS Canada Rougher Flotation Variability Testing Results – Hypogene Ore Sample

Condition

Feed Grade

Concentrate

Rougher Copper

NSC01

1

(% Cu)

Grade (% Cu)

Recovery (%)

0.38

1.8

71.1

NSC02

1

0.88

4.5

81.6

NSC03

1

0.61

3.5

79.2

NSC04

1

0.41

2.3

74.2

NSC05

1

0.84

4.5

76.6

Average

1

0.62

3.32

76.54

NSC01

2

0.38

2.5

84.2

NSC02

2

0.88

5.6

80.8

NSC03

2

0.61

4.6

80.0

NSC04

2

0.41

2.4

92.9

NSC05

2

0.84

2.8

89.7

Average

2

0.62

3.58

85.52

The variation in Hypogene ore head grade was less than that of Supergene, ranging from 0.38% Cu to 0.84% Cu. Increased collector addition in the Hypogene variability tests also resulted in a significant increase in rougher copper recovery from 76.5% to 85.5%. The increase in copper recovery was achieved with no significant effect on rougher concentrate grade. Table 16.11 SGS Canada Rougher Flotation Variability Testing Results – Skarn Ore Sample

Condition

Feed Grade

Concentrate

Rougher Copper

(% Cu)

Grade (% Cu)

Recovery (%)

NSC16

1

0.68

2.1

82.8

NSC17

1

0.21

1.4

82.2

NSC18

1

0.72

3.9

78.1

NSC19

1

0.72

3.0

68.7

NSC20

1

1.10

6.6

72.2

Average

1

0.69

3.4

76.8

NSC16

2

0.68

2.4

89.5

NSC17

2

0.21

1.2

89.6

NSC18

2

0.72

3.4

93.5

NSC19

2

0.72

1.6

86.9

NSC20

2

1.10

3.4

93.7

Average

2

0.62

2.4

90.64

The skarn ore samples exhibited the highest rate of head grade variability, ranging between 0.21% Cu and 1.1% Cu. Again, an increase in collector addition resulted in a significant increase in rougher copper recovery (from 77% to 91%). In the Skarn tests, the increased copper recovery was matched

Page 147


by a marked decrease in concentrate grade, from 3.4% Cu to 2.4% Cu. Although zinc assays were not available it is expected that the concentrate grade decrease was caused by greater flotation of sphalerite into the concentrate. 16.2

DFS METALLURGICAL TESTWORK

The DFS metallurgical test program was designed to provide process parameters that could be used in the design of the Constancia plant flow sheet. The test program was conducted at bench scale and pilot scale and consisted of treatment of individual and composite samples, selected on specific mineral zones. In addition a number of samples from mixed zones were supplied for individual sample tests. The ore zones, basic ore type, approximate number of samples and approximate weights are detailed in Table 16.12. Representative samples were selected for metallurgical testwork, taking account of lithology, mineralisation type, grade and three-dimensional location. To facilitate location and depth discrimination, the ore body was divided into “metblocks”, each metblock being a 100 m cube. Continuous core runs within the metblocks that met the sample selection criteria were then identified. From the available core, a sample set that provided broad spatial representation and was relevant to the likely mine plan for the ore body was selected. Composites were prepared from diamond drill core from all ore zones, except the Mixed Zone samples. These composites were used for flow sheet development tests including grinding, flotation and regrinding. The bulk density of the ore to minus 6 mesh size (~ 3.36 mm), specific gravity (picnometer method), natural pH and the consumption of lime to a pH of 12.0 were determined for each composite. Table 16.12 Ore Zone Samples Average Grade

Ore Description % Cu

g/t Mo

% Zn

g/t Ag 4.77

Supergene

1.13

184

0.04

Skarn – Medium Zinc

1.64

240

1.83

Skarn – High Zinc

0.73

70

1.86

Hypogene

0.56

362

0.05

4.00

Mixed Zone

0.90

487

0.10

4.71

Mine Blend

1.28

230

0.93

Pilot Plant

Number of

Total sample

Samples

weight (kg)

13

1092

9

601

4

204

19

1621

6

462

n/a

240

n/a

Bond Ball Mill tests, Bond Abrasion tests and SMC tests were conducted to determine SAG and ball milling requirements. The flotation program include primary grind and regrind optimisation, reagent optimisation and cleaner circuit development. A number of rougher tests were conducted to provide tailing samples for rheological study. The optimum conditions were then used in locked cycle tests. Products obtained from the respective optimisation tests were analysed for Cu, S, Fe, Zn, Pb and Mo.

Page 148


The final locked cycle test concentrates were analysed for Al, As, Au, B, Ba, Be, Bi, Ca, Cd, Co, Cr, Hg, K, Li, Mg, Mn, Na, Ni, P, Pb, Sb, Se, Sn, Sr, Ti, Tl, U, V, W, Zn, total sulphur, S as sulphate, carbonates, fluorides, chlorides and Re. In addition they were submitted for mineralogical assessment through QEM-Scan. The rougher tailing and the cleaner scavenger tailing of the locked cycle tests were analysed for Al, As, Au, B, Ba, Be, Bi, Ca, Cd, Co, Cr, Hg, K, Li, Mg, Mn, Na, Ni, P, Pb, Sb, Se, Sn, Sr, Ti, Tl, U, V, W, Zn, total sulphur, S as sulphate, carbonates, fluorides, chlorides and Re. Variability tests were performed on each individual sample using the optimum flotation conditions, to evaluate the behaviour of the deposit to the flowsheet. Products obtained from the respective variability tests were analysed for Cu, S, Fe, Zn, Pb and Mo. In addition SMC and Bond ball mill work index tests were conducted on individual samples to determine the throughput over the mine life of 15 years. The pilot plant test program objective was to generate sufficient sample for a laboratory based molybdenum flotation program. Molybdenum rougher kinetic tests were conducted to determine flotation residence time, molybdenum promoter, NaHS addition. The molybdenum concentrate produced was subjected to mineralogical and chemical analysis. In addition, the copper-molybdenum concentrate was used for regrind tests, production of final copper concentrate, and rheology, thickening and filtration tests. Samples of copper rougher and cleaner scavenger tailings were provided to Knight Piésold for rheology and thickening tests. 16.2.1

Mineralogy

In Supergene ore 63% of the copper present was chalcopyrite, with chalcocite (23%), bornite (10%) and covellite (2%) making up the majority of the remainder. Copper minerals that are not considered recoverable by flotation represented less than 2% of the total copper present. The main non-economic sulphide mineral was pyrite, which amounted on average to 4.17% of the sample. Other sulphides represented less than 0.2% of the sample. Non-sulphide gangue was dominated by quartz and Kfeldspar, although there was also a significant amount of sericite/muscovite. Clays represented approximately 2% of the sample. In Hypogene ore, copper was present predominantly as chalcopyrite (approximately 90% of copper). Minor copper minerals included chalcocite (8%) and bornite (2%). Copper minerals that are not considered recoverable by flotation represented less than 1% of the total copper present. Non-sulphide gangue was dominated by quartz and K-feldspar and there was also a significant amount of sericite/muscovite. Clays represented approximately 2% of the sample mass. Pyrite constituted 3.13% of the sample on average, while other sulphides represented less than 0.1%. In Skarn ore, copper was present predominantly as chalcopyrite (approximately 90%). Minor copper minerals included chalcocite (7%) and bornite (3%). Copper minerals that are not considered recoverable by flotation represented approximately 1% of the total copper present. Pyrite constituted 4.3% of the sample on average. An appreciable amount of other sulphides were present, representing 0.71% of the sample. These were dominated by sphalerite, with an appreciable amount of galena.

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16.2.2

Comminution

The distribution of DWi and BWi results for supergene ore are shown in Figure 16.1. Supergene ore is classified as the least competent ore type at Constancia with an 80th percentile DWi of 3.7 kWh/m3. In terms of ore hardness for ball milling, Supergene ore is in the middle of the range with an 80th percentile BWi of 13.1 kWh/t. Figure 16.1 Drop Weight (DWi) and Bond Work (BWi) Index Test Results for Supergene Ore

100

100 SUPERGENE 13 samples

90

80

Cumulative Freqency (%)


90

70 60 50 40 30 20 10

80 70 60

SUPERGENE 13 samples

50 40 30 20 10

0

0

0

1

2

3

4

5 3

DWi (kWh/m )

6

7 8 9 10 11 12 13 14 15 16 17 18 BWi (kWh/t)

The average abrasion index for Supergene ore samples was in the medium range at Ai = 0.1146. The distribution of DWi and BWi results for Hypogene ore are shown in Figure 16.2. Hypogene ore is classified as the most competent ore types at Constancia with an 80th percentile DWi of 7.5 kWh/m3. In terms of ore hardness for ball milling, Hypogene ore is also the hardest with an 80th percentile BWi of 16.3 kWh/t.

Page 150

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 16.2

100

100

90

90

80

80



Drop Weight (DWi) and Bond Work (BWi) Index Test Results for Hypogene Ore

70 60 50

HYPOGENE 19 samples

40 30 20 10

70 60 50

HYPOGENE 19 samples

40 30 20 10

0

0

0 1 2 3 4 5 6 7 8 9 10 3

DWi (kWh/m )

7 8 9 10 11 12 13 14 15 16 17 18 BWi (kWh/t)

The average abrasion index for Hypogene samples was also the highest, Ai = 0.1863. Skarn samples were analysed as two groups, medium-zinc skarn (soft) and high-zinc skarn (hard). The soft Skarn is located in the upper part of the orebody and the hard skarn is deeper. There were eight (8) samples representing medium-zinc Skarn (soft) and only three representing high-zinc Skarn (hard). The distribution of DWi and BWi results for soft Skarn is shown in Figure 16.3. Soft Skarn is classified as the least competent ore type at Constancia with an 80th percentile DWi of 3.8 kWh/m3. In terms of ore hardness for ball milling, Skarn is the softest with an 80th percentile BWi of 10.6 kWh/t. Hard Skarn is classified as competent as hypogene with an 80th percentile DWi of around 7.5 kWh/m3. In terms of ore hardness for ball milling, hard Skarn is slightly harder than soft skarn with an 80th percentile BWi of around 11.5 kWh/t.

Page 151


100

100

90

90

80

80 Cum Freqency (%)

Cum Freqency (%)

Drop Weight (DWi) and Bond Work (BWi) Index Test Results for Skarn Ore

70 60 50 40

Soft SKARN 8 samples

30

70 60 50

30

20

20

10

10

0

Soft SKARN

40

0

0

1

2

3

4

5

6

7

8

7

DWi (kWh/m3)

8

9

10 11 12 13 14 15 BWi (kWh/t)

The average abrasion index for Skarn samples was also the lowest, Ai = 0.0659. 16.2.3 16.2.3.1

Flotation Copper flotation

Supergene Ore The primary grind / flotation program for the Supergene composite included d80 values of 250 µm, 200 µm, 150 µm, 106 µm and 75 µm. The ground material was subjected to 18 minutes of laboratory batch flotation with lime added to the milling and flotation steps to increase the slurry pH to 10.0. MIBC frother (10 g/t) was used, together with 30 g/t A-3033 collector. Tests to evaluate pH included pH values of 7.53 (as received pH), 9.0, 10.0, 11.0 and 12.0. Six different collectors were evaluated on the Supergene composite, as detailed in Table 16.13. Rougher flotation feed, ground to a d80 of 106 µm was conditioned with the selected collector and floated for 18 minutes at a pH of 10.0, with 10 g/t MIBC used as a frothing agent. Table 16.13 Collectors Evaluated in Supergene Rougher Flotation Product

Description

SNF3330

Sodium isopropyl xanthate

Dosage (g/t) 30

PAX

Potassium amyl xanthate

30

A3477

Sodium isobutyl dithiophosphate

30

A404

Blend of dithiophosphates and mercaptobenzothiazole

30

A3926

Alkyl thionocarbamate

30

A3302

Amyl xanthate ester

30

Page 152


The optimum flotation conditions for the supergene composite were a primary grind d80 of 106 µm, a regrind d80 of approximately 37 µm, using either SNF3330 or A-3302 as a collector at a pH of 9.0, with 12 minutes batch flotation rougher residence time and three stages of cleaning. The conditions for each cycle of the Supergene Composite locked cycle test are shown in Table 16.14. Table 16.14 Conditions for Supergene Locked Cycle Tests Reagents (g/t) Stage Grinding

CaO

A 3302

AF 65

30

15

950

Conditioning Rougher

215

Regrinding

200

Time (min)

Grind d80 µm

Cond

Flot

pH

Cumulative

106 3 10

10

37

Cleaner 1

4

11.5

Scavenger

8

11.5

Cleaner 2

3

11.5

Cleaner 3

3

11.5

The locked cycle test results indicate that when treating Supergene ore a concentrate grade of 29.6% Cu can be achieved with a copper recovery of 86% and a molybdenum recovery of 53%. Using the primary grinding and rougher flotation conditions of the locked cycle test individual Supergene samples were floated to determine the effect of ore variability on rougher flotation. Significant variation was seen in both copper and molybdenum feed grades (Table 16.15). This variation exhibited itself in widely fluctuating rougher concentrate grade, with rougher recoveries being more robust. The average copper rougher stage recovery was 89% (standard deviation = 8%), with an average molybdenum rougher stage recovery of 75% (standard deviation = 11%).

Page 153

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 16.15 Results from Variability Testing on Supergene Ore Samples Head grade

Sample

Rougher Concentrate Grade

Rougher Recovery (%)

%Cu

%Mo

%Cu

%Mo

Cu

Mo

M-1

1.16

0.030

4.11

0.109

86.58

90.01

M-2

1.12

0.022

2.79

0.058

93.92

74.20

M-3

0.57

0.009

4.90

0.033

94.15

74.79

M-4

1.11

0.004

1.86

0.011

88.60

65.99

M-5

0.62

0.004

1.98

0.012

68.21

60.24

M-6

1.24

0.008

4.02

0.021

91.45

72.06

M-7

1.99

0.035

9.39

0.155

96.81

90.85

M-8

0.69

0.025

3.87

0.110

82.37

62.79

M-9

0.84

0.029

4.23

0.117

90.70

74.24

M-16

0.82

0.006

2.58

0.015

89.50

66.08

M-21

1.84

0.046

5.13

0.122

95.12

88.48

Ave

1.09

0.020

4.08

0.069

88.86

74.52

St dev

0.5

0.014

2.1

0.1

8.0

10.9

Analysis of the Supergene final concentrate showed detrimental elements such as As, Hg, Sb were below detection limits. Bismuth was reported at the low value of 0.011% which is well below normal penalty limits. With the exception of zinc, other elements were well below possible penalty limits. Zinc (0.88% of concentrate) was below expected penalty levels, but comprises a sufficient portion of the concentrate to have a material effect on transport and smelting charges. Molybdenum comprised 0.3% of the concentrate, providing an acceptable feed stock to the molybdenum flotation plant. Carbon analysis was low, indicating that no floatable carbon minerals will impact on the molybdenum concentrate grade. Silver represented the only payable element other than copper and molybdenum in the concentrate. Hypogene Ore Flotation conditions trialled for the Hypogene composite were the similar to those for the Supergene composite. The conditions for each cycle of the Hypogene Composite locked cycle test are shown in Table 16.16.

Page 154


Table 16.16 Test Conditions for the Hypogene Locked Cycle Test Work Reagents (g/t) Stage

CaO

Grinding A

0.9

Conditioning

200

Rougher

250

Regrinding

55

A 3302

Time (min)

Grind AF 65

d80 µm

cumulative

75 30

10

3 10

10

4

11.5

8 mins

Cleaner 1 Scavenger Cleaner 2

pH

Flot

Cond

8 60

3

Cleaner 3

11.5

3

The hypogene locked cycle test results (Table 16.17) indicate that when treating predominantly hypogene ore through the concentrator a concentrate grade of 23.4% Cu can be achieved with a copper recovery of 88% and a molybdenum recovery of 75.4%. The lower final concentrate grade is largely due to the high lead content. Table 16.17 Results From Hypogene Locked Cycle Tests Product

Grade (%)

Mass (%)

Cu

Fe

Mo

S

Zn

Pb

Cleaner concentrate

2.13

23.4

22.1

1.0

27.2

1.6

4.3

Combined tailings

97.9

0.07

2.57

0.007

1.04

0.028

0.046

Rougher concentrate

11.05

3.27

6.01

0.18

5.22

0.23

0.62

Calculated head grade

100.0

0.57

2.94

0.03

1.59

0.02

0.14

Product

Recovery (%)

Cleaner concentrate

87.9

15.7

75.4

36.3

55.6

66.8

Combined tailings

12.1

84.3

24.6

63.7

44.4

33.2

Rougher concentrate

91.6

32.6

89.2

52.4

60.0

71.2

Total

100.0

100.0

100.0

100.0

100.0

100.0

Variability tests reported Table 16.18.show significant variation in both copper and molybdenum feed grades. The average copper rougher-stage recovery was 90% (standard deviation = 9%), with an average molybdenum rougher-stage recovery of 83% (standard deviation = 8%). A noticeable decrease in copper recovery was seen for samples deeper within the ore body and the possibility that a 106 µm primary grind was too coarse for the latter years of mine life was considered. To test this, samples M23, M27, M32, M38 and M39 were refloated with a 75 µm primary grind. A comparison of the copper recovery and molybdenum recovery for these two samples is shown in Table 16.19. Grinding to 75 µm produced a 7% increase in rougher copper recovery and an 11 % increase in rougher molybdenum recovery with a lower percentage variation in results.

Page 155

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 16.18 Results from Variability Testing on Hypogene Ore Samples Rgh Concentrate Grade

Rougher Recovery (%)

%Cu

%Mo

%Cu

%Mo

Cu

Mo

M-12

0.65

0.029

3.58

0.19

95.81

88.59

M-13

0.34

0.027

1.24

0.12

91.69

80.25

M-14

0.67

0.017

2.36

0.07

96.44

91.17

M-15

0.83

0.054

3.89

0.28

97.63

94.18

M-17

0.79

0.028

3.94

0.15

94.41

85.47

M-18

1.01

0.014

6.95

0.09

74.75

75.42

M-22

0.74

0.012

3.22

0.06

93.89

87.09

M-23

0.34

0.032

1.34

0.16

93.72

88.04

M-24

0.40

0.072

1.84

0.39

94.97

84.43

M-25

0.53

0.059

2.26

0.33

96.40

89.27

M-26

0.64

0.021

2.71

0.10

93.81

88.86

M-27

0.65

0.032

6.86

0.33

92.15

77.17

M-28

0.65

0.012

2.55

0.04

95.05

74.75

M-29

0.41

0.021

1.51

0.09

94.82

89.26

M-32

0.61

0.021

5.27

0.16

85.79

65.60

M-37

0.42

0.058

2.12

0.33

92.75

83.03

M-38

0.26

0.080

2.76

0.93

82.40

79.24

M-39

0.37

0.003

3.55

0.04

60.93

64.09

M-42

0.42

0.033

1.83

0.17

90.98

87.91

Ave

0.56

0.03

3.15

0.21

90.44

82.83

St dev

0.20

0.02

1.68

0.21

9.06

8.34

Sample

Head grade

Table 16.19 Deep Hypogene Samples Tested at 75 µm and 106 µm Rougher Copper Recovery (%)

Rougher Molybdenum Recovery (%)

106µm

75µm

106µm

75µm

M-23

93.72

95.24

88.04

92.58

M-27

92.15

96.41

77.17

91.14

M-32

85.79

96.50

65.60

87.92

M38

82.40

70.39

79.24

88.62

M-39

60.93

93.50

64.09

71.54

Ave

83.00

90.41

74.83

86.36

St dev

13.12

11.25

10.00

8.49

Sample

In the hypogene final concentrate detrimental elements such as As, Hg and Bi were below detection limits in the final concentrates. With the exception of zinc and lead other elements were well below possible penalty limits. Zinc (1.62% of concentrate) and lead (0.2% of concentrate) were below expected penalty levels, but comprise a sufficient portion of the concentrate to have a material effect on transport and smelting charges. Molybdenum comprised greater than 1.0% of the concentrate, although it is anticipated that this will be reduced to approximately 0.1% following treatment in the molybdenum circuit. Carbon analysis was low, indicating that no floatable carbon minerals will impact

Page 156


on the molybdenum concentrate grade. Silver represented the only payable element other than copper and molybdenum in the concentrate. Skarn Ore The optimum flotation conditions for the Skarn composites were a primary grind d80 of 106 µm, a regrind d80 of approximately 37 µm, using either AX343, A3926 or A-3302 as a collector at a pH of 9.0, with 12 to 16 minutes batch flotation rougher residence time, three stages of cleaning and zinc sulphate as a zinc depressant.. The conditions for each cycle of a medium-zinc Skarn composite locked cycle test are shown in Table 16.20. Table 16.20 Conditions Used in Locked Cycle Test Work on Medium-Zinc Skarn Ore Reagents (g/t) Stage

CaO

Grinding

400

Conditioning

30

A 3302

AF 65

30

40

d80 µm

150

Cleaner 1

90

Scavenger Cleaner 2 Cleaner 3

Cond

Flot

pH

cumulative

106 3

Rougher Regrinding

Time (min)

Grind

10

9.0

4

11.0

4 mins 8

60

3

11.5

3

The locked cycle test results are summarised in Table 16.21 and indicate that when treating predominantly medium-zinc skarn ore through the concentrator with no depressant use, a concentrate grade of 22.8% Cu can be achieved with a copper recovery of 90% and a molybdenum recovery of 56%. The lower final concentrate is largely due to the high zinc content. It is expected that with the use of zinc depressants, copper concentrates of 26% copper, with less than 5% zinc will be achievable, at a copper recovery of approximately 85%.

Page 157


Table 16.21 Results from Medium-Zinc Locked Cycle Tests Product

Mass (%)

Grade (%) Cu

Fe

Mo

S

Zn

Pb

Cleaner concentrate

5.67

22.8

24.4

0.2

32.4

9.6

1.4

Combined tailings

94.3

0.15

21.5

0.008

7.4

1.2

0.27

Rougher concentrate

13.0

8.94

20.3

0.127

18.1

3.9

0.56

Calculated head grade

100.0

1.43

21.66

0.023

8.92

1.8

0.33

90.4

6.4

56.3

20.8

31.7

23.6

Combined tailings

9.6

93.6

43.7

79.2

68.3

76.4

Rougher concentrate

89.4

12.8

80.0

28.3

31.9

24.4

Total

100.0

100.0

100.0

100.0

100.0

100.0

Product Cleaner concentrate

Recovery (%)

Elemental analysis of the final concentrate with medium zinc grade skarn ore showed arsenic below detection limits and although values of Hg were reported they were below penalty levels at 0.5 ppm and 0.29%. With the exception of zinc, bismuth and cadmium other elements were well below possible penalty limits. Both cadmium and bismuth were at values around their penalty limits at 0.0325% Cd and 0.029% Bi and will require monitoring during operation. Zinc at 9.9% of concentrate weight is well above penalty limits and if a depression strategy is not enacted may result in either low value or nonsaleable concentrate. Molybdenum comprised less than 0.2% of the concentrate. Carbon analysis was low, indicating that no floatable carbon minerals will impact on the molybdenum concentrate grade. Silver represented the only payable element other than copper and molybdenum in the Cu concentrate. 16.2.3.2


The addition of NaHS to the rougher feed resulted in 75% of copper minerals becoming non-floatable. The remaining material floated at a flotation rate one fifth of that of molybdenum. Subsequent NaHS addition to the first cleaner transferred 73% of copper minerals in the rougher concentrate to the nonfloating fraction. NaHS addition to the second cleaner was less effective with only 10% of first cleaner concentrate transferred to the non-floating fraction. The copper floating fraction in the second cleaner floated at approximately half the kinetic rate of the molybdenum. It is expected that the parameters in the second cleaner will be repeated in subsequent cleaning stages. A small proportion of lead in the rougher feed was fast floating. The addition of NaHS to the rougher feed resulted in 61% of lead minerals becoming non-floatable. The remaining material floated at a flotation rate slightly faster than that of molybdenum. Subsequent NaHS addition to the first cleaner removed the fast floating fraction and transferred 62% of lead minerals in the rougher concentrate to the non-floating fraction. There was no change in the lead flotation rate in the first cleaner. NaHS addition to the second cleaner was again less effective with only 51% of first cleaner concentrate transferred to the non-floating fraction. There was no change in the lead flotation rate in the second cleaner. Zinc mineral flotation was similar to that for copper, although NaHS addition to the second cleaner resulted in 93% of zinc minerals in the first cleaner being transferred to the non-floating fraction.

Page 158


Seventy-nine percent of non sulphide gangue in the rougher feed was non-floatable. The remainder floated at a rate similar to molybdenum. In the first cleaner no non-sulphide gangue was transferred to the non-floating fraction, indicating that NaHS was not effective in depressing the floatable non-sulphide gangue. A similar response was witnessed in the second cleaner, with no transfer of floatable material in the first cleaner concentrate to the non-floating fraction. The kinetic rate of non-sulphide gangue did decrease in the cleaning stages, but not sufficiently to prevent it reporting to the molybdenum concentrate. On the basis of the floatability component analysis a series of simulations were conducted with varying numbers of cleaner stages to replicate locked cycle tests. The bulk copper/molybdenum concentrate grade used as a feed to the simulated molybdenum flotation circuit contained 0.22% Mo, 25.0% Cu, 1.0% Pb, 5.0% Zn and 2.0% pyrite. The simulated circuit molybdenum grades and recoveries for the seven stage cleaner circuit are shown in Table 16.22. Molybdenum recovery decreased up to the fifth cleaning stage, after which a recovery of 74.88% with respect to molybdenum circuit feed was maintained. The concentrate grade increased rapidly to seven cleaning stages after which only minimal improvements were achieved above 35.83% Mo. Table 16.22 Simulated Molybdenum Grade/Recovery as a Function of Cleaning Stages Molybdenum Recovery

Concentrate Grade (%Mo)

Molybdenum Circuit Feed

100.00

0.13%

Rougher Concentrate

91.59%

1.71%

Cleaner 1 Feed

199.65%

1.85%

Cleaner1 Concentrate

79.16%

4.08%

Cleaner Sc Concentrate

103.78%

2.81%

Cleaner 2 Feed

101.90%

2.91%

Cleaner 2 Concentrate

97.62%

5.52%

Cleaner 2 Tailing

4.28%

0.25%

Cleaner 3 Feed

105.56%

4.41%


82.82%

9.91%

Cleaner 3 Tailing

22.74%

1.46%

Cleaner 4 Feed

96.03%

7.69%


88.08%

14.15%

Cleaner 4 Tailing

7.94%

1.27%

Cleaner 5 Feed

92.62%

11.72%


79.42%

21.10%

Cleaner 5 Tailing

13.21%

3.19%

Cleaner 6 Feed

80.24%

18.26%


75.70%

27.82%

Cleaner 6 Tailing

4.54%

2.71%

Cleaner 7 Feed

75.70%

27.82%


74.88%

35.83%

Cleaner 7 Tailing

0.82%

1.30%

Page 159


16.2.3.3

Flotation concentrate quality evaluation

Based on the seven stage cleaner circuit floatability component model and QEMScan analysis, the anticipated molybdenum concentrate is shown in Table 16.23. The significant diluents are expected to be talc, amphiboles, calcite, chlorite and chalcopyrite. Table 16.23 Anticipated Molybdenum Concentrate Mineralogical Content (%)

16.2.3.4

Elemental Content (%)

Molybdenite

59.78

Mo

35.83

Talc

9.49

S

32.78

Amphiboles

7.02

Pb

5.98

Galena

6.91

Si

5.83

Calcite

4.35

Fe

3.23

Chlorites

4.05

Mg

2.74

Chalcopyrite

4.03

Ca

2.50

Ti Oxides

1.13

Cu

1.40

Biotite

0.92

Al

0.86

Fe Oxides/Oxyhydroxides

0.76

Ti

0.36

Zircon

0.62

Zr

0.31

Others

0.38

P

0.05

Apatite

0.30

Na

0.05

Clays

0.26

F

0.01

Sphalerite

<0.01

Zn

<0.01

Pyrite

<0.01

FeCl3 leach

The significant amounts of copper and lead reporting to the molybdenum flotation concentrate require further reduction prior to molybdenum product shipment. A ferric chloride leach over four to six hours, at 80oC and atmospheric pressure is expected to remove 80% of copper minerals, 95% of lead minerals and 50% of silver. The pH modifier used would be hydrochloric acid, which will also dissolve the majority of the calcite and apatite present and approximately 50% of the biotite and chlorite. Considering the wide application of this technique, it is considered the most suitable for Constancia molybdenum concentrates. The reactions governing the leaching of galena and chalcopyrite are: CuFeS2 + 4FeCl3 → CuCl2 + 5FeCl2 + 2So PbS + 2FeCl3 → PbCl2 + 2FeCl2 + So Molybdenum will be only slightly taken into solution before an insoluble chloride is formed, passivating the mineral and preventing further leaching. It is expected that only 0.4% of molybdenum will be dissolved.

Page 160


16.2.3.5

Precipitation

The primary salt of environmental concern in leach liquor is lead. Effective precipitation can be conducted with lime at a pH of 11.5. Ninety-five percent of the lead will precipitate as a stable solid. 16.2.3.6

Final molybdenum product quality evaluation

Based on the ferric chloride leach details above, the final molybdenum product will have an elemental composition as shown in Table 16.24. Table 16.24 Final Molybdenum Product Elemental Composition Elemental Content

16.2.3.7

Mo

40.16

S

36.74

Si

6.53

Mg

2.89

Fe

2.39

Al

0.97

Ca

0.83

Ti

0.40

Zr

0.35

Pb

0.34

Cu

0.31

Na

0.03

F

0.01

P

<0.01

Zn

<0.01

Settling tests

Samples of copper/molybdenum concentrate and tailings were settled to determine respective thickener requirements. The effect of flocculant addition on the filtration rate and the % solids of settled concentrate after 24 hours are shown in Figure 16.4. The concentrate exhibited acceptable settling rates and settled solids (>60%) at flocculant addition rates of 10 g/t.

Page 161

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 16.4 Concentrate Settling Tests

72

1 0.9

70

0.8 0.7 66

0.6

64

0.5 0.4

62

0.3

Settling Rate (mm/s)

Settled % Solids

68

60 0.2 58

0.1

56

0 0

10

Flocculant Addition (g/t) Settled % Solids


The effect of flocculant addition on the settling rate and the % solids of settled tailings after 24 hours are shown in Figure 16.5. The tailings achieved acceptable settled solids density (>52% solids) and settling rate at a flocculant addition rate of 10 g/t.

Page 162

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 16.5 Tailing Settling Tests

54

1.6

52

1.4

Settled % Solids

1 48 0.8 46 0.6 44


1.2

50

0.4

42

0.2

40

0 0

10

Flocculant Addition (g/t) Settled % Solids

16.2.3.8


Filtration tests

All column concentrate samples were collected during the pilot plant campaign. From these samples a composite was formed which was considered representative of future copper concentrate production. Typically this was concentrate with greater than 25% Cu and less than 0.15% Mo. These samples were then filtered in a Larox laboratory pressure filter. The results of these pressure tests are shown in Table 16.25. Filtration tests indicated that the Constancia copper concentrate would be moderately difficult to filter, achieving a minimum filter cake moisture of 9.3%, with a filtration rate of 211 kg/m²h. The filtration rate could be increased to 316 kg/m²h, although the percent moisture of the filter cake would increase to 12%. It should be noted that filtration was conducted with as-settled solids, rather than flocculated solids and some benefit in filtration rate may be achieved with flocculated concentrate.

Page 163


Table 16.25 Pressure Filtration Results TEST NO. Density of slurry kg/l

1

2

3

4

1.296

1.364

1.522

1.842

Filtration Parameters Duration of pumping min

2.0

1.5

1.0

1.0

pressing I min

1.5

2.0

1.5

1.0

drying min

4.0

4.0

3.0

3.0

Calculated cycle time min

11.5

11.5

9.5

9.00

Measured Process Parameters during Filtration Tests Pressure of slurry feed bar

6.0

6.0

6.0

6.0

pressing I bar

14.0

14.0

14.0

14.0

12.5

12.1

Process Results Moisture in cake % w/w

16.3

9.3

9.7

Cake thickness mm

20.0

23.0

25.0

26.0

Wet cake weight kg

4.456

4.991

5.268

5.400

Dry cake weight calculated kg

4.042

4.507

4.610

4.747

Filtration rate (dry solids) kg/m²h

210.9

235.1

291.1

316.4

PROCESS DESIGN CRITERIA

A design rate of 5573 t/h is nominated for primary crushing to achieve the design processing plant capacity of 76 000 t/d, at an availability of 80%. 16.3.1

Grinding

A P80 of 106 µm has been selected for grinding circuit design, based on analysis of the available flotation testwork. The capacity of the grinding circuit is nominally up to 25.26 Mt/a on softer ore. This production rate was determined by practical SAG Mill assessments from similar operations, and is the optimum throughput rate for this orebody. The design criteria selected are based on the 80th percentile DWi and BWi ore parameters for ores expected in the initial years of production. This practice is commonly used by GRD Minproc to achieve a robust design that can meet the annual throughput targets taking account of ore variability. Average and 80th percentile design values are summarised in Table 16.26.

Page 164


Table 16.26 Summary of Design Ore Parameters – Grinding SG Average ore th

Design Post 5 years – 80

RWI

BWI

Ai

3 DWi (kWh/m )

(kWh/t)

(kWh/t)

2.8

15.8*

14.2

0.173*

4.6

2.8

15.8*

16.3

0.173*

6.6

2.8

11.2

11.6

Percentile Design – 80th Percentile

16.3.2

3.6

Copper flotation

Design criteria and other information regarding copper flotation are provided in Table 16.27. Table 16.27 Summary of Copper Flotation Design Criteria Copper Roughers Flow

Design

m3/h

6762

Pulp density

Design

% solids w/w

36

m3

300

Min

27.2

pH

10

Cell volume No. of cells

12

Residence time (total) Copper First Cleaners Flow

Design

m3/h

2216

Pulp density

Design

% solids w/w

25

m3

100

Min

12.2

pH

11


5

Residence time (total) Copper Second Cleaners Flow

Design

m3/h

1132

Pulp Density

Design

% Solids w/w

13.5

Ph

11

Cell Volume

m

3

No. Of Cells

100 2

Residence time (total)

Min

9.5

Copper Third Cleaner (Column) Flow

Design

m3/h

330

Pulp density

Design

% solids w/w

19

pH Column diameter

11 M

No. of columns Residence time (total)

4.27 2

Min

37

Page 165

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 16.27 (cont) Summary of Copper Flotation Design Criteria Copper Cleaner Scavenger Flow Pulp density

Maximum

m3/h

1400

Design

% solids w/w

29.5

pH

11 3

Cell volume

m

No. of cells

100.0 6


Min

24.4

Rougher Concentrate Regrind Mill type

IsaMill

No.

1

Mill new feed (each) Mill new feed (requiring regrind) Regrind work index

t/h

503

t/h

377

kWh/t

8.0

Feed size P80 (nominal)

Microns

80

Product size P80 (nominal)

Microns

25

Pinion power

Installed

Cyclone size Number of operating/standby cyclones Cyclone operating pressure

16.3.3

kW

3000

mm

250

op/standby

8/2

kPa

80


Design criteria and other information regarding molybdenum flotation are provided in Table 16.28 and Table 16.29. Table 16.28 Molybdenum Flotation Recovery and Concentrate Grade Metal Recoveries Mo

55% Concentrate Grade

% Cu

0.5

% Mo

40

% Zn

0.25

Pb

0.5

Page 166


Table 16.29 Summary of Molybdenum Flotation Design Criteria Molybdenum Roughers Flow

Design

m3/h

190

Pulp density

Design

% solids w/w

30.0

m3

10

Min

30.1

pH

11


10


Molybdenum First Cleaner Flow Pulp density

Maximum

m3/h

95.6

Design

% solids w/w

24.1

m3

10

Min

17.1

pH

11


3


Molybdenum Second Cleaner Flow Pulp density

Maximum

m3/h

56.7

Design

% solids w/w

23.7

pH

11 3

Cell volume

m

No. of cells

10 2


Min

19.2

Molybdenum Third Cleaner Flow Pulp density

Maximum

m3/h

35.2

Design

% solids w/w

20.0

pH

11 3

Cell volume

m

No. of cells

3.0 2


Min

9.3

Molybdenum Fourth Cleaner Flow Pulp density

Maximum

m3/h

21.7

Design

% solids w/w

18.4

m3

3.0

Min

15.1

pH

11


2


Molybdenum Fifth Cleaner Flow Pulp density

Maximum

m3/h

13.9

Design

% solids w/w

17.5

m3

3.0

Min

11.8

pH

11


1


Molybdenum Sixth Cleaner Flow Pulp density

Maximum

m3/h

9.4

Design

% solids w/w

16.3

pH

11 3

Cell volume

m

No. of cells

3.0 1


Min

17.4

Molybdenum Seventh Cleaner Flow Pulp density

Maximum

m3/h

6.0

Design

% solids w/w

14.0

pH

11 3

Cell volume

m

No. of cells

3.0 1


Min

27.4

Molybdenum Cleaner Scavenger Flow Pulp density

Maximum

m3/h

58.8

Design

% solids w/w

23.7

m3

10

Min

19.4

pH Cell volume

11

No. of cells Residence time (total)

2

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16.3.4

Copper thickening and filtration

Design criteria and other information regarding copper thickening and filtration are provided in Table 16.30. Table 16.30 Copper Concentrate Thickening and Filtration Specification Product Grade % Cu

26

% Mo

0.4

% Zn

2.0

% Pb

0.5 Thickener Specification

Thickener U/F solids (%)

60

2

Solids loading (t/m h)

0.35

Thickener feed rate (t/h)

74

Diameter (m)

18 Filter Specification

Filter cake moisture (%)

9

2

Filter Test Capacity (t/m )

16.3.5

0.2

Molybdenum thickening and filtration

Design criteria and other information regarding molybdenum flotation are provided in Table 16.31. Table 16.31 Molybdenum Concentrate Thickening and Filtration Specification Product Grade % Cu

0.5

% Mo

40

% Zn

0.1

% Pb

0.5 Thickener Specification

Thickener U/F solids (%)

60

2

Solids loading (t/m h)

0.14

Thickener feed rate (t/h)

0.5

Diameter (m)

2.5 Filter Specification

16.3.6

Filter cake moisture (%)

15

Filter Test Capacity (t/m2 – assumed)

0.2

Bagging moisture (%)

10

Tailings thickening

Design criteria and other information regarding tails thickening are provided in Table 16.32.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 16.32 Tailings Thickening Specification Thickener type Thickener U/F solids

52

Solids loading

2

t/m h

0.64

Thickener feed rate

t/h

3085

% solids w/w

36

m

94

Diameter

16.3.7

High Rate %

Concentrate storage

Design criteria and other information regarding concentrate storage are provided in Table 16.33 Table 16.33 Concentrate Storage Specification Copper Concentrate High Grade

Emergency

days

7

t

12 024

m3

4336

days

7

t

12 024

m3

4336

Molybdenum Concentrate Bagging

16.3.8

days

7

t

92

m3

33

Water services

Design criteria and other information regarding water services for the plant are provided in Table 16.34. Table 16.34 Water Services Specification Potable Water Potable water treatment make-up Storage

m3/hr m

3

hrs

11 528 48

Fire/Raw Water m3

432

Hydrant water requirement

L/s

10

Time for fire fighting

hrs

4

Raw water volume

m3

2506

m3/hr

365

hrs

7

Fire Water consumption during fire Number of operating hydrants

Design flow Raw water capacity

3

Page 169


16.3.9

Reagents

Design criteria and other information regarding reagents are provided in Table 16.35. Table 16.35 Reagent Specification Summary Consumption g/t feed Reagent

Type Design

AF65

Frother

26.00

A3302

Collector

30.00

Z14

Collector

10.00

Light fuel oil

Promoter

2.14

Flocculant

10.60

Ferric Chloride

1.00

Hydrochloric acid

1.00

NaHS

Depressant

23.60

ZnSO4

Depressant

0.00

16.4

PROCESS PLANT DESCRIPTION

16.4.1

Overview

The processing circuit consists of a primary gyratory crusher discharging to a coarse ore stockpile of approximately 50 000 t live capacity. Grinding consists of a single line SABC circuit. Flotation consists of a standard copper flotation circuit, combined with a molybdenum flotation circuit. Depressant will be required depending on the zinc content of mined ore, and a ferric chloride leach will be used to maximise the molybdenum concentrate grade. Concentrates are thickened via high rate thickeners prior to filtration in pressure filters. The tailings streams are combined and thickened prior to discharge to the tailings management facility (TMF). Based on the proposed mine plan and test work, a production forecast was prepared for the selected circuit. The predicted average annual SAG mill throughput varies depending on ore type, as is shown in Figure 16.6. The predicted average copper recovery and copper concentrate grade also vary, as is shown in Figure 16.7.

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Yearly production (MT/Y)

Predicted Plant Production Rates Years 1-16

22 20 18 16 14 12 10 8 6

Design All max

4 2 0

HY max

0

1

2

3

4

5

6

7

8

9

10 11 12 13 14 15 16

Production year Figure 16.7 Predicted Average Copper Recovery and Copper Concentrate Grade

Copper Recovery (%) and Concentrate Grade (%Cu)

100 90 80 70 60 50 40 30 20 10 0 1

2

3

4

5

6

7

8

9

10

11

12

13

14

15

16

Year Of Operation Copper Recovery

Concentrate Grade

Details regarding the plant design are included in Section 18.3.

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17.

MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1

INTRODUCTION

Resource estimation for the Constancia deposit is based on integrated geological and assay interpretations of information recorded from diamond core logging and assaying. Interpretation wireframe data was supplied by Norsemont, including recently updated information prepared by Atticus Associates, Lima using the Leapfrog (Version 2.1) software package. Most preliminary data preparation has been carried using the Micromine software package. Statistical analysis and variography has been carried out in the GeoAccess Professional package, with verification in Micromine and Datamine. Wireframe validation was done in both Micromine and Datamine, and correction of wireframes and flagging of data was carried out in Micromine. Rock and volume model creation was carried out in Micromine, with export to Datamine for verification. All resource estimation work was carried out using the Estima process in Datamine Studio 3, operating in double precision. Extensive use was made of the Datamine macro facilities for almost all of the data processing and analysis, cell modelling, estimation and reporting functions. A single resource block model has been constructed to cover the Constancia and San José parts of the deposit. Coordinates used for all data and models are the National Grid UTM coordinates based on the Provisional South America 1956 (PSAD56) datum. 17.2

DATA PROVIDED TO GRD MINPROC

GRD Minproc was provided with updated data for the project on 14 January 2009, with further updates and corrections being supplied through to 30 January 2009. Atticus Associates provided the data in the form of validated Excel spreadsheet files as used in the geological interpretation and wireframe generation. These files were used as the basis of the resource work to ensure that the data used matched the wireframes provided. Data files provided by Atticus were: 

dh_Alteration.csv



dh_Assay.csv



dh_AssayICP.csv



dh_Collar.csv



dh_CuSeq.csv



dh_lithology.csv



dh_Lithology.csv

Page 172




dh_Puntualcharge.csv



dh_RMR.csv



dh_RQD.csv



dh_SG.csv



dh_Survey.csv



dh_Zonation.csv

A series of detailed DXF files of the latest topography covering the whole of the Constancia project area was previously provided (January 2008). Atticus also provided four sets of wireframe model files in Vulcan wireframe model format, as follows: 

Grade shell wireframes, including 0.2% Cu and 0.15% Zn.



Zonation wireframes: Leached Oxide Cu Supergene Mixed Hypogene



Lithological wireframes: MMP

monzonite micro-porphyry

MP1

monzonite porphyry

MP2

late stage non-mineralised porphyry

QMP

quartz monzonite porphyry

SK skarn SS sandstone 

Alteration wireframes: Argillic Phyllic Potassic Propilitic Silicification

17.3

DATA PREPARATION

All spreadsheet data was imported into Micromine format files and validated. included: 

collar and downhole survey integrity



agreement between collar RL and topography

Checks carried out

Page 173




interval file from-to integrity



no From or To depths in any interval tables (assay, lithology etc) greater than the hole depth entered in the collar table



no entries in the down hole survey depth tables greater than the hole depth entered in the collar table



duplicate sample intervals



overlapping sample interval



unmatched hole identifiers in collar, survey and interval files



inappropriate downhole distances (such as negative values)



downhole desurvey and coordinate generation.

17.4

SURFACE AND SOLID WIREFRAME DATA GENERATION

Atticus and Associates used the assay, mineralogical, lithological, structural and other data as interpreted and provided by Norsemont geologists to construct 3-dimensional solid wireframe models. The geological modelling comprises three different models: the lithological model, defining the 3-D geometries of the mappable lithologies; the copper domain model, defining the distribution of the copper mineral types; and the grade model, identifying the geometry of the total copper envelope. All of the models are based on the capture of geological field data, viz. surface mapping, diamond drill core logging and laboratory analysis of core samples. 17.4.1

Lithology model

The lithology solid model is based primarily on drilling data with a secondary reference to the geology outcrop map. On occasions where the surface geology doesn’t match up with the upper few meters of the logged lithology, the logged drill data is used in preference (the outcrop in this region is limited, and the weathering profile can make identification at surface difficult). The surface map used was completed in January 2009 by Norsemont geologists. The model is defined from summary lithology data from 307 drillholes, 14 of which were drilled by Rio Tinto and the remainder by Norsemont. All of the Rio Tinto drillholes have been re-logged according to Norsemont logging and rock identification procedures, also some of the 2007 and 2008 Norsemont drillhole intervals have been reviewed and re-logged. Due to the nature of the deposit and the orientation of the drilling, it was not possible to develop a solid model from a series of cross-sections and for this reason the model was built directly in 3-D using Zaparo’s Leapfrog (Version 2.1) modelling software to complete the interpretation and build the contact surfaces and solids. The surfaces of each of the solid geometries have been defined by both ‘snapping’ contact points directly to the drill strings and honouring across- and along-strike interpretation lines.

Page 174


Some of the contacts, specifically the Skarn-Monzonite and Skarn-Limestone contacts, are gradational in nature and the contact surface has sometimes been picked as a ‘best fit surface’ between the two rock types. Solids were defined for the following lithologies: QMP – a unit that is predominantly quartz monzonite porphyry; SS – sandstone; SK - skarn, which contains material that has been logged as skarn, endoskarn and exoskarn; MMP – a monzonite micro-porphyry; MP1 – monzonite porphyry, essentially the mineralised porphyry; and MP2 – a later stage non-mineralised porphyry. The increase in amount of drilling in the south and southwest sector of the deposit since previous resource modelling has allowed for better definition of the skarn bodies and the associated marble units. A larger volume of skarn material has been identified, the geometry of which takes the form of sheet-like units extending from the marble thrust into the porphyry (Figure 17.1). The modelled skarn geometry highlights the structural complexity of this section of the deposit, Figure 17.2 and Figure 17.3 show how stacked thrust planes and vertical NW-SE faults control the emplacement and displacement of the skarn unit. Figure 17.1 Orthographic View Looking NE, Showing the Skarn and Marble Units

Page 175

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.2 Plan View Showing the Constancia Skarn Geometry

The relationship between the skarn found at San José and the main Constancia skarn is still unclear; from a plan view it appears that the San José Skarn could be an extension (Figure 17.3), however, the orientation of the mineralising structures appear to be in the opposite direction, suggesting an apparent antiform.

Page 176

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.3 Sectional View Showing the San José and Constancia Skarn Bodies

The increase in drill density across the porphyry has enabled better definition of the geometry of different phases of porphyritic intrusions. The MMP intrusion is found as large bodies to the northern and western edges of the modelled area, and as smaller dyke-like bodies in the central zone of Constancia deposit. The MP2 intrusions exist as narrow sub-vertical dykes running approximately N-S through the MP1 and MMP units. Unlike the MMP, the MP2 intrusions are always unmineralised and hence occurred at a later stage in the genesis of the deposit. The QMP intrusive unit exists as two narrow, sub-vertical, continuous dykes that cross-cut all other lithologies; these dykes are oriented in a NNW-SSE direction. Figure 17.4 shows the relationship between the different intrusives.

Page 177

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.4 Plan View Showing the Geological Model for the Constancia Project

MMP (dark pink), marble (light blue), QMP (brown), Skarn (orange), MP2 (red). MP1 is the rest of model. Drill strings coloured pink.

Plan representation is illustrated in Figure 17.5 and a 3-D image in Figure 17.6. Both are illustrated after import to the Micromine software package.

Page 178

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.5 Lithological Wireframes- Plan View

Page 179

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.6 Lithological Model Looking North West

17.4.2

Mineralisation domain model

The mineralogy model is based on logged zonation, which, since the last model, has been recoded and checked against the sequential leach copper assays values. Although primarily based on the zonation data, the mineralogy model was also reviewed with the sequential copper assay results which helped determine the contacts in the regions where the zonation data proved inconclusive. The Norsemont drill core was analysed for total copper via ICP, and four acid digestion/AAS when values were reported above the upper detection limit for ICP; those samples that returned a total copper content of above 0.2% were sent for additional analysis via sequential copper leach – first, digestion by H2SO4 (Cu_SS, acid soluble or oxide copper), followed by digestion by NaCN (Cu_CN, cyanide soluble, typically chalcocite and bornite – secondary copper minerals), then finally digestion by HCl and HNO3 (Cu_R, residual copper, typically chalcopyrite). The definition of the majority of the domains is based on the percentage of the sequential copper leach compared to total copper content. Oxide contains CuSS>50%; Mixed contains CuSS>25% and 25%50%; Transition contains 25%60%; Hypogene contains CuR>80%. The Mixed and Transition domains may contain zones of intermittent Oxide-Supergene or Supergene-Hypogene mineralisation. The zonation data was defined as LX (leached), OX (oxide), SG (Supergene enriched), MX, (mixed hypogene and supergene enriched) and HY (hypogene). Surfaces between each of these

Page 180


classifications were defined and a solid model constructed. In certain areas the mineralogical zones were not well defined by the zonation data (Figure 17.7 - shows how the modelled supergene zone contains minor sections of logged Hypogene, Leached and Mixed) so information from the sequential assay results and lithology logs were used to help guide the zonation boundary surface. In the sequential assay data, supergene is considered to contain >50% CuCN, and the mixed zone is defined as containing both supergene and hypogene material. The nature of the deposit is such that the mineralogical zone boundaries will not always be clearly defined, and the model has been created from ‘best fit’ surface that takes in data from all the evidence available. Figure 17.7 Sectional View 100m Wide Looking NE

The boundaries between the different mineralisation domains were generated using Leapfrog software, creating a surface that honours all points on drillhole strings that mark the intersection between the registered domains. A solid domain model was then created by connecting the surfaces with an outside edge denoting the limit of information surface approximately 70 m from the outer drillhole collars.

Page 181


The increase in the drill density and the change in classification since the previous modelling of the mineralogical zones has increased the volume of supergene enriched material and generated more material in the mixed category. The greatest change is towards the south of the porphyry where new holes have confirmed and increased the volume of deep supergene enrichment, albeit of mixed category. Figure 17.8 Plan View Showing the Surface Extent of Supergene Enrichment

SG (red), MX (purple), OX (orange) and the drill strings visible are coloured green for LX (leached) and blue for HY (hypogene)

The final domains produced are: 

Oxide Cu



Oxide Leach



Supergene



Mixed



Hypogene

A plan view of the main mineralisation domains is shown in Figure 17.9, while a 3-D view is seen in Figure 17.10.

Page 182

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.9 Mineralisation Domains - Plan View

Figure 17.10 Mineralisation Domains, Looking Northwest

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17.4.3

Copper grade shell model

Total copper assay data were used to develop the modelled grade shell. Due to the complexity of the grade distribution, this model has been defined via 3-D interpolations direct from the drill data using Leapfrog software. Subsequent to the mineralisation domain modelling, it was observed that the controls on the total copper mineralisation vary according to the defined domains. In order to create an accurate representation of the extension of the Cu grade across the deposit, a different interpolant is required for each of the copper domains; thus the angles and orientations of the extension of mineralisation between drill string data points within the skarn is different to that in the supergene. The resultant isosurfaces from this preliminary modelling were refined and modified using a series of control points and lines, to produce a single solid model representing the nominal 0.2% Cu grade shell. Use of the new version (Version 2.1) of Leapfrog has eliminated the issue in the 2007 model of the inclusion of ‘sub-grade’ portions of drillholes within the grade shell. The latest phase of drilling has defined a 0.2% grade shell with a slight decrease (8%) in overall volume from the previous model. Drilling towards the north and west has extended mineralisation and defined the extent of the ore body, which was previously open. Drilling in the region between Constancia Hill and San José has demonstrated that the mineralisation does not extend between the two areas and indicates that a structural control separates the two. Drilling towards the south has defined limits to the porphyry mineralisation, but has picked up copper grade in skarn material. In this region the 0.2% grade shell becomes more discontinuous and restricted in volume. The increase in drill density has resulted in better definition of the grade shell around lithological and structural contacts (Figure 17.11 and Figure 17.12).

Page 184

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.11 Plan View Showing the 0.2% Cu Grade Shell and Drill Strings

Figure 17.12 Orthographic View Looking NE Showing the 0.2% Cu Grade Shell and Drill Strings

Page 185


17.4.4

Zinc grade shell model

Zinc was analyzed by ICP followed by a second analysis by atomic absorption (with four acid digestion) for samples above the ICP upper detection limit, and the data was used to define the zinc grade shell. Modelling of the Zn distribution, which from previous investigations was shown to be limited to the skarn areas, was carried out using the Leapfrog software grade modelling algorithm as this modelled quite well the complex Zn distribution. Grade shells were generated for 0.15% and 0.30% Zn, with the 0.15% Zn shell being used to define the elevated zinc mineralisation for resource interpolation purposes. The distribution of zinc within the Constancia deposit is confined to the skarn and immediate surroundings, with grade extending into both limestone and intrusive rocks. High grade and extensive mineralisation is not only confined to the skarn, but is associated with structures that cross-cut the skarn bodies. The latest phase of drilling has defined much more intensive zinc mineralisation towards the south of the deposit, with indications that the significant zinc mineralisation extends beyond the current limit of the model. (Figure 17.13). Figure 17.13 Plan View Showing the 1500 ppm Zn Shell

Page 186

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.14 Orthographic View Looking NE Showing the Relationship between Skarn Lithology (green) and the 1500 ppm Zn Shell (light blue)

Figure 17.15 Plan View Showing the Relationship between the Skarn Lithology (green) and the 1500 ppm Zn Shell (light blue)

Page 187


Figure 17.16 Orthographic View Looking North Showing the 1500 ppm Zn Shell

17.4.5

Skarn ore type domain

A metallurgical issue has been raised related to skarn mineralisation spatially close to supergene material, and where the two material types are likely to be co-mingled during the mining/processing stage. An ore type field has been defined to cover this. Figure 17.17 and Figure 17.18 show the Skarn wireframe (blue) and the Supergene wireframe (yellow), identifying the main Constancia zone with San José to the west. Areas of obvious proximity can be seen, and are highlighted again in Figure 17.19.

Page 188

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.17 Skarn Wireframe (blue) and the Supergene Wireframe (yellow)

Figure 17.18 Skarn Wireframe (blue) and the Supergene Wireframe (yellow), with a better View of Overlap Areas

Page 189

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.19 Plan View of Skarn and Supergene Wireframes

The issue of Zn has also been considered, and wireframes have been developed to enable interpolation of ‘high-Zn’ domains. In the context of the modelling, these domains have been outlined at a 0.15% Zn threshold. Three-D and plan representations of the high Zn domains are illustrated in Figure 17.20 and Figure 17.21, with high Zn areas shown in a deeper shade of blue. As can be seen, they are almost exclusively related to skarn mineralisation. The definition of the Skarn2 metallurgical type includes the condition of Zn>0.5%. At this stage of modelling, the only Zn definition available is from the wireframes.

Page 190

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.20 Supergene and Skarn, with High Zn Domains Highlighted

Page 191

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.21 Supergene and Skarn, with High Zn Domains Highlighted – Plan View

The next stage has been to develop a series of block models which can be used to identify skarn zones expected to contain supergene material, and where Zn is also high. A simple overlap of wireframes is not sufficient, since the aim is to try and predict time periods during which skarn will be mined with more than 5% of supergene. Proportions of material types within various block sizes have been generated from the wireframes. Definition of ‘Skarn2’ (SK2) type material can then be defined on a block basis, where the proportion of Skarn (SK) and Supergene (SG) are both greater than a selected threshold, and where the Zn grade is ‘high’. Figure 17.22 and Figure 17.23 show the distribution of normal resource model size blocks (25x25x15 m) where skarn occurs and there is greater than 5% of SG. Most occurs within high-Zn wireframes, and most of the material is also contained within several contiguous areas. This is encouraging in that it suggests meaningful spatial domains can be identified.

Page 192

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.22 SK2 in Resource Model Blocks (highlighted in red)

Figure 17.23 Plan View of SK2 in Resource Model Blocks (highlighted in red)

Page 193


To evaluate this approach using different overall volumes to simulate tonnage mined during various time periods, the same exercise has been carried out for block sizes of: 50x50x25 (approximately 160 000 t) 75x75x30 (approximately 450 000 t) 100x100x40 (approximately 1 000 000 t) Results are illustrated in visual form (Figure 17.24 for 25x25x10 m blocks and Figure 17.25 for 100x100x40 m blocks) and in tabular form (Table 17.1). Figure 17.24 SK2 Areas - 50x50x25 m Blocks

SK2 areas (including MX) 50x50x25

Plan of SK2 areas (including MX) 50x50x25

SK2 areas (excluding MX) 50x50x25

Plan of SK2 areas (excluding MX) 50x50x25

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.25 SK2 Areas - 100x100x40 m Blocks





The proportion of total Skarn which becomes SK2 using these calculations has been summarised in Table 17.1. The amount of Skarn classified as SK2 is shown, and compared to the total Skarn (not just mineralised skarn) of 149.9 Mt. Table 17.1 Quantities of SK2 Skarn in Model SK2 (Inc MX)

SK2 (Exc MX)

SK2 %

SK2 %

Mt

Mt

(Inc MX)

(Ex MX)

Wireframe Overlap

5.3

2.6

4%

2%

25x25x10

6.4

3.5

4%

2%

50x50x25

9.2

5.7

6%

4%

75x75x30

11.1

8.0

7%

5%

100x100x40

15.0

10.0

10%

7%

No account is taken in these calculations of Zn grade. Even though the majority of the SK2 material will be in ‘high’ Zn skarn areas, the grade may not always be above 0.5% Zn.

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It appears that using a 100x100x40 m re-block size and re-classifying Skarn as SK2 within these panels essentially wherever Skarn and Supergene (including Mixed) occur within them, gives two distinct contiguous zones of SK2, one at Constancia, the other, smaller one at San José. Both zones are limited in areal and depth extent, and would appear to give a good first pass approach to defining SK2 material for optimisation purposes. It appears that SK2 as it would be expected to present to the mill, accounts for some 10% of total Skarn, or some 10 to 15 Mt of mineralised SK2. 17.4.6

Topography

The topography DXF files for the area covering the Constancia project were imported into Micromine software as strings, and a 3-D surface digital terrain model was created. This is illustrated in Figure 17.26, Figure 17.27 and Figure 17.28. Figure 17.26 Topographic Contours and Features

Page 196

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.27 Topographic Surface Plan

Figure 17.28 Topographic Surface Model Looking Northwest

Page 197


17.5

SAMPLE CODING

17.5.1

Flagging by lithology domain

The assay interval data was flagged into a LITH field by overlaying the lithological domain solid models. The flagging process was validated by visual inspection on section, plan and in 3-D. 17.5.2

Flagging by mineralisation domain

The assay interval data was also flagged into a ZONE field by overlaying the mineralisation domain solid models. The flagging process was validated by visual inspection on section, plan and 3-D. 17.5.3

Flagging by copper and zinc grade shells

The assay interval data was flagged into a 02CU field by overlaying the 0.2% Cu grade shell solid model. The flagging process was validated by visual inspection on section, plan and 3-D. The assay interval data was also flagged into a ZNZONE field by overlaying the 0.15% Zn grade shell solid model. The flagging process was validated by visual inspection on section, plan and 3-D. 17.6

DATA COMPOSITING

The dominant sample length at Constancia is 2 m, but there are many smaller (and some larger) intervals. To provide valid data for statistical and geostatistical analysis, 2 m composites were generated; these composites honoured the boundaries in the assay data of the lithological and mineralisation domains. Histograms of length distribution in samples and composites are shown in Figure 17.29 and Figure 17.30.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.29 Histogram of Assay Interval Length

Figure 17.30 Histogram of Composite Interval Length

Page 199


17.7

STATISTICAL ANALYSIS AND VARIOGRAPHY

17.7.1

Statistical analysis by domain

An analysis of histograms and log probability plots of composites as coded by the various domains was carried out to determine which domains would be used as hard and soft boundaries during interpolation. A combination of the mineralisation domains and the 02CU grade shell, together with exceptions flagged by late stage dykes (as defined in the lithological domains), was found to provide satisfactory discrimination between the populations. A summary of the distribution of the main domains can be seen in Figure 17.31, Figure 17.32, Figure 17.33, Figure 17.34, Figure 17.35, Figure 17.36 and Figure 17.37. 17.7.2

Outlier analysis (capping)

Outlier analysis was carried for Cu, Mo, Ag and Zn using the following methodology: 

Review of histograms and probability plots to identify significant breaks in populations that may be used to interpret possible outliers.



Investigation of spatial clustering of potential outlier data. High grade composites that exhibit clustering may be considered valid members of the population, while isolated high grade composites were considered as possible outliers, requiring cutting and/or search restriction.



Analysis by combinations of domains as determined from domain statistical analysis (see Section 17.7.1)

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.31 Cu Log Probability Plot Overlay of Domains

No capping was applied

Page 201

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.32 Mo Log Probability Plot Overlay of Domains

Capping Oxide Cu

0.12% Mo

Supergene

0.12% Mo

OP (default) 0.10% Mo Hypogene

0.10% Mo

Skarn

0.16% Mo

Page 202

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.33 Ag Log Probability Plot Overlay of Domains

Capping Supergene

60 ppm Ag

Oxide Cu

40 ppm Ag

Skarn

70 ppm Ag

Hypogene

50 ppm Ag

OP

35 ppm Ag

Page 203

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.34 Au Log Probability Plot Overlay of Domains

Capping 1 gm/t for all domains

Page 204

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.35 Zn Log Probability Plot Overlay of Domains

No capping was applied.

Page 205

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.36 Pb Log Probability Plot Overlay of Domains

No capping was applied.

Page 206

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.37 S Log Probability Plot Overlay of Domains

No capping was applied. 17.7.3

Variography

Variography was approached by initially generating horizontal variogram contour fans, then generating variograms in the direction of best continuity and perpendicular to this direction. The known geological orientations of certain features and domains were taken into account when deciding upon primary orientations. Nugget variances were typically found to be best determined using down-hole variograms, with the nugget value then being applied to the corresponding directional models. Variogram modelling was generally achieved using two structure spherical schemes. Modelled variogram parameters are provided in Table 17.2. Note that the variograms are modelled using the normal experimental values and then the modelled gamma values are re-scaled for compatibility with population variances. Note, too, that variography on the oxide mineralisation domains was carried out in “unfolded” space as described in Section 17.9.3.

Page 207

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.2 Variogram Model Parameters HYPOGENE Variable Cu

Mo

Ag

Zn

Pb

S

Au

Axis

Orientation

Nugget

Sill 1

0.55

0.12

Range 1

Sill 2

Major

0° → 000°

Semi-major

0° → 090°

Minor

90° → 000°

145.0

345.0

Major

0° → 000°

46.0

175.0

Semi-major

0° → 090°

Minor

90° → 000°

Major

0° → 000°

Semi-major

0° → 090°

Minor

90° → 000°

85.0

187.0

Major

0° → 000°

269.0

544.0

Semi-major

0° → 090°

Minor

90° → 000°

269.0

544.0

Major

0° → 000°

50.0

187.0

Semi-major

0° → 090°

Minor

90° → 000°

Major

0° → 000°

Semi-major

0° → 090°

Minor

90° → 000°

165.0

360.0

Major

0° → 000°

145.0

264.0

Semi-major

0° → 090°

Minor

90° → 000°

145.0

0.36

0.10

145.0

46.0

345.0 0.33

0.54

46.0

0.46

0.63

0.10

0.06

0.19

85.0

269.0

50.0

0.48

0.18

0.35

0.18

165.0

145.0

187.0

544.0

187.0 187.0

165.0 0.01

175.0 187.0

0.28

50.0 0.37

345.0

175.0

85.0 0.61

Range 2

360.0 0.63

0.47

145.0

360.0

264.0 264.0

SUPERGENE Variable Cu

Mo

Ag

Zn

Axis

Orientation

Major

0° → 055°

Nugget

Sill 1

Range 1

Sill 2

30

Range 2 140

Semi-major

0° → 145°

Minor

90° → 000°

8

36

Major

0° → 000°

40

120

Semi-major

0° → 090°

Minor

90° → 000°

0.16

0.35

0.38

0.25

50

30

0.47

0.40

8

130

200 30

Major

0° → 000°

Semi-major

0° → 090°

Minor

90° → 000°

8

14

Major

0° → 000°

54

298

Semi-major

0° → 090°

30 0.41

0.44

0.12

0.04

45

80

88 0.47

0.52

105

140

Page 208

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.2 Variogram Model Parameters

Variable Cu

Mo

Ag

Zn

Variable Cu

Mo

Ag

Zn

Variable Cu

Mo

Ag

Zn

Axis Major Semi-major Minor Major Semi-major Minor Major Semi-major Minor Major Semi-major Minor

SKARN CONSTANCIA Orientation Nugget Sill 1 0° → 045° 25° → 135° 0.35 0.21 65° → 315° 0° → 045° 25° → 135° 0.13 0.32 65° → 315° 0° → 045° 15° → 135° 0.17 0.52 75° → 315° 0° → 045° 25° → 135° 0.15 0.18 65° → 315°

Range 1 150 195 13 100 164 14 155 206 6 47 71 17

Sill 2


SKARN SAN JOSÉ Orientation Nugget Sill 1 0° → 000° 50° → 270° 0.35 0.21 40° → 090° 0° → 000° 50° → 270° 0.13 0.32 40° → 090° 0° → 000° 50° → 270° 0.17 0.52 40° → 090° 0° → 000° 50° → 270° 0.15 0.18 40° → 090°

Range 1 150 195 13 100 164 14 155 206 6 47 71 17

Sill 2


Orientation 0° → 000° 0° → 090° 90° → 000° 0° → 020° 0° → 110° 90° → 000° 0° → 000° 0° → 090° 90° → 000° 0° → 000° 0° → 090° 90° → 000°

Range 1 50 30 6 170 145 10 60 100 3.5 25 36 15

Sill 2

OXIDE CU Nugget

Sill 1

0.26

0.45

0.09

0.28

0.07

0.12

0.07

0.13

0.44

0.55

0.31

0.67

0.44

0.55

0.31

0.67

0.28

0.62

0.81

0.80

Range 2 246 205 27 401 310 78 178 243 16 136 191 25

Range 2 246 205 27 401 310 78 178 243 16 136 191 25

Range 2 70 75 8 180 280 20 80 185 19 125 198 57

Page 209

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.2 Variogram Model Parameters

Variable


Cu

Mo

Ag

Zn

OXIDE LEACHED Orientation Nugget Sill 1 0° → 020° 0° → 110° 0.25 0.11 90° → 000° 0° → 000° 0° → 090° 0.16 0.08 90° → 000° 0° → 045° 0° → 135° 0.01 0.55 90° → 000° 0° → 000° 0° → 090° 0.08 0.13 90° → 000°

17.8

BLOCK MODEL CONSTRUCTION

17.8.1

Preparation

Range 1 75 77 5.7 45 50 5 95 95 5.5 145 130 13

Sill 2 0.64

0.76

0.44

0.79

Range 2 180 200 16.4 280 170 25 183 200 24 390 400 28

A fully coded cellular (block) model representing all mineralisation, lithological and grade shell subsets was constructed by applying constraints using relevant surface and solid wireframes, with reference to model prototype parameters shown in Table 17.3. The selected parent cell dimensions were based on observations of the typical spatial distribution of data in each of the easting, northing and vertical directions. Minimum sub-cell dimensions were selected so as to reflect the likely geometric resolution of boundary constraints imposed by the various interpretative elements as determined from the overall drillhole spacing. Table 17.3 Model Prototype Parameters Easting

Northing

RL

Model origin (centroid)

200 500

8 399 000

3607.5

Parent cell dimension (m)

25

25

15

Number of cells

100

80

60

Minimum subcell (m)

2.5

5

3

Page 210


17.8.2

Lithological modelling (LITH)

The lithological model was created by an ordered overlaying of sub-models created within the lithological domain solids. Models were sequentially added in the following order: 

MP1



SS



SK



MMP



MP2



QMP

Examples of the model are shown in section and plan in Figure 17.38 and Figure 17.39. Figure 17.38 Lithology Model – Plan View

Page 211

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.39 Lithology Model Northing Section

17.8.3

Mineralisation modelling (ZONE)

The mineralisation model was created by an ordered overlaying of sub-models created within the mineralisation domain solids. Models were sequentially added in the following order: 

Hypogene



Leached



Mixed



Oxide



Supergene

Examples of the model are shown below in section and plan in Figure 17.40 and Figure 17.41.

Page 212


Figure 17.40 Mineral Zonation Model Plan

Figure 17.41 Mineral Zonation Model Northing Section

Page 213


17.8.4

Cu grade shell modelling (02CU)

The 0.2% Cu grade shell solid was used to create a single model. 17.8.5

Zn grade shell modelling (ZNZONE)

The 0.15% Zn grade shell solid was used to create a single model. 17.8.6

Sub-model consolidation

The four sub-models were combined to form a single model, and as final step, this model was further overlain with a “rock” model generated beneath the topographic surface to produce a final volume model. 17.9

GRADE ESTIMATION

The block model was generated to include both the San José and Constancia zones. GRD Minproc conducted kriging neighbourhood analysis (KNA) on the domain-coded Cu data in the Constancia zone to determine optimal parameters for grade estimation. Goodness-of-fit statistics were generated to assess the efficiency of the various parameters. primary statistics used were the kriging efficiency (KE) and the slope of regression3.

The

KE calculates the overlap expected between the estimated block grade histogram and the ‘true’ block grade histogram. A high KE indicates a good match between estimated and ‘true’ grades, while as parameters become less optimal, KE drops. The slope of regression estimates the correlation between estimated and ‘true’ grades; a value approaching 1.0 indicates a good fit. In addition, other statistics, such as the percentage of negative weights generated in a kriging plan can be considered. A number of key input parameters can be tested in this way, including: 

Block size



Number of discretisation points



Search ellipse dimensions



Minimum and maximum numbers of samples in a search plan.

Cu, Mo, Ag, Pb, Au, S and Zn were estimated by Ordinary Kriging using individual variograms for each element and domain.

3

The kriging efficiency is calculated as (block variance-kriging variance)/block variance, where block variance is the total sill less the

variance contained within a block. The slope of regression is calculated as (block variance – kriging variance + µ)/ (block variance – kriging variance + 2µ).

Page 214


Inverse distance squared interpolation was also carried out for each element and domain, using the same search parameters as for the Ordinary Kriging estimate. This was carried out to aid in validation of the final resource model. 17.9.1

Domain control on estimation

The mineralisation domains are used as the primary domain control on interpolation; however, subdomains are also defined within the hypogene and skarn domains by the 0.2% Cu grade shell (inside and outside). A third layer of sub-domain occurs within the hypogene caused by the need to separate out the barren late-stage dykes defined by the lithology domains QMP and MP2. Zn% estimation also uses further sub-domains as defined by the 0.15% Zn grade shell. 17.9.2

Search strategy

For the mineralised domains, the preferred orientation of the search ellipsoids has been selected to be in the plane of the variograms. Note that the variogram directions are related to observed directions of continuity within each domain. Search distances are related essentially to the drillhole spacing in the plane of the mineralised domain, and to the nature of the grade continuity in the vertical or across-structure direction. For the hypogene and skarn domains the vertical component is a significant proportion of the horizontal components; for the flatter-lying supergene and various oxide domains, the vertical component is smaller (see also Section 17.9.3 on unfolding). The selected search parameters are summarised in Table 17.4.

Page 215

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.4 Search Parameters

ZONE

Search Distance

Rotation

Search 1

(m)

(deg)

(no. samples)

Volume

Search 2

Volume

(no. samples)

Search 3 (no. samples)

X

Y

Z

1

2

3

Min

Max

Factor

Min

Max

Factor

Min

Max

Sk Con (min)

75

75

25

45

25

0

12

30

2

12

30

4

2

30

Sk Con (unmin)

75

75

25

45

25

0

12

30

2

12

30

4

2

30

Sk SJ (min)

75

75

25

0

-50

0

12

30

2

12

30

4

2

30

Sk SJ (unmin)

75

75

25

0

-50

0

12

30

2

12

30

4

2

30

Sg

75

75

25

0

0

0

12

30

2

12

30

4

2

30

Hy (min)

75

75

50

0

0

0

12

30

2

12

30

4

2

30

Hy (LG)

45

75

50

0

0

0

12

30

2

12

30

4

2

30

Hy (MP2)

45

75

75

75

-75

0

12

30

2

12

30

4

2

30

Hy (QMP)

45

75

75

75

-75

0

12

30

2

12

30

4

2

30

Mx

75

75

25

0

0

0

12

30

2

12

30

4

2

30

Oxide

75

75

25

0

0

0

12

30

2

12

30

4

2

30

Leached

75

75

25

20

0

0

12

30

2

12

30

4

2

30

Op

75

75

45

0

0

0

12

30

2

12

30

4

2

30

17.9.3

Unfolding

The oxide and leached copper mineralisation domains are essentially undulating, sub-horizontal domains, which present issues with search ellipse orientations. They generally have limited vertical grade continuity and limited vertical thickness, but locally highly variable RL of the domain. Thus to construct a conventional search ellipse which captures sufficient sample numbers for satisfactory grade estimation often means using a vertical search component which can be much larger than the thickness of the domain, thereby causing over-smoothing in the vertical direction. Limiting the vertical search, conversely, will leave many cells unestimated. One simple solution to this issue is “unfolding” or “flattening” of the domain, where the data and block model are re-aligned to a plane representing the top of the domain. This has been performed for the supergene, oxide leach, mixed and oxide copper domains. A comparison of original and “unfolded” supergene blocks on a typical section is illustrated in Figure 17.42 and Figure 17.43. Note that the variography for these domains was carried out on composites in unfolded space.

Page 216

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.42 Data before Unfolding

Figure 17.43 Data in Unfolded Space

17.9.4

High grade capping

The high grade capping analysis is summarised in Section 17.7.2. 17.9.5

Other parameters

Other parameters used in the estimation were: 

Discretisation 3x3x3 (X,Y,Z)



Parent cell estimation



Maximum number of holes = 6.

17.10

DENSITY ASSIGNMENT

The 1108 density measurements taken by Norsemont have been flagged with the mineralisation and lithological models. Average density values were then determined for each of the modelled lithological units, taking mineralisation style into account, and applied to the block model.

Page 217


A summary of the average density values applied to the block model is presented in Table 17.5. Table 17.5 Density Assignment Mineralisation

17.11

Lithology

Density

Hypogene

MP1

2.55

Hypogene

MP2

2.63

Hypogene

MMP

2.62

Hypogene

QMP

2.48

Hypogene

Skarn (High Zn)

2.86

Hypogene

Skarn (Low Zn)

3.05

Supergene

Undifferentiated

2.38

Oxide

Undifferentiated

2.37

Leached

Undifferentiated

2.46

Mixed

Undifferentiated

2.39

Unmineralised

MP1

2.62

Unmineralised

MP2

2.63

Unmineralised

MMP

2.62

Unmineralised

QMP

2.48

Unmineralised

Skarn

2.77

Unmineralised

Sandstone

2.48

RESOURCE CLASSIFICATION

Procedures for classifying the resources were undertaken within the context of the Canadian Securities Administrators National Instrument 43-101 (NI 43-101), with consideration of the following criteria: 

Quality and reliability of raw data (sampling, assaying and surveying)



Dill hole spacing



Confidence in the geological interpretation



Number, spacing and orientation of intercepts through mineralised zones



Grade continuity information gained from observations and variography



Search volume and number of samples in search



Overall number of samples and spatial continuity of domain



Kriging variance.

GRD Minproc considers that there is sufficient drilling and sampling information, and that this information is of sufficient quality to classify the mineral resource in Measured, Indicated and Inferred categories.

Page 218


Criteria for classification are summarised in Table 17.6, and an overview of the classification into Measured, Indicated and Inferred categories is illustrated in Figure 17.44, Figure 17.45, Figure 17.46, Figure 17.47 and Figure 17.48. Table 17.6 Resource Classification Criteria Domain

Lithology

Cu Envelope

Conditions Search Pass 1 Samples=40 Kriging Variance< 0.1 RL > 4000m Search Pass 1 RL > 3780m

Class

Hypogene

MP1

Inside

Hypogene

MP1

Inside

Hypogene

MP1

Inside/Outside

Search Pass 2/3 Search Pass 1 Samples=40 Kriging Variance< 0.1 RL > 4000m Search Pass 1 RL > 3780m

Inferred

Hypogene

MMP

Inside

Hypogene

MMP

Inside

Hypogene

MMP

Inside/Outside

Inferred

Inside

Search Pass 2/3 Search Pass 1 Samples=30 Kriging Variance< 0.2 RL > 4000m

Hypogene

SKARN

Hypogene

SKARN

Inside

Search Pass 1

Indicated

Hypogene

SKARN

Inside/Outside

Search Pass 2/3

Inferred

Hypogene

QMP

Hypogene

MP2

Measured Indicated

Measured Indicated

Measured

Unclassified Unclassified

Supergene

Inside

Supergene

Inside

Mixed Mixed

Search Pass 1 Samples=30 Kriging Variance< 0.4

Measured

Inside

Search Pass 1 Search Pass 1 Samples=30 Kriging Variance< 0.3

Indicated

Measured

Inside

Search Pass 1

Indicated

Leached

Inferred

Oxide

Unclassified

Other

Unclassified

Page 219

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.44 Resource Classification Schematic Level 4300 mRL

Figure 17.45 Resource Classification Schematic Level 4200 mRL

Page 220

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.46 Resource Classification Schematic Level 4100 mRL

Figure 17.47 Resource Classification Schematic Level 4000 mRL

Page 221

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.48 Resource Classification Schematic - Transform Section

17.12

MODEL VALIDATION

Model validation was carried out in several ways: Statistics were generated to confirm that interpolated model cell grade field values fall within acceptable bounds. Results of comparison of composite means with model means are contained in Table 17.7 for Lithology and Table 17.9 for Zonation.

Page 222

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.7 Statistical Comparison between Means of Composite and Model Data – Lithology Model Composites

Model Cu

LITH MMP

0.14

0.16

MP1

0.23

0.19

MP2

0.04

0.07

QMP

0.04

0.05

SK

0.45

0.45

SS

0.02

0.03 Mo

LITH MMP

0.004

0.005

MP1

0.007

0.006

MP2

0.001

0.001

QMP

0.002

0.005

SK

0.008

0.008

SS

0.000

0.001 Ag

LITH MMP

2.10

2.41

MP1

2.58

2.38

MP2

0.69

1.05

QMP

0.81

1.95

SK

5.11

5.02

SS

0.79

0.97 Pb

LITH MMP

0.03

0.03

MP1

0.04

0.04

MP2

0.02

0.03

QMP

0.03

0.04

SK

0.07

0.07

SS

0.01

0.04 Zn

LITH MMP

0.05

0.05

MP1

0.06

0.07

MP2

0.04

0.06

QMP

0.06

0.07

SK

0.43

0.41

SS

0.03

0.06

Page 223

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.8 Statistical Comparison between Means of Composite and Model Data – Lithology Model Au

LITH MMP

0.02

0.03

MP1

0.04

0.03

MP2

0.01

0.02

QMP

0.01

0.03

SK

0.07

0.07

SS

0.02

0.02 S

LITH MMP

2.46

2.38

MP1

1.51

1.56

MP2

0.49

0.69

QMP

0.43

0.94

SK

4.77

4.53

SS

1.51

0.85

Table 17.9 Statistical Comparison between Means of Composite and Model Data – Mineral Zonation Model Composites

Model Cu

ZONE HY

0.21

0.19

LX

0.10

0.11

MX

0.50

0.46

OP

0.08

0.06

OX

0.59

0.60

SG

0.58

0.56 Mo

ZONE HY

0.007

0.005

LX

0.007

0.006

MX

0.013

0.010

OP

0.003

0.002

OX

0.004

0.005

SG

0.008

0.008 Ag

ZONE HY

2.31

2.34

LX

3.62

3.35

MX

3.67

3.23

OP

1.53

1.21

OX

3.45

3.78

SG

5.01

4.63

Page 224

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.9 (cont) Statistical Comparison between Means of Composite and Model Data – Mineral Zonation Model Pb

ZONE HY

0.04

0.04

LX

0.05

0.05

MX

0.04

0.04

OP

0.06

0.04

OX

0.05

0.06

SG

0.05

0.05 Zn

ZONE HY

0.10

0.13

LX

0.06

0.06

MX

0.09

0.09

OP

0.11

0.06

OX

0.12

0.14

SG

0.05

0.06 Au

ZONE HY

0.03

0.03

LX

0.04

0.04

MX

0.04

0.05

OP

0.02

0.02

OX

0.08

0.09

SG

0.05

0.05 S

ZONE HY

1.96

2.23

LX

0.31

0.32

MX

1.94

2.15

OP

0.71

0.98

OX

0.23

0.34

SG

1.98

2.02



The grade estimates in the block model were thoroughly scrutinised using graphical visualisation utilities. Model and drillhole data were overlain and viewed in various 2D section and plan view slices, with colour highlighting of grade or zonal attributes. Examples are shown in Figure 17.49 Figure 17.50 and Figure 17.51.



Model grade spatial distribution patterns were also reviewed using 3-D facilities, presented variously as section planes, point clouds and cell faces.



A comparison was carried out with the previous resource estimate (GRD Minproc, 2008). It was found that grades in the current model are generally slightly higher than in the previous estimate. Tonnage has increased approximately 10%.

Page 225




A comparison was carried out against declustered block grades.



A comparison of block grades estimated using Inverse Distance Squared was carried out. Figure 17.49 Cu Grade Model Validation - Section 202050 East

Page 226

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.50 Cu Grade Model Validation - Transform Section Cu Validation

Page 227

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.51 Cu Grade Model Validation - Level Plan 4225 mRL

17.13

MINERAL RESOURCE REPORTING

Computations of global tonnes and grade estimates (Table 17.10 and Table 17.11) were checked and verified using two independent software packages (Micromine and Datamine). The copper cut-off grade of 0.25% corresponds to that currently applied by Norsemont and does not represent any independent assessment by GRD Minproc of an economic cut-off. The resource is also estimated at a variety of cut-off grades, and by area and domain. It is important to note the following when considering the grade and tonnage estimates: 

Mineral Resources that are not Ore Reserves do not have demonstrated economic viability.



Measured and Indicated Mineral Resources are that part of a Mineral Resource for which quantity and grade can be estimated with a level of confidence sufficient to allow the application of technical and economic parameters to support mine planning and evaluation of the economic viability of the deposit.



An Inferred Mineral Resource is that part of a Mineral Resource for which quantity and grade can be estimated on the basis of geological evidence and limited sampling and reasonably assumed, but not verified continuity.

Page 228




Although 0.25% Cu is considered a potential cut-off grade for this deposit, the actual cut-off grade has not been confirmed by the appropriate economic studies.



The Constancia deposit includes the Constancia and San José zones. Table 17.10 Constancia Project Global Mineral Resource Estimate 0.25% Cu Cut-off Category

Cut off

Tonnes M)

Cu%

Mo%

Zn%

Ag ppm

Pb%

Au ppm

S%

MEASURED

0.25

119.00

0.47

0.014

0.083

3.73

0.039

0.047

2.22

INDICATED

0.25

195.00

0.48

0.010

0.159

4.17

0.047

0.058

2.62

MEAS+IND

0.25

315.00

0.47

0.012

0.130

4.00

0.044

0.054

2.47

INFERRED

0.25

38.00

0.47

0.009

0.212

4.91

0.057

0.072

3.26

Table 17.11 Constancia Project Global Mineral Resource Estimate 0.20% Cu Cut-off Category

Cut off

Tonnes (M)

Cu%

Mo%

Zn%

Ag ppm

Pb%

Au ppm

S%

MEASURED

0.20

138.00

0.44

0.013

0.080

3.54

0.037

0.044

2.15

INDICATED

0.20

254.00

0.42

0.010

0.146

3.81

0.044

0.053

2.43

MEAS+IND

0.20

392.00

0.42

0.011

0.123

3.72

0.042

0.050

2.33

INFERRED

0.20

62.00

0.37

0.008

0.158

4.09

0.048

0.061

2.55

Additional reporting has been completed for a range of cut-off grades, by orebody (Constancia and San José) and by domain (Table 17.12, Table 17.13, Table 17.4, Table 17.5, Table 17.6 and Table 17.17).

Page 229


Table 17.12

Category MEASURED

INDICATED

MEAS+IND

INFERRED

Constancia Project Global Mineral Resource Estimate at Various Cu Cut-off Tonnes Au (M) gm/t Cut off Cu% Mo% Zn% Ag gm/t Pb%

S%

0.50

34.2

0.77

0.015

0.12

5.09

0.04

0.07

2.85

0.40

56.2

0.64

0.015

0.11

4.58

0.04

0.06

2.60

0.35

72.5

0.58

0.015

0.10

4.30

0.04

0.06

2.49

0.30

95.1

0.52

0.014

0.09

3.99

0.04

0.05

2.35

0.25

119.2

0.47

0.014

0.08

3.73

0.04

0.05

2.22

0.20

138.3

0.44

0.013

0.08

3.54

0.04

0.04

2.15

0.15

146.8

0.42

0.013

0.08

3.46

0.04

0.04

2.12

0.50

55.5

0.79

0.012

0.26

6.46

0.07

0.09

3.98

0.40

90.2

0.66

0.012

0.22

5.46

0.06

0.08

3.38

0.35

120.0

0.59

0.011

0.19

4.93

0.05

0.07

3.04

0.30

152.6

0.53

0.011

0.18

4.54

0.05

0.06

2.83

0.25

195.3

0.48

0.010

0.16

4.17

0.05

0.06

2.62

0.20

254.2

0.42

0.010

0.15

3.81

0.04

0.05

2.43

0.15

376.2

0.34

0.008

0.13

3.24

0.04

0.05

2.15

0.50

89.7

0.79

0.013

0.21

5.94

0.06

0.08

3.55

0.40

146.4

0.65

0.013

0.17

5.12

0.05

0.07

3.08

0.35

192.5

0.59

0.013

0.16

4.69

0.05

0.06

2.83

0.30

247.7

0.53

0.012

0.14

4.33

0.05

0.06

2.64

0.25

314.5

0.47

0.012

0.13

4.00

0.04

0.05

2.47

0.20

392.5

0.42

0.011

0.12

3.72

0.04

0.05

2.33

0.15

523.0

0.36

0.009

0.11

3.30

0.04

0.04

2.14

0.50

7.5

0.75

0.011

0.32

7.54

0.08

0.09

6.06

0.40

11.7

0.64

0.010

0.30

6.53

0.07

0.08

5.58

0.35

15.5

0.58

0.010

0.27

5.88

0.07

0.08

5.12

0.30

20.8

0.51

0.009

0.25

5.27

0.06

0.07

4.64

0.25

28.5

0.45

0.009

0.22

4.75

0.06

0.07

4.11

0.20

48.8

0.35

0.008

0.16

3.82

0.04

0.06

3.09

0.15

144.6

0.23

0.006

0.09

2.53

0.03

0.04

1.98

Page 230

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 17.52 Constancia Resource Reporting by Confidence Classification

Table 17.13 Constancia Project Global Mineral Resource Estimate – Constancia Main Orebody Category

Cut off

Tonnes (M)

Cu%

Mo%

Zn%

Ag gm/t

Pb%

Au gm/t

S%

MEASURED

0.25

115.2

0.47

0.014

0.08

3.71

0.04

0.05

2.22

INDICATED

0.25

185.4

0.47

0.011

0.16

4.09

0.05

0.06

2.59

MEAS+IND

0.25

300.6

0.47

0.012

0.13

3.94

0.04

0.05

2.45

INFERRED

0.25

26.0

0.45

0.009

0.22

4.63

0.05

0.07

4.00

MEASURED

0.20

133.9

0.43

0.013

0.08

3.52

0.04

0.04

2.15

INDICATED

0.20

238.7

0.42

0.010

0.14

3.75

0.04

0.05

2.40

MEAS+IND

0.20

372.6

0.42

0.011

0.12

3.67

0.04

0.05

2.31

INFERRED

0.20

42.7

0.36

0.008

0.16

3.81

0.05

0.06

3.07

MEASURED

0.15

142.2

0.42

0.013

0.08

3.44

0.04

0.04

2.12

INDICATED

0.15

351.8

0.34

0.008

0.12

3.19

0.04

0.04

2.10

MEAS+IND

0.15

494.0

0.36

0.009

0.11

3.26

0.04

0.04

2.11

INFERRED

0.15

132.0

0.23

0.006

0.08

2.49

0.03

0.04

1.91

Page 231


Table 17.14 Constancia Project Global Mineral Resource Estimate – San José Orebody Category

Cut off

Tonnes (M)

Cu%

Mo%

Zn%

Ag gm/t

Pb%

Au gm/t

S%

MEASURED

0.25

4.0

0.66

0.009

0.08

4.25

0.03

0.10

2.29

INDICATED

0.54

0.007

0.18

5.77

0.05

0.09

3.30

0.25

9.9

MEAS+IND

0.25

13.9

0.58

0.008

0.15

5.33

0.04

0.09

3.01

INFERRED

0.25

2.4

0.45

0.006

0.22

5.96

0.06

0.05

5.35

MEASURED

0.20

4.3

0.63

0.009

0.08

4.09

0.03

0.10

2.29

INDICATED

0.20

15.4

0.43

0.008

0.21

4.78

0.05

0.08

2.92

MEAS+IND

0.20

19.7

0.47

0.008

0.18

4.63

0.04

0.08

2.78

INFERRED

0.20

6.1

0.31

0.006

0.14

3.88

0.04

0.05

3.22

MEASURED

0.15

4.5

0.61

0.009

0.07

3.98

0.03

0.09

2.28

INDICATED

0.15

24.4

0.34

0.007

0.24

3.90

0.04

0.07

2.77

MEAS+IND

0.15

28.9

0.38

0.007

0.21

3.91

0.04

0.07

2.69

INFERRED

0.15

12.6

0.24

0.006

0.17

2.90

0.03

0.05

2.70

Page 232


Table 17.15 Constancia Project Global Mineral Resource Estimate – Domain MP1 Hypogene MP1 Category MEASURED

INDICATED

MEAS+IND

INFERRED

Cut off

Tonnes (M)

Cu%

Mo%

Zn%

Ag gm/t

Pb%

Au gm/t

S%

0.50

11.6

0.65

0.017

0.07

4.36

0.03

0.07

2.22

0.40

28.4

0.53

0.017

0.06

3.94

0.04

0.06

2.02

0.35

42.0

0.48

0.016

0.06

3.69

0.04

0.05

1.94

0.30

62.4

0.43

0.015

0.06

3.43

0.04

0.05

1.86

0.25

84.9

0.39

0.014

0.06

3.21

0.04

0.04

1.80

0.20

102.4

0.36

0.013

0.05

3.04

0.04

0.04

1.77

0.15

110.2

0.35

0.013

0.05

2.96

0.04

0.04

1.77

0.50

10.8

0.63

0.015

0.09

4.57

0.04

0.08

2.85

0.40

30.5

0.51

0.013

0.07

3.80

0.04

0.07

2.07

0.35

51.3

0.45

0.013

0.06

3.46

0.03

0.06

1.82

0.30

75.0

0.41

0.012

0.06

3.26

0.03

0.05

1.73

0.25

109.0

0.37

0.011

0.05

3.08

0.03

0.05

1.69

0.20

160.4

0.32

0.010

0.05

2.88

0.03

0.05

1.65

0.15

272.7

0.26

0.008

0.05

2.49

0.03

0.04

1.56

0.50

22.4

0.64

0.016

0.08

4.46

0.04

0.07

2.52

0.40

58.9

0.52

0.015

0.06

3.86

0.04

0.06

2.05

0.35

93.3

0.46

0.014

0.06

3.57

0.04

0.06

1.87

0.30

137.4

0.42

0.013

0.06

3.34

0.04

0.05

1.79

0.25

193.9

0.38

0.012

0.05

3.14

0.03

0.05

1.74

0.20

262.8

0.34

0.011

0.05

2.94

0.03

0.04

1.70

0.15

382.9

0.29

0.009

0.05

2.63

0.03

0.04

1.62

0.50

1.6

0.58

0.018

0.16

4.35

0.04

0.07

5.20

0.40

3.2

0.51

0.017

0.14

4.07

0.04

0.07

4.77

0.35

4.9

0.46

0.015

0.11

3.71

0.04

0.06

4.03

0.30

7.7

0.41

0.014

0.09

3.33

0.03

0.06

3.40

0.25

12.9

0.36

0.012

0.07

3.24

0.03

0.06

2.94

0.20

30.9

0.28

0.008

0.04

2.75

0.02

0.05

2.03

0.15

123.8

0.20

0.006

0.04

2.06

0.03

0.03

1.51

Page 233

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.16 Constancia Project Global Mineral Resource Estimate – Domain Skarn Skarn Category MEASURED

INDICATED

MEAS+IND

INFERRED

Cut off

Tonnes (M)

Cu%

Mo%

Zn%

Ag gm/t

Pb%

Au gm/t

S%

0.50

5.1

0.91

0.02

0.49

7.36

0.05

0.12

7.63

0.40

6.2

0.83

0.02

0.49

7.00

0.05

0.11

7.67

0.35

7.0

0.78

0.02

0.48

6.73

0.05

0.11

7.72

0.30

7.6

0.74

0.01

0.48

6.53

0.05

0.10

7.66

0.25

8.0

0.72

0.01

0.48

6.40

0.05

0.10

7.61

0.20

8.3

0.70

0.01

0.50

6.29

0.05

0.10

7.58

0.15

8.4

0.70

0.01

0.50

6.26

0.05

0.10

7.56

0.50

21.5

0.88

0.01

0.56

8.92

0.10

0.11

6.67

0.40

29.4

0.77

0.01

0.53

7.85

0.09

0.10

6.17

0.35

34.3

0.71

0.01

0.52

7.36

0.09

0.10

5.94

0.30

39.2

0.66

0.01

0.51

6.98

0.09

0.09

5.75

0.25

44.0

0.62

0.01

0.51

6.64

0.09

0.09

5.54

0.20

48.7

0.58

0.01

0.52

6.36

0.09

0.08

5.41

0.15

55.5

0.53

0.01

0.54

5.99

0.08

0.08

5.19

0.50

26.7

0.89

0.01

0.55

8.62

0.09

0.11

6.86

0.40

35.6

0.78

0.01

0.52

7.70

0.09

0.10

6.43

0.35

41.3

0.72

0.01

0.51

7.25

0.08

0.10

6.24

0.30

46.9

0.67

0.01

0.50

6.90

0.08

0.09

6.06

0.25

52.0

0.63

0.01

0.50

6.60

0.08

0.09

5.86

0.20

57.0

0.60

0.01

0.52

6.35

0.08

0.09

5.73

0.15

63.9

0.55

0.01

0.53

6.02

0.08

0.08

5.50

0.50

5.9

0.80

0.01

0.36

8.40

0.09

0.09

6.29

0.40

8.5

0.69

0.01

0.36

7.45

0.08

0.08

5.88

0.35

10.6

0.63

0.01

0.34

6.89

0.08

0.08

5.63

0.30

13.1

0.57

0.01

0.34

6.40

0.08

0.08

5.36

0.25

15.6

0.52

0.01

0.34

6.00

0.08

0.07

5.08

0.20

18.0

0.48

0.01

0.36

5.66

0.08

0.07

4.91

0.15

20.8

0.44

0.01

0.40

5.33

0.08

0.07

4.77

Page 234

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 17.17 Constancia Project Global Mineral Resource Estimate – Domain Supergene Supergene Category MEASURED

INDICATED

MEAS+IND

Cut off

Tonnes (M)

Cu%

Mo%

Zn%

Ag gm/t

Pb%

Au gm/t

S%

0.50

17.5

0.81

0.014

0.054

4.92

0.037

0.052

1.87

0.40

21.6

0.74

0.013

0.055

4.74

0.038

0.048

1.91

0.35

23.5

0.71

0.012

0.055

4.67

0.039

0.047

1.92

0.30

25.1

0.69

0.012

0.054

4.63

0.039

0.045

1.94

0.25

26.4

0.67

0.012

0.054

4.60

0.039

0.044

1.94

0.20

27.6

0.65

0.011

0.053

4.58

0.039

0.044

1.94

0.15

28.2

0.64

0.011

0.053

4.59

0.039

0.043

1.94

0.50

23.2

0.79

0.012

0.068

5.06

0.044

0.068

2.02

0.40

30.3

0.71

0.012

0.068

4.82

0.046

0.062

1.99

0.35

34.4

0.67

0.011

0.067

4.70

0.046

0.059

1.99

0.30

38.4

0.64

0.011

0.067

4.56

0.047

0.056

2.00

0.25

42.4

0.60

0.010

0.067

4.42

0.046

0.053

2.00

0.20

45.1

0.58

0.010

0.068

4.38

0.048

0.052

1.98

0.15

47.9

0.56

0.010

0.068

4.31

0.048

0.050

1.98

0.50

40.7

0.80

0.013

0.062

5.00

0.041

0.061

1.95

0.40

51.9

0.72

0.012

0.063

4.79

0.043

0.056

1.96

0.35

57.8

0.69

0.011

0.062

4.68

0.043

0.054

1.96

0.30

63.5

0.66

0.011

0.062

4.58

0.044

0.052

1.97

0.25

68.8

0.63

0.011

0.062

4.49

0.043

0.050

1.98

0.20

72.7

0.60

0.010

0.062

4.45

0.045

0.049

1.97

0.15

76.2

0.59

0.010

0.062

4.41

0.045

0.047

1.96

Page 235


17.14

MINERAL RESERVES

The Constancia Mineral Reserve is the Measured and Indicated Resource contained in the open pit mine that can be processed at a profit and is scheduled for treatment in the DFS Life-of-Mine (LOM) plan. Since revenue is derived from four payable components (copper, molybdenum, silver plus minor payable gold) the Mineral Reserve reporting cut-off is based on a Net Smelter Return (NSR) cut-off that is estimated using the metal prices and other treatment, recovery and concentrate realisation parameters as detailed in Section 18.15. The Mineral Reserve estimate, comprising Proven and Probable categories, is summarised in Table 17.18. Table 17.18 Constancia Project Global Mineral Reserve Estimate Category

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

PROVEN

161.8

0.45

0.012

3.68

0.05

PROBABLE

115.6

0.40

0.011

3.70

0.05

TOTAL

277.4

0.43

0.012

3.69

0.05

Further reporting of Mineral Reserves is included in Section 18.1.

Page 236


18.

OTHER RELEVANT DATA AND INFORMATION

18.1

MINING STUDIES

18.1.1

Introduction

The Constancia copper porphyry deposit is located in high altitude (4000-4500 masl), with undulating to steep surface topography. Rock strengths are moderate in more weathered supergene and skarn mineralisation to hard in the hypogene ore type. There is some ground-water, and significant rainfall occurs during the wet season from October to March. The copper porphyry deposit is massive which makes it amenable to low cost open pit bulk mining techniques, using shovels mining 15 m benches. Some difficulties result from the relatively low grade and higher cost of mining and processing the hard hypogene ore, and the challenges in removing zinc impurity from the copper concentrates produced from a blend of supergene and skarn ore. Pit optimisation was carried out to maximise the Net Smelter Return (NSR) of the deposit whilst following the geotechnical slope constraints. The resulting mine consists of two adjoining pits: the major Constancia pit and the much smaller San José pit. The Constancia deposit is designed to be mined in four pit stages and San José in a single stage. The cash positive ore blocks were categorised as high, medium, or low NSR blocks based on their profit margins. A number of mining and processing schedule iterations were run. The schedule option which yielded the highest Net Present Value (NPV) was selected. This schedule calls for preferential processing of the high and medium profit margin ore blocks (raised cut-off), supplemented by low margin material when there is a shortfall of high and medium margin ore. Ore mined is hauled and fed directly to the crusher, milled and beneficiated at the processing plant. There is limited surge stockpile capacity at the ROM pad with minimum rehandle envisaged. The bulk of the low marginal material will be carted and stockpiled as part of the long-term potentially acid generating (PAG) waste rock facility (WRF). Low margin material is fed to the plant only during the ramp-up period (the first three months of processing) when metal recoveries are expected to be low, and when there is insufficient high and medium margin ore available to feed the plant, such as during the transition period between pit stages and towards the end of the mine life. Initial waste mined will be non-PAG and will be used to construct haul roads, back-fill pioneering ramps, and build the tailings management facility (TMF) embankment dam. Suitable initial waste mined may also be used to backfill civil construction sites in the mine. PAG waste rock mined will be dumped in the PAG WRF located to the south of the pit. To meet throughput and bulk mining requirements, mining uses 220 t class haul trucks, 32 m3 electric shovels and diesel-powered support equipment. Due to the high altitude, the use of electric mining equipment appears to be advantageous, as performance is not greatly affected by low levels of oxygen. This, however, needs to be weighed against the increased capital cost for such equipment.

Page 237


Total mining of approximately 45 Mt/a includes concentrator feed and waste rock. The average waste to ore ratio is 0.92:1 (based on inclusion of all low margin ore). Mining operations continue over 15 years, inclusive of an approximate eight month pre-production period. Mine planning activities performed during the course of the DFS included: 

Pit optimisation



Ultimate and staged pit design



Mineral Reserve estimation



Mine and process scheduling



Mine fleet assessment



Estimation of mine operating and capital costs.

18.1.2

Pit optimisation

The resource model prepared by GRD Minproc was based on a 25x25x15 m parent block size with minimum sub-cell size of 2.5x5x3 m. Two regularised mining models were prepared (25x25x10 m and 25x25x15 m) to simulate the impact on dilution and mining losses relative to the in-situ resource model. A comparison between the regularised models shows insignificant difference in grade and metal content. The 15 m model was selected as the basis for mine planning due to the higher productivity associated with operating on 15 m benches. Pit optimisation of the mining model (Measured and Indicated mineralisation only) was carried out using Whittle Four-X software on the NSR of each mining block. Revenue is received from copper and molybdenum concentrates with payable silver and gold in the copper concentrate. Optimisation input parameters were based on then-current information. Overall slope input (varying from 45° to 50° depending on domain – see Figure 18.1) was based on April 2009 advice from Knight Piésold. Metal prices were supplied by Norsemont with a copper price of $1.80/lb, molybdenum price of $12.00/lb, silver price of $11.00/oz and gold price of $750/oz. Processing costs and metal recoveries were provided by GRD Minproc, and mining costs were from the PFS mining cost modelling. The full list of optimisation parameters is shown in Table 18.1 to Table 18.4. A number of different scenarios and sensitivities were investigated to select pit optimisation shells. Ultimate and staged pit designs were developed from the selected optimisation shells, incorporating access ramps.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.1 Pit Optimisation Parameters - NSR Calculation Parameter

Unit

Estimate

Copper Price

$/lb

Molybdenum Price

$/lb

Silver Price

$/oz

1.80 12.00 11.00 750.00

Revenue Metal Price

Gold Price Payable Contained Metal Payable Contained Copper

$/oz

Payable Contained Moly

%

Payable Contained Silver

%

Payable Contained Gold Deductions Copper Deduction

%

% Per unit of payable Cu

Moly Discount

% Mo in Moly Con

Minimum Silver Deduction in Cu Concentrate

Ag g/dmt Cu conc

Gold Deduction Payable Metal in Concentrate Payable Copper in Concentrate

Au g/dmt Cu conc

Payable Moly in Concentrate

% Mo in Moly conc

Payable Silver in Copper Concentrate Payable Gold in Copper Concentrate Selling Costs Marketing Costs - Concentrates Transport - Cu concentrates - truck to port

% Cu in Cu conc g payable Ag/dmt Cu conc g payable Au/dmt Cu conc $/t conc $/wmt conc

Transport - Mo concentrates - truck to port

$/wmt conc

Port charges concentrate

$/wmt conc

Insurance

$/wmt conc

Shipping - Cu concentrates

$/wmt conc

Shipping - Mo concentrates Smelting Charges Smelting charges - copper concentrate (dry) Roasting charges - moly concentrate (dry) Price Participation Copper Range Cu Trigger Price Cap and Floor Applicable PP Refining Charges Refining charges - Payable Cu

$/wmt conc $/dmt Cu conc $/dmt Mo conc

% $/lb Cu $/lb Cu $/dmt Cu conc $/lb Cu

Refining charges - Payable Mo

$/lb Mo

Refining charges - Payable Ag

$/oz Ag

Refining charges - Payable Au Royalties State Royalty Minera Livitaca and Katanga

$/oz Au % of NSR % of NSR % of NSR (capped at $10M)

96.5% 100% 90% 98.0% 1.00 0% 30.00 1.00 30.49% 40.00% 209.5

0.00 32.30 77.57 5.86 1.78 35.00 0.00 65.00 1784.0

0% 1.20 0.05 0.00 0.07 0.00 0.40 1.20 3.5% 3.0% 0.5%

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.2 Optimisation Parameters - Operating Costs and Throughput Rates Parameter Operating Costs Processing Hypogene

Unit

Estimate

$/t

4.42 3.29 2.59 2.87

Supergene

$/t

Skarn 1

$/t

Skarn 2 Incremental Processing Cost by Element Copper

$/t

Moly

$/lb Mo

Silver

$/oz Ag

Gold

$/oz Au

0.00 0.00 0.00 0.00

$M/year

9.50

$/t concentrator feed

General & Administration Annual G & A Cost G & A Unit Costs Hypogene

$/lb Cu

Supergene


Skarn 1


Skarn 2


0.60 0.40 0.42 0.42


0.000

$/t mat'l mined

Variable 0.033

Owner's Costs (off-site but project specific) Mining Cost - Open Pit Mining cost ore/waste Incremental mining cost per 10m bench

$/t/bench

Mining Recovery

%

Mining Dilution

%

Scheduling Parameters Grinding Throughput Rate by Ore Type Hypogene (1624 t/h)

100% 0%

Supergene (2703 t/h)

mtpa

Skarn 1 (2880 t/h)

mtpa

Skarn 2 (2880 t/h)

mtpa

16 24 23 23

%

7%

Discount Rate

mtpa

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.3 Pit Optimisation Parameters - Concentrator Recoveries Parameter Recoveries into Concentrator Hypogene Ore Concentrator Recovery Copper recovery into Copper concentrate

Unit

Estimate

%

Moly recovery into Moly concentrate

%

Silver recovery into Copper concentrate

%

Gold recovery into Copper concentrate

%

Zinc recovery into Copper concentrate

%

Lead recovery into Copper concentrate Supergene Ore Concentrator Recovery Copper recovery into Copper concentrate

%

89.0% As per formula 80.0% 60.0% 50.0% 25.0%

%


%


%


%


%

Lead recovery into Copper concentrate Skarn1 Ore Concentrator Recovery Copper recovery into Copper concentrate

% %


%


%


%


%

Lead recovery into Copper concentrate Skarn2 Ore Concentrator Recovery Copper recovery into Copper concentrate

% %


%


%


%


%

Lead recovery into Copper concentrate

%

89.0% As per formula 80.0% 60.0% 50.0% 25.0% 85.0% As per formula 80.0% 60.0% 20.0% 10.0% 85.0% As per formula 80.0% 60.0% 25.0% 12.5%

Page 241

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.4 Pit Optimisation Parameters - Concentrate Grades Parameter Concentrate Grades Hypogene Ore Concentrates Copper concentrate grade Copper Molybdenum

Unit

Estimate

% Cu

As per formula n/a As per formula As per formula As per formula As per formula

% Mo

Silver

g Ag/dmt

Gold

g Au/dmt

Zinc Lead Moly concentrate grade Copper Molybdenum

% Zn % Pb % Cu % Mo

Silver

g Ag/dmt

Gold

g Au/dmt

Zinc Lead Supergene Ore Concentrates Copper concentrate grade Copper Molybdenum

% Zn % Pb

% Cu % Mo

Silver

g Ag/dmt

Gold

g Au/dmt


% Zn % Pb % Cu % Mo

Silver

g Ag/dmt

Gold

g Au/dmt

Zinc Lead Skarn1 Ore Concentrates Copper concentrate grade Copper Molybdenum

% Zn % Pb

% Cu % Mo

Silver

g Ag/dmt

Gold

g Au/dmt


% Zn % Pb % Cu % Mo

Silver

g Ag/dmt

Gold

g Au/dmt

Zinc

% Zn

Lead

% Pb

n/a 40.0% n/a n/a n/a n/a

As per formula n/a As per formula As per formula As per formula As per formula n/a 40.0% n/a n/a n/a n/a

As per formula n/a As per formula As per formula As per formula As per formula n/a 40.0% n/a n/a n/a n/a

Page 242

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.4 (cont) Pit Optimisation Parameters - Concentrate Grades Parameter Skarn2 Ore Concentrates Copper concentrate grade Copper Molybdenum

Unit

Estimate

% Cu

As per formula n/a As per formula As per formula As per formula As per formula

% Mo

Silver

g Ag/dmt

Gold

g Au/dmt

Zinc

% Zn

Lead Moly concentrate grade Copper Molybdenum

% Pb % Cu % Mo

Silver

g Ag/dmt

Gold

g Au/dmt

Zinc

% Zn

Lead

% Pb

Concentrate Moisture Contents Copper Concentrate Moisture Content

%

Moly Concentrate Moisture Content

%

Penalties

n/a 40.0% n/a n/a n/a n/a

7.5% 10.0% No penalties applied

Whittle Four-X software was used to generate optimal pits based on analysis of the resource model. Whittle Four-X allows the generation of a series of nested optimal pits where each successive outline is for a slightly higher product price than the previous one. This is done for a range of prices, from the lowest for which ore can be profitably mined to the highest expected in the future. These pits are then interrogated at the base case costs and prices to establish their respective values. The selection of the optimal pit is based on maximizing the project’s NPV. This selection criterion incorporates time-discounting of money and assumes two extreme mining sequences (best and worst cases) for optimal pit selection. The best-case mining sequence simulates the mining of the nested pits, starting with the smallest pit outline and mining subsequent pits until the largest pit is mined out. The worst-case mining sequence mines to the final pit outline bench by bench. The average of the best and worst possible NPV for each pit shell is calculated, and the pit shell with the maximum average NPV is selected as the optimal pit shell. A number of different scenarios and sensitivities were produced and shells were selected to form the basis for the ultimate and staged pit designs. 18.1.3

Pit design

Ultimate and staged pit designs were created by GRD Minproc from selected optimisation shells incorporating access ramps, catch berms, and internal haul roads. Potential pit design stages are identified by analysing the incremental changes between the series of nested pit shells leading to the selected ultimate pit shell. The incremental analysis of the pit optimisation results is done both

Page 243


numerically and graphically. Significant increases in cash flow and tonnage of ore and waste between succeeding pit shells is also analysed graphically to determine if the pit shell with the notable change lends itself as a basis for designing the next pit stage. The pit shell representing the next pit stage must satisfy the minimum mining width of 50 m and that it can be rationally integrated with the previous and succeeding pit stages. As many pit stages as possible that satisfies the minimum mining width are selected since the higher the number of pit stages, the better the resulting NPV of the project. Four pit stages were identified in the Constancia area while only a single stage was selected in San José due to its small size. The smallest practicable pit shell that can provide two years supply of ore with the highest profit margin and the lowest strip ratio was selected as the basis for the stage 1 or starter pit design. Subsequent pit stages with diminishing profit margins need to satisfy the minimum mining width of 50 m from the previous pit stage. There are instances when pit shells merge to form common walls. Where there is a common wall between pit stages, the accessibility of either side of the merged walls are considered in the design. Where accessibility becomes an issue, the walls of either the current or precursor pit are adjusted allow a minimum mining width of 50 m between the two stages. The pit design parameters were determined in conjunction with what were then current geotechnical slope recommendations from Knight Piésold as shown in Figure 18.1. Ten geotechnical sectors were established by Knight Piésold based on geometric, geological, and rock mass quality characteristics. The geotechnical sectors were then grouped by Knight Piésold into four geotechnical design regions based on the potential mode of failure (i.e. circular and or wedge/planar failures). Final review and analysis of the pit designs were carried out by Knight Piésold according to the pit wall locations defined by the optimisation study, using the specific geotechnical conditions assigned to those pit wall locations. In two of the ten design sectors, these checks showed that catch bench performance and global safety factors would be slightly lower than had been calculated according to the previous, more general geologic interpretations. After further review and assessment, and due to timing constraints, design changes were deferred until the next stage of the study. The revised recommendations are summarised in Section 18.2.

Page 244

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.1 Design Sectors and Proposed Inter-ramp and Bench Slopes

Page 245


In recognition of the increased risk associated with this approach the following mining operational considerations and cost allowances have been taken into account in the study: 

Controlled blasting of the interim slopes and presplitting of the final pit perimeter to physically decouple production blasting from the final wall and protect the wall rock structures and joints



Scaling down of bench faces with a small excavator to remove all loose rock prior to the bench face becoming inaccessible other than via berms



Maintaining access to, and regular clean-up of catch berms using small mobile equipment such as a tracked dozer or an excavator



Use of protective devices such as rock fences in critical areas where personnel or equipment may be at risk from a ravelling slope



Provision in mine plans for alternative access into the pit, should the west pit wall ramp serviceability become affected.

Wall heights are 30 m to be a multiple of the regularised mining model block height of 15 m. Batter angles are 70 degrees. Berm widths were calculated based on the overall slope recommendations from Knight Piésold, and the depth of pit and number of ramps in the walls as per the PFS pit design. Pit access ramps are 30 m wide with 1:10 gradient and were designed to accommodate 220 t class trucks including allowances for the construction of safety windrows and drainage. A 15 m single lane ramp was designed for the last 60 vertical meters of pits and sub-pits. Pit design inventories reported at a zero profit margin reporting cut-off and at an elevated ($3.00 profit margin) reporting cut-off are summarised in Table 18.5. The ultimate and staged pit designs are illustrated in Figure 18.2 to Figure 18.8.

Page 246

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.5 Pit Design Inventories

Constancia Pit Stage Scheduling Data at zero profit margin reporting cutoff Pit Stage Const Stg 1a Const Stg 2a Const Stg 3a Const Stg 4a San Jose 1a Total:

Ore (kts) 66 117 63 228 74 699 109 024 13 722 326 790

Cu (%) 0.61% 0.42% 0.33% 0.28% 0.44% 0.39%

Zn (%) 0.13% 0.13% 0.12% 0.07% 0.11% 0.11%

Total Concentrator Feed Inventory Mo Ag Au Pb CuCon (g/t) (g/t) (g/t) (%) (%) 128.5 4.61 0.06 0.04% 29.33% 98.8 3.78 0.04 0.05% 28.65% 131.3 3.34 0.03 0.04% 28.48% 88.9 2.63 0.04 0.05% 28.89% 70.4 3.80 0.09 0.03% 28.64% 107.8 3.46 0.04 0.04% 28.83%

Mo rec Sulphur Ore (%) (%) (kbcm) 55.00% 3.06% 26 794 55.00% 2.69% 25 106 55.00% 2.34% 28 753 55.00% 1.82% 41 653 55.00% 2.82% 5 334 55.00% 2.40% 127 641

Total Waste (kts) 51 846 56 588 79 057 101 773 10 763 300 026

Strip Ratio

Mo rec Sulphur Ore (%) (%) (kbcm) 55.00% 3.09% 25 800 55.00% 2.74% 21 178 55.00% 2.40% 21 394 55.00% 1.91% 24 516 55.00% 2.81% 4 163 55.00% 2.54% 97 050

Total Waste (kts) 54 358 66 712 98 205 146 628 13 791 379 693

Strip Ratio

0.78 0.89 1.06 0.93 0.78 0.92

at $3.00 profit margin reporting cutoff Pit Stage Const Stg 1a Const Stg 2a Const Stg 3a Const Stg 4a San Jose 1a Total:

Ore (kts) 63 605 53 104 55 551 64 168 10 694 247 123

Cu (%) 0.63% 0.47% 0.38% 0.35% 0.52% 0.46%

Zn (%) 0.13% 0.14% 0.12% 0.08% 0.11% 0.12%

Total Concentrator Feed Inventory Mo Ag Au Pb CuCon (g/t) (g/t) (g/t) (%) (%) 131.8 4.68 0.06 0.04% 29.48% 108.7 4.08 0.05 0.05% 28.95% 154.0 3.72 0.03 0.04% 28.87% 106.7 3.03 0.05 0.05% 29.38% 74.1 4.30 0.10 0.04% 29.15% 122.8 3.89 0.05 0.04% 29.19%

0.85 1.26 1.77 2.29 1.29 1.54

Figure 18.2 Ultimate Pit Design Plan View

Page 247

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.3 Constancia Stage 1 Design Plan View

Figure 18.4 San José Pit Design Plan View

Page 248

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.5 Constancia Stage2 Design Plan View

Figure 18.6 Constancia Stage 3 Pit Design Plan View

Page 249

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.7 Constancia Stage 4 Pit Design Plan View

Figure 18.8 Constancia Ultimate and Staged Pit Designs Plan View

Page 250


18.1.4

Mineral Reserve

The Constancia Mineral Reserve is the Measured and Indicated Resource contained in the open pit mine that can be processed at a profit and is scheduled for treatment in the DFS Life-of-Mine (LOM) plan. Since revenue is derived from four payable components (copper, molybdenum, silver and minor payable gold) the reserve reporting cut-off is based on a NSR cut-off that is estimated using the metal prices and other treatment, recovery and concentrate realisation parameters as detailed in Section 18.14.5. The Mineral Reserve estimate, comprising Proven and Probable categories, is summarised in Table 18.6. Table 18.6 Constancia Project Global Mineral Reserve Estimate Category

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

PROVEN

161.8

0.45

0.012

3.68

0.05

PROBABLE

115.6

0.40

0.011

3.70

0.05

TOTAL

277.4

0.43

0.012

3.69

0.05

The Constancia mine consists of a main pit and a small satellite orebody, San José, that overlaps with the Main pit at the end of the mine life. The main Constancia orebody is scheduled to be mined in four consecutive pit stages while the San José pit will be mined in a single stage. Table 18.7 summarises the Mineral Reserve by pit stage. Table 18.7 Constancia Project Mineral Reserve by Pit Stage Category

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

CONSTANCIA STAGE 1

64.6

0.62

0.013

4.65

0.06

CONSTANCIA STAGE 2

57.0

0.45

0.010

3.94

0.04

CONSTANCIA STAGE 3

63.0

0.36

0.014

3.54

0.03

CONSTANCIA STAGE 4

81.1

0.31

0.010

2.80

0.04

SAN JOSE

11.7

0.49

0.007

4.08

0.10

TOTAL

277.4

0.43

0.012

3.69

0.05

The mineralisation within the Constancia deposit has been sub-divided into three major ore types being Skarn, Supergene and Hypogene. This fundamental sub-division was established during resource modelling and carried through into the mining model to allow the different recovery, concentrate quality and plant throughput impacts associated with each ore type to be properly identified and tracked in the LOM processing schedule. Subsequent to resource modelling, a further ore type based on Zn to Cu ratios (“High Zinc”) was established. Ore types and their metallurgical response are discussed further in Section 16. Table 18.8 summarises the Mineral Reserve by ore type.

Page 251

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.8 Constancia Project Mineral Reserve by Ore Type Category SUPERGENE

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

71.0

0.58

0.011

4.43

0.05

SKARN

19.3

0.64

0.013

5.01

0.08

HYPOGENE

163.6

0.35

0.012

2.95

0.04

HIGH ZINC

23.5

0.40

0.010

5.51

0.07

TOTAL

277.4

0.43

0.012

3.69

0.05

In order to maximise project value, higher operating margin ore was treated preferentially through the LOM processing schedule. In order to achieve this outcome, the in-pit mineralisation was sub-divided and identified in the mining model as being one of three operating margin ranges: 1. $0.00 to $3.00 margin per tonne of ore (“low margin ore”) 2. $3.00 to $5.00 margin per tonne of ore (“ medium margin ore”) 3. >$5.00 margin per tonne of ore (“high margin ore”). Table 18.9 summarises the Mineral Reserve by operating margin. Table 18.9 Constancia Project Mineral Reserve by Operating Margin Category

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

> $5.00/t

198.9

0.51

0.013

4.20

0.05

>$3.00/t to $5.00/t

48.2

0.26

0.008

2.63

0.03

$0.00/t to $3.00/t

30.3

0.19

0.007

2.02

0.03

277.4

0.43

0.012

3.69

0.05

TOTAL

A number of LOM plan scenarios were developed in order to identify the case that delivers the optimal financial outcome for the project. The adopted strategy treats low margin mineralisation on an opportunistic basis to maintain ore supply to the plant while maintaining total mining at a consistent rate. As a result, some 49 Mt of low margin mineralisation is mined during the project life, but is not scheduled for treatment. Although this mineralisation will theoretically generate a positive operating margin, if processed, it has been excluded from the reserve statement since: 

It is not scheduled for treatment, and associated capital costs (tails dam expansion) have not been allowed in the study



The material is stockpiled for a period of between 1 and 14 years. It is unknown to what extent the flotation response and associated metal recovery will be compromised as a result of oxidation during this protracted period.

This mineralisation represents a potential upside to the project if metal prices, rehandle costs, tails storage costs and flotation recovery at the time result in a favourable economic outcome.

Page 252


Low margin mineralisation scheduled to be mined but excluded from the Mineral Reserve statement is summarised in Table 18.10. Table 18.10 Low Margin Mineralisation excluded from the Mineral Reserve Category

Tonnes (M)

Cu%

Mo%

Ag g/t

Au g/t

49.4

0.18

0.006

2.21

0.03

$0.00 to $3.00/t

18.1.5

Mine and process schedules

Bench reporting of reserve information was performed for the pit stages and imported into a purposebuilt mine scheduling spreadsheet. A range of mining rates and flotation plant production profiles was investigated, before a final mining rate of around 45 Mt/a was adopted as the optimum sustainable rate that will bring forward the mining and treatment of higher grade concentrator feed. Ore mined will be hauled and fed directly to the crusher. There is limited surge capacity at the ROM pad with minimal rehandle envisaged. The bulk of the low operating margin material will be stockpiled in the footprint to the long-term PAG waste rock dump. Low margin material is fed to the plant only during ramp-up over the first three months (when plant recoveries are expected to be low) and when insufficient better margin material is available, e.g. during transition between pit stages and towards the end of mine life. Mine and process scheduling was carried out on a monthly basis for the pre-strip (Year-1) and first year of production, quarterly for Years 2 through 5 and annually thereafter. Figure 18.9 illustrates the mining production rate by pit stage over the mine life.

Page 253

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.9 Mining by Pit Stage (Mt) 50

Const Stg 4a Mt

45

Const Stg 3a Mt

40

San Jose 1a Mt Const Stg 2a Mt

35

Const Stg 1a Mt

30 25 20 15 10 5

Yr 22

Yr 23

Yr 24

Yr 25

Yr 22

Yr 23

Yr 24

Yr 25

Yr 21

Yr 20

Yr 19

Yr 18

Yr 17

Yr 16

Yr 15

Yr 14

Yr 13

Yr 12

Yr 11

Yr 10

Yr 9

Yr 8

Yr 7

Yr 6

Yr 5

Yr 4

Yr 3

Yr 2

Yr 1

Yr‐1

0

The presentation of potential ore by material type is summarised in Figure 18.10. Figure 18.10 Ore Mining by Material Type (Mt) 35

High Zn Mt 30

Skarn 2 Mt Skarn 1 Mt

25

Supergene Mt Hypogene Mt

20

15

10

5

Yr 21

Yr 20

Yr 19

Yr 18

Yr 17

Yr 16

Yr 15

Yr 14

Yr 13

Yr 12

Yr 11

Yr 10

Yr 9

Yr 8

Yr 7

Yr 6

Yr 5

Yr 4

Yr 3

Yr 2

Yr 1

Yr‐1

0

Page 254


Table 18.11 shows the annual mining production schedule. Concentrator ore processing is shown in Figure 18.11, and long-term ore stockpile inventories are illustrated in Figure 18.12.

Page 255

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.11 MINING Float Ore Cu Zn Mo Ag Au Pb Sulphur Waste Total Mining Strip Ratio

(kts) (%) (%) (g/t) (g/t) (g/t) (%) (%) (kts) (kts)

Total 326 790 0.39% 0.11% 107.8 3.46 0.04 0.04% 2.40% 300 026 626 816 0.92

CLOSING STOCKPILES High NSR (kts) Medium NSR (kts) Low NSR (kts) Total (kts) PROCESSING Float Ore Cu Zn Mo Ag Au Pb Sulphur

(kts) (%) (%) (g/t) (g/t) (g/t) (%) (%)

277 392 0.43% 0.11% 116.7 3.69 0.05 0.04% 2.46%

CONCENTRATE PRODUCTION - COPPER dry tonnes 3 766 324 Cu (%) 28.1% Zn (%) 2.93% Ag (g/t) 217.2 Au (g/t) 2.10 Pb (%) 0.79% Cu in Con (t) 1 057 992 Ag in Con (kozs) 26 302

Yr-1 822 0.37% 0.07% 42.8 3.77 0.08 0.05% 0.53% 12 763 13 585 15.53

Yr 1 21 952 0.58% 0.10% 89.8 4.23 0.07 0.04% 2.28% 23 597 45 549 1.07

Yr 2 22 971 0.58% 0.11% 116.4 4.43 0.06 0.04% 3.01% 21 804 44 776 0.95

Yr 3 23 143 0.63% 0.15% 152.4 5.07 0.05 0.04% 3.40% 21 786 44 930 0.94

733 58 31 822

527 610 1 138

713 1 505 2 218

783 3 126 3 910

829 6 617 7 445

799 12 176 12 976

751 16 988 17 739

21 636 0.59% 0.10% 90.1 4.24 0.07 0.04% 2.24%

21 892 0.60% 0.11% 119.8 4.50 0.06 0.04% 3.06%

21 451 0.67% 0.15% 162.1 5.25 0.06 0.04% 3.50%

21 055 0.50% 0.10% 108.2 4.35 0.05 0.06% 2.42%

19 642 0.49% 0.10% 98.7 4.02 0.07 0.04% 2.91%

357 411 28.9% 3.55% 205.3 2.45 0.57% 103 154 2 359

409 706 28.8% 2.66% 192.4 1.83 0.49% 117 883 2 534

450 872 28.5% 2.72% 199.9 1.60 0.52% 128 692 2 898

332 281 28.6% 2.79% 220.3 1.77 0.91% 94 943 2 353

2 424 40.0% 970

3 606 40.0% 1 443

4 782 40.0% 1 913

9 457 13 194 946 23 597

10 823 5 513 5 468 21 804

6 775 10 019 4 993 21 786

-

-

CONCENTRATE PRODUCTION - MOLY dry tonnes 44 245 Mo (%) 40.0% Moly in Con (t) 17 698 Waste Mining Summary NPAG (kts) Oxide (kts) Rest (kts) All Waste (kts)

54 877 61 510 183 639 300 026

Annual Mining and Processing Schedule Yr 4 Yr 5 Yr 6 Yr 7 24 590 25 173 24 624 16 886 0.46% 0.42% 0.39% 0.34% 0.10% 0.10% 0.14% 0.19% 98.3 88.2 96.9 106.8 4.04 3.58 3.63 3.75 0.04 0.06 0.05 0.04 0.06% 0.04% 0.04% 0.06% 2.42% 2.84% 2.86% 2.55% 20 326 19 670 20 209 28 022 44 916 44 843 44 832 44 908 0.83 0.78 0.82 1.66

11 463 1 260 40 12 763

Yr 8 22 687 0.32% 0.12% 110.2 3.15 0.03 0.04% 2.65% 22 098 44 785 0.97

Yr 9 23 214 0.33% 0.11% 158.7 3.17 0.03 0.04% 2.29% 21 527 44 741 0.93

Yr 10 18 790 0.34% 0.12% 154.4 2.99 0.03 0.03% 2.07% 25 991 44 780 1.38

Yr 11 19 335 0.26% 0.05% 90.3 2.51 0.03 0.05% 1.60% 25 522 44 857 1.32

Yr 12 24 956 0.27% 0.09% 100.6 3.13 0.04 0.07% 2.05% 19 699 44 655 0.79

Yr 13 30 572 0.28% 0.09% 78.4 2.69 0.05 0.04% 2.10% 12 344 42 916 0.40

Yr 14 22 651 0.31% 0.06% 86.4 2.31 0.04 0.03% 1.70% 4 667 27 318 0.21

Yr 15 4 425 0.41% 0.03% 108.4 2.72 0.06 0.03% 0.86%

16 274 16 274

21 758 21 758

404 27 461 27 866

30 018 30 018

33 244 33 244

41 565 41 565

311 54 788 55 099

60 893 60 893

49 398 49 398

19 860 0.44% 0.15% 107.5 3.96 0.06 0.04% 3.00%

18 351 0.32% 0.18% 102.8 3.72 0.04 0.05% 2.50%

17 203 0.37% 0.12% 125.1 3.44 0.04 0.04% 2.74%

17 106 0.38% 0.12% 187.3 3.56 0.03 0.04% 2.42%

16 638 0.36% 0.11% 164.9 3.16 0.03 0.03% 2.09%

16 109 0.28% 0.04% 95.9 2.49 0.03 0.04% 1.50%

16 635 0.31% 0.09% 120.2 3.47 0.05 0.07% 2.10%

17 038 0.35% 0.10% 94.1 3.25 0.06 0.06% 2.36%

16 856 0.36% 0.07% 96.1 2.52 0.05 0.03% 1.85%

15 920 0.25% 0.04% 76.0 2.13 0.04 0.03% 1.51%

304 352 28.3% 2.64% 207.5 2.58 0.63% 86 010 2 031

287 650 27.2% 3.64% 218.8 2.30 0.75% 78 187 2 024

210 241 25.6% 5.11% 259.5 1.99 1.17% 53 879 1 754

211 970 27.0% 3.14% 223.2 1.75 0.82% 57 279 1 521

215 532 27.4% 2.99% 226.2 1.40 0.75% 59 036 1 567

196 298 28.1% 2.82% 214.1 1.38 0.57% 55 198 1 351

134 200 29.0% 1.48% 239.5 2.41 1.34% 38 899 1 033

163 780 27.6% 2.71% 281.9 2.79 1.83% 45 177 1 485

190 048 27.7% 2.85% 233.4 3.12 1.24% 52 555 1 426

183 100 28.6% 2.02% 185.7 2.58 0.62% 52 429 1 093

118 882 29.2% 1.59% 228.1 2.84 0.96% 34 672 872

3 131 40.0% 1 252

2 666 40.0% 1 066

2 937 40.0% 1 175

2 594 40.0% 1 037

2 960 40.0% 1 184

4 406 40.0% 1 763

3 772 40.0% 1 509

2 124 40.0% 850

2 749 40.0% 1 100

2 204 40.0% 882

2 227 40.0% 891

1 663 40.0% 665

2 976 6 994 10 356 20 326

4 188 6 250 9 233 19 670

3 040 3 886 13 283 20 209

1 272 2 482 24 269 28 022

116 1 039 20 943 22 098

3 258 7 040 11 229 21 527

710 2 410 22 872 25 991

611 1 375 23 536 25 522

126 49 19 524 19 699

38

25

12 305 12 344

4 642 4 667

4 425

Page 256

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.11 Float Ore Processing Plant Feed by NSR Range (Mt) 25

Low NSR Mt

20

Medium NSR Mt High NSR Mt 15

10

5

Yr 24

Yr 25

Yr 24

Yr 25

Yr 23

Yr 22

Yr 21

Yr 20

Yr 19

Yr 18

Yr 17

Yr 16

Yr 15

Yr 14

Yr 13

Yr 12

Yr 11

Yr 10

Yr 9

Yr 8

Yr 7

Yr 6

Yr 5

Yr 3

Yr 4

Yr 2

Yr 1

Yr‐1

0

Figure 18.12 Closing Ore Stockpile Inventories (Mt) 70

Low NSR Mt 60

Medium NSR Mt High NSR Mt

50

40

30

20

10

Yr 23

Yr 22

Yr 21

Yr 20

Yr 19

Yr 18

Yr 17

Yr 16

Yr 15

Yr 14

Yr 13

Yr 12

Yr 11

Yr 10

Yr 9

Yr 8

Yr 7

Yr 6

Yr 5

Yr 4

Yr 3

Yr 2

Yr 1

Yr‐1

0

Page 257


18.1.6

Ore types and batching

The three major ore types identified in the Constancia resource model are Skarn, Supergene and Hypogene. Mineralisation that is high in zinc and proximal to the Supergene ore may be problematic in terms of producing elevated Zn levels in the copper concentrate, potentially affecting its marketability. In order to better understand potential impacts, a spatial/analytical routine was developed to sub-divide Skarn in the resource model into Skarn1 and Skarn2, where Skarn2 mineralisation is high in zinc and proximal to Supergene mineralisation. The resource and regularised mining model were modified to include the resultant skarn sub-division. Constancia ore and waste loading operations are scheduled to be performed using large electric shovels capable of mining on benches with a 15 m face height. This approach minimises mining costs, but results in poor ore type selectivity relative to other mining approaches. To simulate realistic mining selectivity, the resource block model was regularised, meaning that the grade attributes of the constituent ore types within a parent cell were averaged. Ore types in the mining model were then assigned on a majority basis i.e. if a parent block contained more than 50% Supergene it is designated as Supergene. As a result of ore regularisation, those blocks around ore type margins are a mixture of ore types, but are identified only by the majority constituent. Subsequent to creating the mining model, a further “High Zinc” ore type was created. Any ore blocks with a Zn/Cu ratio of greater than 0.66 were designated as high zinc. High zinc ore is principally a subset of the Skarn1 and Skarn2 mineralisation. However, a significant quantity (>20%) was from Hypogene mining blocks, while a lesser amount (<10%) was derived from Supergene mining blocks. It is most probable that the majority of high zinc ore from these latter two mineralisation types is actually related to majority Hypogene and Supergene mining blocks with minor (<50%), but significant Skarn content. During processing, the metallurgical performance and copper concentrate quality will be optimised if high zinc and Supergene ores are not mixed for treatment. However, mixing of other ore types is not a treatment issue, so the process methodology will be to batch treat ore types through the concentrator with the feed consisting of: 1. Supergene and/or Hypogene 2. Skarn and/or Hypogene. Process operations advice is that batch durations can be as short as one week, although a longer period would be desirable. With the short duration, plant batching should normally be controllable by managing the respective ore types in the pit, with a minimal requirement for short term stockpiling and re-handling of ore. In any event, the mining estimate allows for routine re-handling of approximately 20% of ROM feed as a result of crusher stoppages or congestion during normal operations.

Due to the bulk mining approach and the complex mineralisation shapes it is inevitable that there will be some mixing of high zinc and Supergene ore types in the plant feed. This mixing will be controlled by ore control practices comprising:

Page 258




Routine face mapping to identify and predict Skarn/Supergene contacts



Routine logging and sampling of blasthole cuttings to identify ore types



Batch designation as part of short term planning and grade control practices.

Where Skarn/Supergene mixing occurs, it should normally be possible to identify problematic ore prior to mining and apply appropriate management strategies. This will be a learning process that will evolve as the short term plant and concentrate response is better understood during daily operations. 18.1.7 18.1.7.1

Mine fleet assessment Design considerations

The following factors were considered in the selection and sizing of mining equipment for Constancia: 

Nature and occurrence of the orebody



Mining dilution and ore losses



Ore blending requirements



Match with the production schedule



Shovel and haul truck size to match crusher to minimise ROM rehandle



Performance at high altitude



Standardisation of the equipment fleet



Mining fleet integrated with civil construction fleet – mine trucks to haul non-PAG waste to construct the TMF dam



Local knowledge and the dealers network



Maintenance of the equipment

Constancia is a massive copper porphyry deposit and hence is amenable to conventional bulk mining techniques using shovels and trucks. The massive nature of the orebody also means that mining dilution and ore loss are not expected to be significant issues. The selected mining schedule requires the mining fleet to move a total of approximately 45 Mt of ore and waste material annually. The first step in the fleet sizing and selection process was to determine the most suitable loading fleet with the capacity to load the required annual total material movement. The size of the major loading unit is a balance between the required capacity, the width and number of available mining areas, blending requirements, mining selectivity, and matched to the crusher capacity. Constancia will essentially be a single pit operation with the small San José pit located to the westnorthwest joining the Constancia pit close to the surface. The initial pit stages have relatively small mining areas, which will be limited in number during the early stages of mining. GRD Minproc considers that two digging units will be most suitable.

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The grade distribution of the Constancia deposit varies by depth. Higher grade ore is located close to the surface with grade decreasing at depth. It is envisaged that no blending between grade categories will be required. Some processing difficulties have been identified in mixing certain types of ore (i.e. supergene with skarn) due to the high zinc contents of the latter. This will be addressed by batch treating instead of blending. Electric powered mining equipment has advantages over diesel powered equipment for mines located in high altitude as performance is not as affected by low levels of oxygen. Electric equipment also has lower operating costs than diesel equipment due to electric power being cheaper than diesel fuel. Electric equipment is, however, significantly more expensive to purchase than diesel equipment and also requires an electrical power distribution infrastructure. Electric equipment also needs supporting cable crew and equipment. On the other hand, electric equipment lasts longer than diesel equipment and hence is more capital effective for longer life mines as replacement capital is reduced. After analysis of available data, the study was based on electric shovels, and diesel powered drills. A detailed trade-off study between electric and diesel equipment needs to be carried out prior to final selection of mining equipment for the operation. 18.1.7.2

Mining fleet definition

The selected mine fleet is as follows: 

Shovels A 32 m3 capacity electric shovel was selected as the primary digging unit for its lower operating cost and better performance compared to diesel powered shovels in bulk mining applications at high altitudes. The total material movement requirement of around 45 million tonnes per year can be satisfied by two electric shovels with a dipper capacity of 32 m3. Electric shovels in the 3035 m3 class can operate productively at the 15 m bench height. The selected shovels also have production rates that are a good match for the peak throughput rates of the primary crusher. Each electric shovel is estimated to be capable of loading around 25 Mt/a of ore and waste material.



Drills The selected bench height is 15 m to optimise machine productivity and efficiency and reduce blasting costs. Production drills are expected to drill the 15 m bench in a single pass, thereby improving productivity by reducing lost time in joining drill rods. Crawler-mounted diesel drills have been selected for productivity and operational flexibility. Three drill machines capable of 15 m single pass drilling at 279 mm diameter will be purchased to cater for the drilling requirements of the project. Drilling operations will be carried out by Norsemont. A pre-spilt drill will also be purchased to minimise blast damage to the pit walls.



Haul Trucks In order to keep operating costs low, 220 t class haul trucks and support equipment have been selected. A fleet estimation software, TALPAC, was used to calculate truck cycle times. The haul road profile strings for ore and waste were prepared using SURPAC software. Truck performance

Page 260


was adjusted to the 4500 masl altitude by derating the rimpull force, engine power, and retardation performance based on the equipment suppliers high altitude derating factor specifications. The truck haulage cycle time was estimated for each haulage profile, and for each year of mine life. Based on the above information, the truck requirement for each year of mine was estimated. After smoothing, the average truck requirement is for 13 units, with a build-up occurring over the first six months as shown in Figure 18.13. Figure 18.13 Haul Truck Requirements

Period Truck Requirement ROM Ore PAG5 Waste LTSP Rehandle Trucks Required

LG PAG10 Waste Trucks Required

Non‐PAG5 Waste PAG15 Waste Trucks Purchased

25

Trucks Required

20

15

10

5

0



FEL A multi-functional, FEL with 18 m3 bucket has been specified to serve the following functions: Provide production loading back-up in the pit when a primary shovel is unavailable Provide truck loading for rehandle of marginal ore feed stockpiles Rehandle (by tramming) from short-term operational stockpiles located on the ROM pad Provide pioneering support to establish initial shovel digging areas.



Support Equipment A mine support equipment fleet is required for maintenance of haul roads, waste dumps, stockpiles, ROM pads, face cleaning and preparation of production faces near the shovels. The support fleet consists of two track dozers, one wheel dozer, two water carts, and two motor graders. A small integrated tool carrier is used for material handling purposes.

Page 261


A small excavator with rock breaker attachment is required for breaking big boulders produced during blasting operations. A computerised dispatch system has been allowed, to monitor equipment, provide production statistics and provide the information to measure and improve fleet productivity. Table 18.13 includes a summary of the major equipment selected. 18.1.8

Mine operating cost

The DFS mining concept is that mining equipment will be owned, operated and maintained by Norsemont, with support by specialist contractors in the following areas: 

Down-the-hole explosive supply



Diesel supply, storage and dispensing services.

Explosive supply will be by a local vendor providing a down-the-hole service. ANFO will be the main explosive with waterproof explosive allowed for potential wet conditions. After assessment of rock properties, powder factors of 0.30 to 0.35 kg/t in waste and 0.35 to 0.40 kg/t in ore have been adopted. Higher powder factors in ore will deliver the fragmentation required to minimise the presentation of oversize rocks to the primary crusher and reduce the plant energy requirements. Mine operating costs have been built up from first principles. Major cost drivers are labour, diesel, power, and explosives. Key operating cost drivers are summarised in Table 18.12. Diesel prices were provided by Norsemont. Table 18.12 Key Operating Cost Inputs Item Diesel AN explosive

Value

Unit

Comment

0.66

$/litre

includes storage & dispensing ANFO (20% wet with heavy ANFO)

533

$/t

Powder factor - ore

0.35 to 0.40

kg/t

Powder factor - waste

0.30 to 0.35

kg/t

4 500

hrs

Truck tyre life

Truck fleet requirement has been built up from loading unit and truck productivities on nominated haul profiles using TALPAC software adjusted for altitude. Mining equipment vendors provided equipment hourly major parts costs. Other significant costs collated on an hourly basis included: diesel or power consumption, ground engaging tools (GET), tires, servicing, and lubricants. The proposed major equipment fleet make-up is summarised in Table 18.13, together with key equipment assumptions used to build up the operating cost estimate. While specific equipment models have been used to build up the estimate, actual fleet configuration would be subject to a further tendering and evaluation process to establish the most cost-effective mining solution.

Page 262

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.13 Equipment Fleet and Hourly Costs

Type

Equipment Class

Shovel 32 m3 Dump Truck 220 tonne FEL 18 m3 Track Dozers 391 kW Wheel Dozers 362 kW Graders 209 kW Water Truck 91 kl 2.5 m3 Integrated Tool Carrier Rock Breaker H130S Hammer Excavator General Duties 12m3 Cable Reeler Production Drill Diesel 279 mm Pre‐split Drill 127 mm Service Truck Tyre Handler Low Loader Light Vehicle Passenger Bus Lighting Plants 9m Hydraulic Mast

Fleet Units

2 13 1 2 1 2 2 1 1 1 1 3 1 1 1 1 14 2 8

Operating Operating Purchase Hours Costs Price Hrs/yr US$/hr US$M 6 701 6 701 5 585 5 046 4 840 5 606 3 723 3 723 2 234 3 723 1 489 4 906 5 957 3 723 2 102 491 1 862 1 862 3 723

283.99 267.28 345.81 129.11 183.38 114.18 105.66 44.29 54.62 188.65 118.75 276.64 159.68 20.00 32.86 103.54 9.45 11.20 8.17

Expected Life Hrs

16.29 100 000 3.43 65 000 3.74 50 000 0.94 30 000 0.82 30 000 0.67 30 000 1.54 60 000 0.23 40 000 0.42 35 000 2.79 60 000 0.86 30 000 1.54 60 000 0.70 25 000 0.17 25 000 0.15 30 000 0.63 20 000 0.03 10 000 0.27 15 000 0.03 25 000

Note: 5% vendor budget pricing price reduction applied to reflect competitive equipment tender market.

Figure 18.14 illustrates the mining operating unit cost with time, showing the major operating cost components. The mining costs are inclusive of all material rehandling from stockpiles.

Page 263

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.14 Operating Costs by Time ($/t)

$4.00 $3.50 $3.00

Other

Labour

Support Equip

Blast

Drill

Hauling

Loading

$2.50 $2.00 $1.50 $1.00 $0.50

Yr15

Yr14

Yr13

Yr12

Yr11

Yr10

Yr9

Yr8

Yr7

Yr6

Yr5

Yr4

Yr3

Yr2

Yr1

Yr‐1

$0.00

The average mine operating cost for the life of the mine including pre-strip is $1.19/t mined. Unit costs in the pre-production year are higher as a result of earthworks associated with pioneering activities to establish the electric shovels and construction of access roads and ramps. High unit costs at the end of the mine life are the results of a high level of stockpile rehandle activities off-setting reduced material movement from the mine. Figure 18.15 shows the mine operating cost broken down into its major components. It can be seen in this chart that haulage is the biggest contributor at $0.48/t, followed by production blasting at $0.22/t, then by support equipment at $0.19/t. Loading and labour cost are identical at $0.09/t.

Page 264

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.15 Mine Operating Costs Components

Unit Operating Costs (US$/t) $0.09

$0.09

Load

$0.02

Haul $0.19 Production Drilling Presplit Drilling

$0.02

Production Blasting $0.01 $0.48

Presplit Blasting Blasting Contractor Support Equipment

$0.22

Labour $0.01 $0.07

18.1.9

Other

Mine capital cost

The total mining capital cost (including capitalised mining operating costs in the pre-production period, equipment replacement, rebuilds and sustaining capital) is estimated at $164.93 M, as detailed in Table 18.14.

Page 265


Table 18.14 Mining - Capital, Sustaining and Replacement Costs ($ M) Mine Area

Yr-1

Loading Hauling

Yr 1

Yr 2

Yr 3

Yr 4

Yr 5

Yr 6

Yr 7

Yr 8

Yr 9

36.73

0.06

0.31

0.06

0.06

0.26

0.06

0.06

3.98

45.63

0.45

Drill & Blast

3.86

1.54

Support

7.11

2.79

Other

7.19

Prestrip Total:

0.34

0.61

0.34

5.45

0.45

Yr 10

Yr 11

Yr 12

Yr 13

Yr 14

0.06

0.06

0.26

0.06

0.06

Yr 15

42.04

23.91

70.44

0.35 4.03 0.36

0.61

0.07

5.82

3.62

17.56

0.23

0.73

10.08

24.55

4.41

19. 0 119.51

Total

19.0 0.31

0.51

0.42

4.29

0.67

0.06

3.98

0.26

0.06

0.06

0.07

164.93

Page 266


Table 18.15 outlines the major components of the mine capital costs estimated for the project.

Mine capital costs do not include the infrastructure costs related to the mine including a $1.0 M provision to provide initial fuel, explosive and workshop facilities for the civil fleet. Those costs are included under general infrastructure facilities. Table 18.15 Mine Capital Costs by Component

CAPITAL COSTS Units ITEMS Equipment Purchases Replacement Purchases & Overhauls Start up Spares Trailing Cables Replacement & Repair - Electric Shovel Cable Towers and Crossing Ramps, cable trays Tyres ( one set only) Crane rental for equiment commisioning Survey Equipment Shovel dipper Truck Tray Mining Software & Systems Heavy Equip Assembly ‐ Crane hire, transport etc Blasting Contractor Mob/Demob Dispatch and fleet management system Prestrip Total

$ $ $ $ $ $ $ $ $ $ $ $ $ $ $ $

Total 99 580 055 37 591 030 2 761 760 1 008 000 1 000 000 252 387 180 000 80 000 500 000 891 292 150 000 600 000 136 000 1 200 000 19 002 416 164 932 941


The mine capital cost estimates are based predominantly on budget quotations from equipment suppliers. The quotes include delivery to site, and expenses related to commissioning of the equipment and training of operators. A detailed list of standard equipment and accessories has been included in the quotes. Table 18.16 shows the capital cost of the mining fleet envisaged for the Constancia project including the expected life of the equipment. The electric shovels are by far the most expensive individual piece of equipment at about $16.8 M each, but the largest area of expenditure is for haul trucks ($45.9 M).

Page 267

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.16 Equipment Unit Capital Cost - Mining Fleet

Equipment Summary (with expected competitive tender market price reduction) Type

Shovel Dump Truck FEL Track Dozers Wheel Dozers Graders Water Truck Integrated Tool Carrier Rock Breaker Excavator General Duties Cable Reeler Production Drill Diesel Pre‐split Drill Service Truck Tyre Handler Low Loader Light Vehicle Passenger Bus Lighting Plants

Capacity

32 m3 220 tonne 18 m3 391 kW 362 kW 209 kW 91 kl 2.5 m3 H130S Hammer 12m3 279 mm 127 mm

9m Hydraulic Mast

Purchase Price US$

Tyres US$

Start‐up Spares US$

16 287 138 492 671 3 176 172 252 387 100 300 3 579 061 156 412 104 377 937 040 25 404 819 556 21 698 670 554 17 698 1 536 978 42 418 234 546 7 407 422 932 13 356 2 788 424 79 056 863 497 19 500 1 544 676 37 725 698 250 21 300 166 250 5 250 153 644 4 852 626 463 19 783 25 935 819 266 000 8 400 34 885 1 102

Fleet Units

Expected Life Hrs

2 13 1 2 1 2 2 1 1 1 1 3 1 1 1 1 14 2 8

100 000 65 000 50 000 30 000 30 000 30 000 60 000 40 000 35 000 60 000 30 000 60 000 25 000 25 000 30 000 20 000 10 000 15 000 25 000


18.2

GEOTECHNICAL STUDIES

18.2.1 Introduction Knight Piésold carried out geotechnical and hydrogeological investigations at the site in order to (1) map and characterise the geologic units and structures, (2) gain an understanding of the subsurface soil and rock types, their depths and engineering properties, (3) characterise the ground water and hydrogeological flow regimes, (4) identify and characterise potential borrow material sources to evaluate their suitability for use as construction materials, and (5) obtain soil and rock samples for laboratory testing. The geotechnical investigation included the following: 

Geologic site mapping and reconnaissance



Excavation and sampling of test pits



Dynamic Probing Light (DPL) testing to determine the depths of soft, wet organic deposits in certain areas

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

Geotechnical drilling and sampling at selected sites



Standard Penetration Testing (SPT) and Large Penetration Testing (LPT) in the geotechnical drillholes



In-situ permeability testing in the geotechnical drillholes



Piezometer installation and monitoring in the geotechnical drillholes



Sample recovery and testing for index, compressibility, strength and permeability properties.

The field work was conducted in phases between October 2007 and June 2009 and focused on areas where the main project structures will be located. These included the open pit, TMF, PAG WRF and the process plant. Plan view illustrations of the geotechnical investigation program and the results of the geological and structural mapping work are presented in Figure 18.16 and Figure 18.17, respectively.

Page 269


Figure 18.16 Geotechnical Site Investigation

Page 270

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.17 General Geologic Map

Page 271


The geotechnical investigation program consisted of the following activities: Open pit area 

Ten angled and oriented core drillholes in the Constancia pit area, with depths ranging from 140 to 318 m. Most of the oriented core holes were located and oriented specifically to intercept major fault zones that were considered at the time to have potential stability impacts (in plane shear or as wedge components) on the final pit walls. These holes were generally oriented at angles opposite to the slope directions of the pit walls.



Geomechanical logging and lithological descriptions of the cored materials from these holes and 244 additional exploration drillholes completed by Norsemont between 2005 and 2008.



Surface outcrop mapping using conventional structural geological mapping techniques, augmented with detailed geomechanical measurement of rock fabric at 49 sites using window mapping techniques. There is a paucity of surface outcrops available at the site, so most of the geomechanical and structural mapping was made from rock exposures in the San José pit, trenches dug across major structure projections, drill pad cuts and access road cuts.



Field testing including SPT tests, downhole permeability tests in soil and rock, and point load tests on recovered drill core. The point load tests were carried out routinely on exploration as well as geotechnical drill core, and were supported by laboratory compressive strength (UCS) tests on recovered drill core.



Rock laboratory testing including index properties, slake durability, elastic properties (Young’s modulus and Poisson’s ratio); triaxial (confined compression); point load index, UCS (unconfined compression); Brazilian indirect tension, sonic velocity, and small-scale direct shear of discontinuities in samples of recovered core and hand-dug samples of fault gouge

Tailing Management Facility and Associated Structures 

Seven geotechnical drillholes with depths ranging from 35 to 150 m



Six hydrogeological drillholes with depths ranging from 38 to 125 m



Geomechanical logging and lithological descriptions of the cored materials from these holes



Installation and completion of six open standpipe piezometers in selected gotechnical holes



Excavation of 86 test pits



Field testing including SPT testing, down hole permeability tests in soil and rock, point load tests on recovered rock core samples, and DPL tests in the soft, wet organic bog areas



Laboratory testing of representative rock and soil samples to characterise both foundation and potential borrow materials.

Waste Rock Facility and Associated Structures 

Six geotechnical drillholes with depths ranging from 70 to 120 m.



Six hydrogeological drillholes with depths ranging from 50 to 140 m.



Geomechanical logging and lithological descriptions of the cored materials from these holes.

Page 272




Installation and completion of eight open standpipe piezometers (four singles and two multi-level open stand pipe in selected geotechnical holes.



Excavation of 26 test pits.



Field testing including SPT testing, down hole permeability tests in soil and rock, point load tests on recovered rock core samples, and DPL tests in the soft soils areas.



Laboratory testing of representative rock and soil samples to characterise both foundation and potential borrow materials.

Process Plant 

Three geotechnical drillholes with depths ranging from 30 to 65 m



Geomechanical logging and lithological descriptions of the cored materials from these holes



Installations of three open standpipe piezometers



Field testing including SPT testing, down hole permeability tests in soil and rock, point load tests on recovered rock core samples, and DPL tests in the soft, wet organic bog areas



Excavation of 36 test pits



Laboratory testing of representative rock and soil samples to characterise both foundation and potential borrow materials

18.2.2 Open pit geotechnical investigations, design parameters and slope angles The final Constancia pit area will measure 1.6 km by 1.3 km with a maximum depth (difference between the highest point of the pit rim and the lowest point within the pit) of 570 m. About 85% of the rock consists of intrusive rocks (chiefly porphyritic monzonites with minor components of diorite and andesite); 5% consists of sandstones and 10% of skarn and marbleised limestones. The structural map prepared for the pit is presented in Figure 18.18.

Page 273


Figure 18.18 Pit Structural Map

Page 274


Structurally the Constancia pit area is controlled by four major structural systems expressed as regional faults and local faults, some of which follow regional trends. These structures are significant because they directly influence the quality of the rock mass and therefore the stability of the pit walls, and the allowable slope angles. The four main fault systems are indicated in Table 18.17. Table 18.17 Major Structures System in Constancia Pit Area System

Average Strike

Average Dip

System 1

N 11º W

65° SW

System 2

N 56º E

69° SE

System 3

N 80º W

79 ° SW

System 4

N 06º W

65° NE

As noted earlier, rock fracture measurements were obtained from geologic surfaces and cell/window mapping in road cuts and outcrops, and from oriented core holes drilled specifically for geotechnical purposes. Geotechnical data relating to fracture intensity and other parameters were collected from the ore reserve characterisation holes but these holes conveyed no fracture orientation data. Data were identified as major structures (faults significant enough to be shown on surface geologic mapping), minor structures, and individual discontinuities (joints, bedding planes, and shears) and were identified separately in the database. The fracture orientation data were adjusted for orientation bias (Terzaghi, 1965) resulting from the alignment of the source of the observations (for example, drillhole orientation or cut slope azimuth). A geotechnical database was established based on the mineral exploration drilling conducted by Norsemont, dating back to 2005, and the 2008/9 geotechnical site investigation results. The geotechnical database includes details of alteration, oxidation, lithology, rock strength, and rock fracturing characteristics including spacing, roughness, infilling, Rock Quality Designation (RQD) and continuity (or persistence) estimated according to the fracture type and infilling, in accordance with ISRM recommended procedures (1981). All of the pit wall slope stability analyses assumed that the water table would be lowered behind the pit walls by perimeter dewatering wells and in-pit dewatering holes. This is consistent with the pit dewatering plan and allows the pore water pressures to be neglected. Unit weight data and classification parameters assumed damp/wet conditions. The rock mass parameters were compiled into estimates of rock mass quality through the Rock Mass Rating (RMR) system of Bieniawski (1989), and further into the Geomechanical Strength Index (GSI); (Hoek, Marinos, and Benissi, 1998; and Hoek, Wood and Shah, 1992). The RMR classification data by core run show that 58% of the values obtained correspond to “poor” rock, 38% to “fair” and 4% to “good” rock. In the Constancia pit area, the rock mass ratings are controlled by weathering, alteration, fracture intensity, and the filling material in fault zones, but are not greatly influenced by the lithology. From the rock mass quality ratings, the geotechnical conditions were discretised into six major geotechnical units, or domains, as follows:

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

Geotechnical Unit I – overburden consisting of Quaternary deposits formed mainly by glacial till and areas of bog material or wetlands. The maximum thickness of this unit is approximately 40 m in the pit area, but is typically much less.



Geotechnical Unit II – residual soil to extremely weathered rock with intense oxidation, extremely weak to very weak and fragmented: characteristically similar to fault zones bearing gouge or intensely fractured/crushed rock fillings.



Geotechnical Unit III – highly weathered to fair rock with moderate oxidation, weak and highly fractured. GSI values are typically in the 30-40 range.



Geotechnical Unit IV – moderately weathered rock, medium strong to strong but fractured. GSI values typically range between 40 and 50.



Geotechnical Unit V – slightly weathered rock, strong to very strong, slightly fractured. GSI values are typically in excess of 50.



Geotechnical Unit VI – slightly weathered to fresh rock, strong to very strong, slightly fractured. These rock masses approximate the strength of the intact rock. GSI values are typically in excess of 60.

The investigations showed that for about three-quarters of the ultimate pit slope perimeter, slope stability is influenced chiefly by discrete major structures and block faults that intercept broad zones of more competent rock. These areas were analyzed by considering the discrete structures kinematically, and by performing continuum, limit-equilibrium stability analyses on sections representing the discrete structures and the intervening rock mass zones. The remainder (1/4) of the ultimate pit slope areas are influenced chiefly by a more homogeneous condition of poor rock quality. These areas were analyzed by limit equilibrium and continuum modeling methods. In both cases, the applicable fault zone and rock mass strength conditions were characterised from rock quality values developed from core inspection and surface mapping, coupled with laboratory geotechnical testing on recovered fault gouges and rock discontinuities. The pit slope recommendations are associated with nine pit design sectors (Figure 18.19), which were established to group areas of the proposed pit having similar geometric, geological and rock mass quality characteristics. Each pit sector was characterised by one or more geotechnical units, each unit representing a region of generally consistent rock mass character. Major fault structures were considered as individual units. Minor fault structures, as well as alteration, oxidation, and lithology were included in the overall rock mass characterisation within a particular unit. Rock structure orientation data were applied for each design sector. For each design sector, the data included rock structural observations and fracture statistics from surface observations and oriented core holes pertinent to that particular sector. Therefore, the kinematic analyses reflect both surface and subsurface data for each sector. The recommended slope angle may be determined by the catch bench requirements (catch bench design) or by global stability (which considers the rock mass at the inter-ramp) or global (overall pit wall) scale. Catch bench integrity may be driven by kinematics, which identifies the distribution of localised discontinuities along which catch bench degradation may take place in toppling, plane shear, or wedge failure modes. Catch bench integrity may also be affected by rock mass failures responding to overall

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weak rock mass conditions. Methods used to determine the appropriate pit slope angles based on catch bench design or global stability included detailed kinematic stability at bench and inter-ramp scales; limit equilibrium analysis; probabilistic analysis by the limit equilibrium method; and stress analysis using finite element methods in certain design sectors. If the slope design is governed by global stability the resulting catch benches may be wider than the required 11.5 m minimum, assuming that the same bench face angle occurs. For purposes of checking the catch bench design, kinematic and limit equilibrium stability analyses indicate that a bench face angle of 65-70o is expected to be achievable, in most places, for the Constancia pit walls. A 30 m high double bench configuration, requiring catch benches at least 11.5 m wide (Ryan and Pryor, 2000) has been considered for pit development (the calculated minimum catch bench width includes an additional 1 m allowance for block loss at the crest).

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.19 Design Sectors and Proposed Inter-ramp and Bench Slopes

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The global stability or inter-ramp design may be affected by kinematics as well; in this case, however, consideration is given to major structures or very strong trends of localised discontinuities that may support the development of failure zones in plane shear, wedge, or passive block failure modes. The global pit slope stability was also checked by limit equilibrium techniques using Hoek and Brown failure criteria applied to characteristic geotechnical sections within each sector that include the various geotechnical domains and their rock mass properties. It was assumed that controlled blasting techniques and other measures would be employed to reduce rock mass disturbance from blasting and stress relief. In critical sections of design sectors I, II and VII, stress reduction analyses were carried out using the finite element method. The relationship between inter-ramp angle, overall angle, bench face angle, and catch bench width is shown in. Figure 18.20. Figure 18.20 Relationship Between Inter-ramp Angle, Overall Angle, Bench Face Angle, and Catch Bench Width

Summarizing the above, three types of stability analysis were carried out as appropriate to each design sector: 

kinematic stability analysis controlled by rock mass structures

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

inter-ramp or global slope stability analysis controlled by kinematic capability and rock mass strength



Probabilistic analysis for bench face stability.

The minimum acceptable factors of safety (FoS) estimated by the limit equilibrium method adopted for stability analyses are 1.2 under static conditions; and, in accordance with the Peruvian regulations, 1.0 (earthquake loading) using the peak ground acceleration for a 100 year return event. Based on a seismic risk study for the Constancia Project, a peak ground acceleration of 0.08 g was adopted for a return period of 100 years. According to the analysis procedures explained above, the recommended inter-ramp slope angle and bench angles are shown in Table 18.18. The pit slope angles for each design sector have been recommended in accordance with acceptance criteria for each of these design methods, taking into consideration the expected maintenance and operational impacts. The tabulated bench face angles are expected to occur with a calculated overall frequency of at least 65%. Since there is potential for catch bench failures and localised slope movement at the inter-ramp scale, clean-up, maintenance, and special attention to rock fall protection will be necessary. In general, a higher probability of failure may be tolerable where the slopes are low, where the mining configuration permits rock fall catchment, where the service time of the slope is short, or when important facilities such as pump stations or ramps are not vulnerable to the slope in question. Table 18.18 Inter-ramp Slope Angles – Constancia Open Pit OVERALL SECTOR

PIT SLOPE (º)

INTER-

BENCH

POTENTIAL FAILURE MODE

RAMP

CALCULATED

BENCH

BENCH

SLOPE

BENCH FACE

HEIGHT

WIDTH

(º)

ANGLE (º)

(M)

(M)

PLANAR

WEDGE

X

I

50

54

70

30

11.5

X

II

50

54

70

30

11.5

X

III

50

54

70

30

11.5

IV

48

52

70

30

11.5

X

X

V

50

54

70

30

11.5

X

X

TOPPLING

CIRCULAR

VI

45-47.5

48-52

65-70

30

13.5

X

VII

45

48

65-70

30

13.5

X

VIII

47

50

70

30

11.5

IX

45

48

65-70

30

13.5

X

X X

Recommended inter-ramp slope angles range from 48-54o depending on the design sector. Overall slope angles are the flatter of the angles resulting from kinematic or limit equilibrium analyses for the global stability and catch bench cases. Considering the current open pit configuration, the calculated probability of failure ranges between 8% and 27%, in about 77% of the bench face slopes. Based on the current geotechnical understanding of the Constancia pit geology, it has been calculated that a probability of failure of 55% pertains to the

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remaining 23% of the bench face slopes. In these areas additional geotechnical investigation needs to be carried out in the next stage of engineering design. 18.2.3 Geotechnical investigations at the TMF, PAG WRF and plant site Subsurface conditions at the locations of the TMF, PAG WRF, and plant site are closely related to the geologic units present, as well as the influences of erosion and glaciation. Bedrock at the site is mainly comprised of sedimentary rock of the Cretaceous period and intrusive rock of the Paleogene period. More recent soil formations are of the Quaternary period, and include organic bogs with some laccustrine deposits, alluvium and glacial till deposits. Figure 18.21 presents the locations and limits of the geologic units mapped within the project area, and includes a brief description of the material characteristics associated with each unit. Figure 18.21 TMF Geotechnical Investigation and Geologic Map

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The following summarises the more significant geologic units as they relate to the geotechnical subsurface conditions: 

Bog deposits are generally found in low-lying areas of the valleys and in valley depressions where standing water has historically accumulated. Associated deposits generally consist of saturated peats and organic-rich soils comprised of silts and clays. These deposits are not suitable as foundation or construction materials as the material has high compressibility and low strength.



Alluvium is generally located in the low-lying areas of the valleys in and around active rivers, streams and drainages. Alluvial deposits consist mainly of sub-rounded to rounded sandy gravels with silt. These materials are generally well-sorted and tend to be higher in permeability.



Glacial till deposits, largely in the form of lateral and end moraines, are located at higher elevations above the northwestern portion of the site where alpine glaciation occurred. Glacial till material consists of sandy and clayey gravels with cobbles. The materials vary in consistency from medium dense to dense, and are generally lower in permeability. These materials are generally carbonaceous having originated from limestone outcrops further to the northeast.



Intrusive Rocks of the Apurimac batholith consist of diorites, monzonites and monzodiorites. The most prevalent of these are diorites which are generally fine grained, and highly to extremely weathered at the ground surface, becoming less weathered at depth. In areas of extreme weathering, the diorite is completely weathered to fine sand and silt. This is common on the ground surface in the southern half of the TMF site.



Sedimentary Rocks of the Chilloroya Formation consist of medium-grained sandstone with interbedded siltstone and mudstone. Sandstone is found throughout the site and is generally moderately to highly weathered near the ground surface, becoming less weathered at depth. At higher elevations, where topography is steeper, the sandstone tends to be less weathered.

Under the site of the TMF, bedrock consists of diorite and sandstone. Figure 18.21 presents the TMF geotechnical investigation and geological map. The contact between these units extends from beneath the northwest side of the site to the southeast with the sandstone located to the north and the diorite located to the south. The boundary is just south of the main section of the proposed embankment dam, and will place the majority of the embankment on sandstone with the west end on diorite. The TMF site features two prominent south to north orientated valleys and in these valley bottoms the bedrock is overlain by bog deposits. The bogs are generally only a few meters deep but reach depths of up to 12 m in their centres. Current development plans for the TMF call for completely removing the bog deposits. It is estimated that a suitable foundation for the TMF embankment will be reached at depths of approximately 2 to 5 m below the existing ground surface and below the bottoms of the bogs. In the diorite, however, the bedrock appears to be extremely weathered to depths greater than 20 m in some areas. Further investigations will be required to evaluate the need for removing the diorite to significant depths under the embankment. Bedrock under the site of the PAG WRF also consists of diorite and sandstone; however, a large portion of the site is overlain by consolidated glacial till deposits and/or outwash deposits which vary in thickness from between 25 and 95 m. In the lowest lying areas of the site, bog and alluvial soils were encountered. Unsuitable foundation soils are currently planned for removal from beneath the critical

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slopes of the PAG WRF in order to maintain adequate stability of the slopes. Excavation depths in these areas are anticipated to be between 0.5 and 1.0 m, and further investigations will be required to confirm these depths. At the site of the process plant, foundation materials consist of sandstones, diorites and monzonites. Competent rock (RMR values greater than 30) was found at depths between 7 and 10 m under the locations of the main structural components of the plant. Residual soils and very poor quality rock overlying the more competent bedrock will most likely need to be removed during construction. 18.2.4 Borrow materials During the geotechnical investigations, potential borrow sources were identified for construction materials. Table 18.19 provides a summary of construction materials, potential borrow sources, locations of borrow sources and the processing (if any) that will be required of the materials. Table 18.19 Construction Materials Summary Construction Material

Potential Borrow Source

Location

Random/common fill

Sandstone, intrusive rocks

Throughout Project site

Structural fill

Intrusive rocks

Pit, TMF basin

Drainage material

Alluvium / intrusive rocks

Chilloroya River / TMF basin

Processing Selective excavation and moisture conditioning Selective excavation and placement in controlled lifts Crushing and/or screening Selective excavation,

Low permeability core/ sub-base material

Glacial till

PAG WRF footprint

screening, moisture conditioning and placement in controlled lifts

Filter material

Intrusive rocks

Wearing course

Sandstone

Rip-rap

Intrusive rocks

Pit, TMF basin, throughout

Crushing and/or screening

Project site

and washing

Throughout Project site

Moisture conditioning

Pit, TMF basin, throughout

Crushing and/or screening,

Project site

interlocked placement

Additional investigations and laboratory testing will be necessary during the next stage of the design to fully characterise the suitability of these materials and their available quantities. Current plans call for the non-PAG waste rock mined from the pit to be used for constructing a large portion of the TMF embankment, as well as constructing haul roads and operator access roads at the site. Approximately 55 Mt of this material will be available over the life of the project, and, based on the mine plan and a mass balance calculation, sufficient quantities of this material will be available for other uses after the initial construction stages of the project. 18.2.5 Natural hazards and slope stability The slopes in the area of proposed project structures appear stable and display no observed risk of mass displacement, sliding or landslides. However, in isolated areas of the project away from proposed

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project structures, some sinkholes were observed, providing evidence of historic dissolution and erosion. These observed features pose no potential risk to the proposed project structures. 18.2.6 Seismic risk analysis The site is located in the Province of Cusco, which has experienced historical seismicity related mostly to tectonic activity associated with an extensional zone of the subducting Nazca plate passing under the South American plate. The province is located within Zone 2 (intermediate) of the 2003 Peruvian Seismic Resistance Code. Deterministic and probabilistic analyses were carried out for the site. The deterministic analysis defined the potential source zones, types and properties of earthquakes that could be experienced at the site based on its seismic setting, and included defining potential Maximum Credible Earthquakes (MCEs) that the site could experience. The probabilistic analysis produced a statistical relationship between peak ground acceleration (PGA) and return period for earthquakes that the site could experience based on the recorded seismic history. The deterministic analysis identified two potential MCEs, as follows: 

Magnitude 8.0 deep intraplate subduction earthquake, producing a mean plus one standard deviation PGA at the site of 0.38 g.



Magnitude 7.0 shallow crustal earthquake artificially located 25 km directly below the site, producing a mean plus one standard deviation PGA at the site of 0.25 g.

The first event is considered to represent the Maximum Design Earthquake (MDE) since it has been calculated to produce a significantly higher PGA. As a higher magnitude deep intraplate event, it would also be expected to produce a greater number of significant cycles during shaking and have a longer duration of shaking than a shallow crustal event. The probabilistic analysis gave PGAs at the site, in the free field, for selected return period events, as follows: 

50 year event = 0.06 g






475 year event (10% chance of exceedance in 50 years) = 0.15 g






2475 year event (2% chance of exceedance in 50 years) = 0.24 g






10 000 year event = 0.34 g

The above listed PGAs are associated with a 63% probability of exceedance, which is representative of the case where the design life of the facility in years equals the return period of the event.

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18.3

HYDROGEOLOGICAL STUDIES

18.3.1

Introduction

Hydrogeological studies were conducted for the open pits, TMF and PAG WRF areas. Studies consisted of springs mapping and inventories; mapping of karstic limestone areas; drilling and installation of multi-level piezometers; hydraulic testing of soils and bedrock; baseline monitoring of ground water quality, spring flows and water levels; installation of large diameter wells and pumping tests within high permeability fault zones; water balance and recharge calculations using GIS; threedimensional numerical modelling of the open pit, PAG WRF and TMF; impact analysis; and dewatering cost estimations. The following section outlines the key hydrogeological issues, the conceptual ground water flow model, pit inflow estimations and hydraulic control measures, and seepage estimations from the mine waste facilities based on information obtained to date. The key hydrogeological issues investigated during the EIA and feasibility studies are listed in Table 18.20 with related mitigative measures. Table 18.20 Key Hydrogeologic Issues and Mitigative Measures Key Hydrogeologic Issues Mitigative Measures Potential reduction of baseflows and/or water

Ground water from the dewatering wells with

quality in the streams and springs within hydraulic

acceptable water quality will be pumped to the

influence of the open pit.

affected areas as needed during the dry season.

Potential for reduced ground water supply and

Fractured bedrock aquifers could be developed

failure to supply the mill with the minimum water

beneath the Chilloroya River valley as a back-up

requirements if drought conditions occur during the

water supply. These investigations will be carried

mine life.

out in a subsequent stage of design.

Potential for poor drainage of the pit slopes within

Depressurisation of pit walls can be overcome with a

the less fractured or less permeable rock masses.

network of horizontal drains or an increased number of in-pit vertical wells within the low-permeability rock masses.

Potential for high inflows to the open pit from

Increased number of pumping wells and pumping

transmissive fault zones.

rates will resolve this; however, increased pumping will increase the hydraulic impact on surface water resources along fault zones.

Potential water quality impacts to the Chilloroya

Seepage bypass from the waste facilities is expected

River from mine waste facilities.

to be low, with minimal to negligible environmental impacts. Ground water monitoring wells located immediately downgradient of the facilities should detect water quality impacts long before impacts can occur to surface water or springs. Mitigative measures will consist of interception and treatment.

Potential for poor water quality seepage from the

Ground water seepage surrounding the open pit will

open pit during post mining pit flooding.

be monitored throughout the mine life initially after closure and will be collected and treated if necessary.

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18.3.2

Hydrogeological investigations and results

Open pit The hydrogeological investigation at the open pit included the following: 

Quarterly hydrogeological mapping of the open pit area and vicinity



Ground water sampling and laboratory analysis of metals, anions, nutrients and bacteria. The regional ground water is generally of good quality with occasional elevated Al, As, B, Fe, Mn, Na, Ni, Pb, and SO4 near the mineralised areas



1160 m of diamond drilling and an equivalent length of packer (hydraulic) testing within the bedrock units



Installation of 40 standpipes and multi-level piezometer wells, and hydraulic testing on the completed piezometers, 10 of which were installed within the plant site area



Monthly monitoring of water levels in the piezometric network. The seasonal variation in phreatic water level is low, typically between 1 and 5 m



Installation of two large diameter (8 inch) wells to depths of 162 m and 170 m along the alignment of the Yanak Fault on the north and south side of the pit, respectively



Pumping tests on the large diameter wells to assess bulk hydraulic parameters of the fault zone



A cross-hole dye tracer test using Uranine within the Yanak fault.

Five hydrostratigraphic units were identified for the open pit: 

Glacial moraine deposits are present to the north, east and south of the pit area, but not in the pit area itself. The geometric mean conductivity value for this unit was low (8.4x10-8 m/s), and it is considered a hydraulic barrier to recharge.



Apurimac intrusive diorite and monzonite are the dominant bedrock types in the pit area. The geometric mean hydraulic conductivity for both the diorite and monzonite was 2.4x10-7 m/s. These two lithologies were combined as one hydrostratigraphic unit since they have similar properties.



The Chilloroya Formation sandstone/siltstone is present along the southern wall of the proposed pit area. The hydraulic conductivity of this formation is typically low, at 2.3x10-7 m/s throughout the project area, and about an order of magnitude higher within the pit area.



Arcurquina limestone is present to the north and east of the open pit. The karstic limestone has been identified as a recharge unit in the overall hydrogeological model of the Constancia area. Karstification is common to estimated depths of 200 m within the ridge areas. Limestone is not intersected by the ultimate pit shell.



Major fault zones are significant to the ground water regime throughout the Constancia project area. The Yanak fault is of particular significance in the Pit area. The fault width varies from about 20 to 70 m with an estimated hydraulic conductivity of 5x10-5 m/sec. The fault appears to be both a barrier and conduit to ground water movement with strong artesian ground water pressures (~20 psi) and flows up to 20 L/sec from drillholes.

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Tailings management facility The hydrogeological investigation of the TMF area included the following components: 

Quarterly hydrogeological mapping of the TMF area and vicinity



Spring surveys of location, flows and water quality



Ground water sampling and laboratory analysis of metals, anions, nutrients and bacteria



Geophysical surveys within the Chilloroya Valley consisting of 10 electrical resistivity soundings to target drilling locations, and approximately 1000 m of seismic refraction survey to define the bedrock profile



Over 1200 m of diamond drilling and an equivalent length of packer (hydraulic) testing within the bedrock units. The dominant rock types in the TMF area are sandstone and diorite



Installation of 18 standpipes and multi-level piezometer wells and hydraulic testing on the completed piezometers



Installation of two multi-level piezometers to depths of 38 m and 39 m within valley infill of the Chilloroya valley immediately downstream of the proposed TMF site



Monthly monitoring of water levels in the piezometric network.

Four hydrostratigraphic units were classified from the lithologies and conductivities found within the TMF area, as follows: 

Valley infill deposits, consisting of alluvium and possibly glaciofluvial deposits, have a measured thickness of 12 m. The geometric mean hydraulic conductivity was estimated at 6x10-6 m/s.



The Apurimac intrusive diorite is present over the southern portion of the TMF area with a hydraulic conductivity of 1.3x10-7 m/s.



The Chilloroya Formation sandstone/siltstone is present in the lower parts of the TMF area to the northeast and underlies much of the dam foundation. The formation has a hydraulic conductivity of approximately 2.3x10-7 m/s.



Fault zones were considered as separate hydrostratigraphic units. The hydraulic conductivity of the fault zones is approximately 10 times higher than the bulk hydraulic conductivity of the adjacent relatively unfractured blocks.

Waste rock facility The hydrogeological investigation at the PAG WRF area included the following components: 

Quarterly hydrogeological mapping of the PAG WRF area and vicinity



Spring surveys of location, flows and water quality: a total of 19 springs were mapped within the limestone formation to the northeast of the facility layout. No ground water springs were identified within the PAG WRF footprint



Ground water sampling and laboratory analysis of metals, anions, nutrients and bacteria. Good quality water was typically found, except for elevated sulphate within Quebrada Cunahuiri



Over 1015 m of diamond drilling and an equivalent length of packer (hydraulic) testing within the bedrock units. The dominant bedrock in the PAG WRF is sandstone of the Chilloroya Formation

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

Installation of 16 standpipes and multi-level piezometer wells and hydraulic testing on the completed piezometers



Monthly monitoring of water levels in the piezometric network



15 permeameter tests within the Quaternary moraine deposit which covers over 50% of the proposed PAG WRF area.

Six hydrostratigraphic units were identified within the PAG WRF area. 

The alluvial material has a maximum measured thickness of 27 m along the axis of Quebrada Cunahuiri. The average hydraulic conductivity of the alluvium is approximately 2.5x10-5 m/s.



Consolidate glacial till or outwash covers most of the planned PAG WRF area with a thickness of up to 33 m. Permeameter measurements of hydraulic conductivity in the moraine are considered representative, giving values in the order of 1x10-7 m/s.



The Chilloroya Formation sandstone/siltstone is present over most of the PAG WRF area. It is underlain by intrusives at a depth of 72 m in the northeastern part of the PAG WRF. The Chilloroya sedimentary unit has a hydraulic conductivity of approximately 6x10-7 m/s.



The Apurimac intrusives are believed to underlie the sedimentary rocks at depth. The intrusive unit has an average hydraulic conductivity of 5x10-7 m/s.



Arcurquina limestone is present to the east and north of the PAG WRF. It has not been found in outcrop or drilling within the PAG WRF footprint, and is unlikely to have a local effect on the WRF ground water regime.



Fault zones were found to have properties similar to those found in the tailings facility.

18.3.3

Numerical seepage modelling and affects analysis

The hydrogeological regime beneath the Constancia project area was modelled with FEFLOW, a commercially available finite element code for ground water and unsaturated zone flow. The modelling comprised four phases, as follows: 1.

Steady state, pre-mining calibration phase

2.

Steady state and transient models of pit operation and closure

3.

Steady state models of the tailings impoundment

4.

Steady state models of the waste rock dumps.

The base model and the pit model have ten layers. Additional layers were added, as required, to simulate the tailings impoundment and the waste rock dumps. The model domain covers the entire mine site and extends outwards to significant streams on its periphery which serve as boundary conditions. The model hydraulic conductivities were calibrated to water levels measured in site monitoring wells. A measure of the goodness-of-fit of the model is the normalised root mean squared error. As a rule of thumb, a value of less than 10% is considered acceptable for groundwater flow modelling. The Constancia model has a normalised root mean squared error of 3.6%.

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18.3.4

Open pit seepage and dewatering

The open pit is expected to have final dimensions of approximately 1.5 km x1.3 km. The base of the final pit will be at RL 3876 m and about 125 m below the Chilloroya valley. The north and south highwalls are expected to have heights of approximately 575 m and 375 m, respectively. The phreatic surface ranges from approximately RL. 4400 m to RL 4225 m; significant ground water inflow to the pit will not occur until mining extends below these elevations. It is estimated that between 10 to 30% of rainfall recharges the bedrock system in the local catchment area, depending on rock type, degree of fracturing and karstification. Field investigations and numerical modelling suggests that there is sufficient ground water supply from a combination of pit dewatering wells and groundwater supply wells, thus negating the need for a surface water reservoir. The exception to this will be during the start-up of operations when water temporarily stored in the TMF will be utilised. Pit sumps and deep dewatering wells have been incorporated into the dewatering system design for the feasibility study. A phased sump design will be required as the pit develops. Temporary sumps will collect both surface water and ground water along the ramps and at the pit base, increasing in size as the pit expands. Dewatering wells must be activated on start-up and run continuously throughout the mining and milling operations. The pit dewatering system will consist of eight dewatering wells situated around the perimeter of the open pit targeting prominent fault zones, and ten in pit wells. An additional four dewatering wells in vicinity of the plant may be required to assist with the supply of water at startup. Numerical dewatering simulations indicate localised drawdown in the early years within the intrusive areas, with a preferential drawdown extending to the north along the Yanak Fault. As the pit deepens, its hydraulic influence will expand to an estimated maximum diameter of 5 km by 6 km, using 1 m head drop as the cut-off point. Dewatering rates are sensitive to fault permeability and recharge rates, and are limited by well depths of 300 m; however, dewatering rates are considered to be easily manageable. Deeper dewatering wells are possible if larger drilling rigs are imported, with costs increasing substantially. Due to the high permeability of the fault zones, the need for pit wall depressurisation ahead of the pit advancement using in-pit horizontal drains or vertical wells is not anticipated; however, contingencies for horizontal drains have been made. Based on the drilling and mapping within the open pit area, it is unlikely that the pit shell will come in contact with any of the karstic limestone. Hydraulic connection between the pit and the karstic aquifers depends largely on the permeability of the rock mass between the pit shell and the limestone, and the permeability of the fault zones. Hydraulic connection with the karstic aquifer is significantly increased with the presence of permeable fault zones. Two potential adverse impacts could occur to the limestone aquifers and related springs – (1) baseflow reduction to Sacrani Valley and Caja Pampa Spring located within the Chilloroya Valley west of the open pit, and (2) baseflow reduction to springs and wetland areas within two of the three lake basins situated north of the mine within the limestone terrain, which are used for community irrigation water supply. The Yanak Fault passes immediately to the east of the three large northern lakes located between 2 km and 5 km north of the open pit. The lakes are bounded to the east and west by splays of the Yanak fault and on the north and south by E-W trending faults. These factors make these lakes particularly vulnerable to pit dewatering. Numerical modelling indicates a high probability that the two closest lakes will experience reduced inflows originating from direct ground water discharge to the lake and/or reduced baseflow from upstream channels. This assumes that the fault zone is permeable over its

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entire length; however, the confidence of this assumption is low. Excess water from the dewatering operations will be used to mitigate any potential impacts on community water supplies. Once mining has ceased, ground water levels will recover with time and stabilise at a new elevation which will be controlled by the spill point at southwest side of the pit at approximately RL 4225 m. Preliminary estimates are that the pit will fill to the spill point at RL 4225 m in approximately 100 years. Post-rebound outflows at the spill point are estimated to be between 10 and 15 L/s. The potential exists for acid rock drainage (ARD) from the pit walls, but this will be addressed by in situ neutralisation as discussed in Section 18.12.4. 18.3.5

Tailings management facility seepage

The proposed TMF will be located within a small catchment on the south side of the Chilloroya River, approximately 3 km southwest of the open pit. The numerical model was used to evaluate seepage losses from the TMF. The tailings deposit was modelled as a two-zoned material: an upper layer with a hydraulic conductivity of 1x10-7 m/s, and a lower layer with a hydraulic conductivity of 5x10-8 m/s. The dam was modelled with a hydraulic conductivity of 1x10-4 m/s, and the dam core was treated as a discrete feature within the model with a hydraulic conductivity of 1x10-8 m/s. During operations, the entire tailings surface can be assumed to be completely saturated. For the ultimate TMF configuration, the pond surface was set to an elevation of 4125 m, 5 m below the surface of the tailings. An initial model run was made without an underdrain system or geomembrane liner, and the seepage loss from the impoundment was estimated at 25 L/s. The predicted seepage rate is dependent upon the hydraulic conductivity of the tailings and the foundation materials. The addition of a geomembrane under the tailings results in a reduction in seepage losses of 22 L/s. When both an underdrain and geomembrane are simulated, the estimated uncontrolled seepage losses drop to 4 L/s. This modelled condition is representative of the TMF design case. At post-closure, the tailings surface will be covered and vegetated. This will have the result of reducing the recharge rate over the tailings. 18.3.6

Waste rock facility seepage

The PAG WRF will be developed within the Cunahuiri Valley situated immediately south-southeast of the proposed open pit. The approximately 100 m deep, U-shaped glacial valley will receive all PAG and unsuitable waste rock over the mine life. Three conceptual seepage paths from the PAG WRF were identified and assessed with the model: (1) south to the Telaracaca Valley; (2) north to the open pit; and (3) west along the base of the Cunahuiri Valley in the direction of the Chilloroya Valley. The hydrogeology of the PAG WRF was evaluated for operation and post closure scenarios. In both analyses, the ultimate PAG WRF configuration was modelled, with an assumed hydraulic conductivity of 1x10-7 m/s. During operations, the waste rock system and underlying foundation material are expected to be unsaturated. The rate of seepage loss is dependent upon the assumed recharge rate through the facility and the hydraulic conductivity of the dump material. The recharge rate to the facility has been assumed to be 26% of the annual precipitation rate, or approximately 234 to 260 mm/year, depending on the elevation of the dump surface. This translates to an annual average recharge rate of

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approximately 17 L/s through the PAG WRF. During operation, pumping of the dewatering wells and the presence of the open pit will prevent any losses of waste dump-derived seepage to the environment. Any infiltration not captured by the drain system under the PAG WRF will be captured by the open pit and dewatering system during operation and while the open pit is flooding. At post-closure, it is assumed the final cover will reduce the recharge rate to a maximum of 16% of the annual average precipitation. If the underdrainage system continues to work in perpetuity, the numerical model predicts negligible seepage losses from the facility. In the event that the underdrain system fails after closure, the numerical model predicts seepage losses to the Chilloroya Valley of approximately 2.8 L/s 18.4

PROCESS PLANT DESIGN

18.4.1

General layout and description

The layout of the mine, processing plant, waste dumps, and tailings facilities are shown in Figure 18.22. The process plant is located to the west of the open pit mine, with the TMF located approximately 3 km south of the plant. The PAG WRF is located adjacent to the pit. Road access to the process plant and pit is from a connection to the existing north-south road 2 km to the western side of the plant. The construction and operations accommodation camp is located in walking distance of the process plant. The mine is sited in hilly terrain. The plant site has been selected as being the closest site available to the mine that is both economical and practical to develop, but significant earthworks are required to establish a multi-tiered pad. Major plant areas are all located on in-situ material, i.e. constructed on cut. Some general buildings are constructed on engineered embankment. A suitably sized construction laydown area is provided for the establishment of management offices and construction requirements. Power supply is provided to the mine from the existing network via a 138/220 kV transmission line from the Tintaya substation. The crushing facility is benched into the hillside to minimise expensive ROM pad construction. The layout allows the crushing facility to be free-draining ensuring practical and economical maintenance. The mine office and workshops are located immediately to the north of the process plant, and is located outside the 500 m designated blast zone from the pit boundary. Significant costs would have been required to expand the existing process plant site to accommodate these buildings nearer to the plant.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.22 Site Layout

18.4.2

Facility description

The plant is designed to process a maximum of 76 000 t/d of ore (25.26 Mt/a at 91.3% plant availability) from the Constancia and San José ore bodies. It will operate 52 weeks per year, 7 days per week, 24 hours per day, at 91.3% availability. This throughput will reduce to 45 480 t/d as the ore properties change (harder ore expected from Year 5 onwards). The expected mine life is 15 years. Annual concentrate production rates ramp up from 350 000 t in the first year to a peak of 450 000 t in Year 3. Production then drops to around 300 000 t/a until Year 10, after which it falls to 200 000 t/a and below until mine closure in Year 15.

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Annual production rates are 70 533 t/a of copper metal and 54.5 t/a of silver metal contained in copper concentrate. Molybdenum concentrate production ramps up from 2400 t in Year 1 to a peak of 4800 t/a in Year 3. It then fluctuates between 2500 and 3000 t/a until another high is reached in Years 9 and 10, after which it drops to 2000-2500 t/a until closure. The primary crusher, belt conveyors, thickeners, tanks, pebble crushers, flotation cells, ball mills and various other types of equipment will be located outdoors without buildings or enclosures. To facilitate the appropriate level of operation and maintenance, the SAG mill (semi-enclosed), molybdenum plant, filters and concentrate storage will be housed in clad structural steel buildings. 18.4.2.1

Processing plant – general

The general layout of the processing plant is shown in Figure 18.23. Figure 18.23 Processing Plant Layout

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The processing plant has been laid out in accordance with established good engineering practice for traditional grinding and flotation plants. The major objective is to make the best possible use of the natural ground contours to minimise pumping requirements by using gravity flows and also to reduce the height of steel structures. To optimise the cost of the major footings for the SAG mill, the height of the SAG mill above grade has been minimised by locating the cyclone feed sump as low as possible. The mill cyclones have been located so that the cyclone overflow can gravitate into the rougher conditioning tank. At the other end of the flotation bank, the copper tailings thickener has also been located to facilitate gravity flow and eliminate the requirement for another large set of pumps. Due consideration has been given to the layout of the molybdenum plant to facilitate good housekeeping and occupational health and safety requirements. The copper and molybdenum plants are independent of each other and separated by a reasonable distance. Further, the molybdenum concentrator is sited downwind of major high occupancy plant areas (prevailing wind on site is from the north). Similarly, the reagents area can release hazardous gases and so is located downwind of high occupancy plant areas, but close enough to the flotation area to avoid long runs of piping. Rainwater runoff and general spillages from the processing plant are fully contained and directed to the dirty water sediment pond. Again, the natural fall of the site is used to direct flow to this pond, located at the western end of the site at the lowest point, to avoid the need for pumps. Wherever possible, major substations and motor control centre (MCCs) have been located as close as possible to the drives to minimise cable run lengths. 18.4.2.2

Crushing

Primary Crushing and Conveying The ore is delivered to the ROM pad by haul trucks (nominally 220 t capacity) and is dumped directly into the primary crusher dump pocket. ROM ore top size is 1 m. The ROM pad is approximately 175 m long and 125 m wide. It provides haul truck access to the primary crusher dump pocket from two sides. Provision will be made on the ROM pad for two emergency ROM stockpiles, each with areas of 2000 m2, which would provide a total emergency capacity for approximately 24 000 t of ore, sufficient for 5-6 hours of primary crusher operation. The emergency stockpiles can be loaded into the crusher dump pocket by FEL. The potential exists for additional storage areas for “off-spec” ore to be located away from the ROM pad. A design feed rate of 5573 t/h (133 752 t/d) is nominated for the crushing plant: an availability of 80% is assumed. The crushed coarse ore stockpile will allow the plant to continue to run during crusher outages. A 60” x 89” primary gyratory crusher is fed by rear-dumping from two dump points by trucks, or by FEL from a stockpile. The third face at the crusher dump pocket is fitted with a drive-up ramp

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allowing bobcat access into the dump pocket for clean-out prior to maintenance work being undertaken or for clearing blockages during normal operations. A static rock breaker is situated at the feed bin to break oversise rocks and to clear blockages. Dust suppression water is sprayed within the dump hopper during the tipping of each truck load. The ore is crushed using a closed side setting of 175 mm. Expected maximum lump size is around 250 mm, with a P80 of 175 mm. The primary crusher is equipped with a 600 kW drive motor. The crusher discharges into a pocket located directly above a 2.4 m wide apron discharge feeder which transfers crushed ore to a 40 m long by 2.1 m wide crusher transfer belt conveyor, which in turn feeds onto the 1.8 m wide stockpile feed conveyor at up to 5573 t/h. The crusher transfer conveyor acts as a sacrificial conveyor to reduce wear on the much longer stockpile feed conveyor and also protects against damage to the longer belt. A metal detector is fitted above the crusher discharge conveyor and a self-cleaning magnet is suspended at the transfer point from the conveyor to the stockpile feed conveyor. Tramp metal is discharged to a collection bin. The stockpile feed conveyor is equipped with a weightometer. A dust suppression system is provided for the primary crushing area. The primary crusher area is supervised from a local control room. The crusher MCC is located on ground level to house the electrical and instrumentation ancillaries. It is supplied from the primary crusher substation. Crusher lubricant and hydraulic fluid are water-cooled. Closed circuit cooling water will be sourced from a cooling tower in the stockpile and pebble crusher area. An air blower provides sealing air to prevent dust entering the crusher eccentric in the bottom shell, via the dust seal chamber. A 250 t crawler jib crane is retained on site for use in crusher maintenance. An air compressor supplies crusher air services. Water is removed from the area by a sump pump. Crushed Coarse Ore Stockpile The stockpile feed conveyor has a design capacity of 5573 t/h, and discharges crushed coarse ore onto the open stockpile. The stockpile has a live capacity of 50 278 t, or approximately 16 hours of plant running time. Three 1.8 m wide variable speed reclaim belt feeders, with design capacities of 1895 t/h each, reclaim ore from beneath the stockpile and discharge onto the 1800 mm wide, 400 kW, SAG mill feed belt conveyor. A tunnel houses the reclaim feeders and the loading end of the SAG mill feed conveyor. A smaller personnel escape tunnel is also provided. The tunnel is ventilated by fans. The required air, water, and fire suppression services will be run in the tunnel. Eight hoists are provided for maintenance of the belt feeders and the portion of the SAG mill feed conveyor inside the reclaim tunnel. A dust suppression system is provided for the pebble crusher discharge loading points. Water is removed from the reclaim area by a sump pump. The SAG mill feed conveyor is equipped with a reclaim weightometer, pebble crusher discharge loading points, and a SAG mill feed weightometer. The SAG mill feed conveyor design capacity for combined

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crushed ore feed and crushed mill pebbles is 4737 t/h (3158 t/h of crushed ore and 790 t/h of recycled pebbles). SAG mill grinding media balls (125 to 150 mm nominal) are added to an 872 t SAG mill ball storage bin by FEL. A variable speed SAG mill grinding media feeder feeds balls onto the coarse ore on the SAG mill feed conveyor, at a controlled rate. 18.4.2.3

Grinding

Grinding reduces the particle size of the crushed ore to a target grind size of 106 µm, suitable for flotation. The SAG/ball/crush (SABC) grinding circuit contains a variable speed SAG mill, a SAG mill discharge screen, two pebble crushers, two cyclone feed pumps, two cyclone clusters, and two fixed speed, dual-pinion ball mills. The SAG mill is equipped with a variable speed 21.5 MW gearless mill drive motor. The 11.6 m dia, 7.9 m EGL (38’ x 26’) mill shell is fitted with a disc brake. Each of the two 7.3 m dia, 12.2 m EGL (24’ x 40’) fixed-speed twin-pinion ball mills is driven by two 6.5 MW motors with reduction gearing and an inching drive. Each of the two pebble crushers is equipped with a 600 kW drive. Each of the two cyclone feed pumps is equipped with a 1.5 MW drive. The variable speed SAG mill gearless mill drive motor is started and controlled by a proprietary “Ehouse”. The four fixed speed motors driving the two ball mills are started by conventional liquid resistance starter circuits. The mill lubricants and “E-house” components are water-cooled. Metallic liners are used in both the SAG mill and the two ball mills. Mill liner handling tools, mill liner bolt tools, monorails, and mill feed chute trolleys are provided for mill liner maintenance. The grinding circuit is fed from both the coarse ore stockpile and the two pebble crushers to the SAG mill feed chute, via the mill feed conveyor. The expected maximum lump size is around 250 mm with a P80 of 175 mm. The SAG mill grinds a nominal 3158 t/h of fresh crushed ore from the stockpile and a nominal 790 t/h of recycled ore pebbles. The pebbles will be crushed and recycle through the SAG mill. Ore entering the SAG mill feed chute is combined with process water, forming the feed to the SAG mill. Size reduction is realised by grinding within the mill with the charge and steel balls. The SAG mill slurry exits through the discharge grate (25 mm aperture) with pebble ports (60 mm) and passes over the 3.6 m x 7.3 m SAG mill discharge vibrating screen which has a 15 mm aperture. A trommel screen is not used. The 15 to 60 mm ore pebbles from the oversize of the SAG mill discharge screen are recycled to the 1300 t pebble crushing bin, via the 1421 t/h, 1200 mm wide pebble conveyors, fitted with weightometers, metal detectors and tramp metal magnets. SAG mill ball scats and tramp metal are separated from the ore pebbles and sent to either of the two pebble tramp bins. Excess pebbles can overflow from the pebble crushing bin to a pebble crusher overflow bunker. Pebbles may be bypassed around the pebble crushers to the SAG mill feed conveyor.

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Two 2.4 m wide pebble crushing belt feeders split the feed of ore pebbles from the pebble crushing bin to the two pebble crushers located above the SAG mill feed conveyor where it exits the stockpile tunnel. The pebble crushers are cone type crushers producing a 13 mm nominal product. Pebble crusher discharge is returned to the SAG mill feed conveyor. Crusher lubricant and hydraulic fluid are cooled by oil-to-water coolers. An overhead crane is provided for pebble crusher maintenance. Water is removed from the pebble crushing area by a sump pump. SAG mill discharge slurry passes as underflow through the vibrating screen and enters the SAG mill discharge hopper, together with the undersize discharge from both ball mill trommel screens. Dilution process water is added, before the combined slurry (SG 1.64) is pumped at 9408 m3/h (maximum) by the parallel circuit, cyclone feed pumps, to two hydro-cyclone clusters for classification. Each of the cyclone feed pumps has a 1.5 MW drive motor. A spare bare shaft cyclone feed pump is to be stored at site to allow immediate change-out of a defective pump. The scats from the two ball mill trommel screens discharge to two scats bunkers. A maximum 3016 t/h of oversized slurry particles reports to each of the two cyclone cluster underflows and is directed back to the two ball mill feed chutes. A maximum 1579 t/h from each of the two primary grinding cyclone overflows, having a particle size P80 of 106 µm, is transferred to cyclone overflow boil boxes, and from there via gravity to the copper rougher flotation feed conditioning tank. The design ball mill power at the pinion (expected operating) is predicted at 11.8 MW each. With the optimum grind size for flotation and a Bond ball mill index of 16.3 kWh/t, two 13 MW ball mills were selected. Ball mill grinding media (steel balls – 65 mm nominal size) are fed by a FEL to the ball mill storage bins. Balls are discharged to a kibble attached to the jib crane and transferred to ball mill ball feeders and storage bins. Balls are added to either of the two ball mills by diverting balls from the feeder to one or other of the two ball mill feed chutes. Lubricating oils are cooled by oil-to-water heat exchangers. The SAG mill gearless (ring or wraparound) drive motor’s starter is cooled by cooling fans and air/water heat exchangers. The SAG mill motor’s “E-house” is also water-cooled. Closed circuit cooling water for the entire grinding area is circulated through a cooling tower located at the stockpile and pebble crusher area. Water is removed from the grinding area by two sump pumps. A large mobile crane retained at the site would be available for maintenance in the mill and cyclone areas. 18.4.2.4

Copper flotation

The function of the copper flotation area is to recover copper and molybdenum into a copper/molybdenum concentrate, reject zinc to the cleaner scavenger tailings stream and reject coarse non-sulphide gangue to the rougher tailings. Cyclone overflow from the primary grinding hydrocyclones reports via gravity to the agitated rougher conditioning tank in the copper rougher circuit. The feed is conditioned with slaked lime to ensure the

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circuit pH is maintained at its set value. The conditioned feed reports to two copper rougher flotation banks each consisting of six 300 m3 mechanical flotation cells, operated at a pH of 10.0. These produce a bulk concentrate containing copper, silver, and molybdenum which reports to the copper regrind cyclone feed sump. The copper tailings are fed to the tailings thickener for eventual disposal at the TMF. Spillage and water are removed by two flotation area sump pumps. The copper rougher bulk concentrate is pumped to a cyclone cluster. Finer material, which does not require regrinding, reports to the cyclone overflow. The coarse material reports to the cyclone underflow and is pumped to the copper cleaner regrind mill. A 3 MW IsaMill is used. The mill gearbox and motor lubricants are cooled using the closed circuit cooling water system. The regrind mill discharge is combined with the copper regrind mill cyclone cluster overflow, and is pumped to the agitated copper cleaner conditioning tank. In the copper cleaner circuit the pH is increased to 11.0 with slaked lime. Zinc sulphate (ZnO4) is used as a depressant for sphalerite. The copper cleaner circuit consists of three stages of cleaning and one bank of cleaner scavenger flotation cells. The first copper cleaning circuit consists of five 100 m3 1st stage cleaner flotation cells. Concentrate containing copper and molybdenum is recovered and reports to the 2nd stage copper cleaner feed sump and the 2nd stage copper cleaner flotation cells for further upgrading. The tails report to the six copper cleaner scavenger flotation cells. The 2nd stage copper cleaner flotation stage consists of two 100 m3 mechanical flotation cells. The concentrate reports to the 3rd copper cleaner flotation stage for further upgrading. The tails from the 2nd copper cleaner report back to the copper cleaner conditioning tank. The 3rd copper cleaner flotation stage consists of two 172 m3 flotation columns, 4.27 m diameter x 12 m high. The column feed pumps and circulation pumps have a common 100% standby. The columns produce a copper/molybdenum concentrate which reports to the molybdenum feed thickener in the molybdenum flotation circuit for further treatment. The tails stream reports to the head of the 2nd stage copper cleaner flotation stage via the feed sump. The flotation columns are supplied with compressed air from dedicated copper flotation column compressors. The copper cleaner scavenger stage consists of six 100 m3 mechanical flotation cells. Low grade concentrate is recovered and reports back to the copper cleaner conditioning tank. The tails are fed to the tailings thickener. The flotation cells, other than the flotation columns, are supplied with low pressure air from dedicated low pressure air blowers. The copper/molybdenum concentrate from the copper 3rd stage cleaner flotation column reports to an 18.0 m diameter copper/molybdenum feed thickener. The main function of the thickener is to reduce carryover of reagents from copper flotation into the molybdenum flotation circuit. Thickener overflow is returned to the copper circuit.

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18.4.2.5


The function of the molybdenum flotation area is to recover molybdenum into a molybdenum concentrate and reject copper to the tailings as a copper concentrate. The stages of molybdenum flotation are: 

Rougher



1st to 7th stage cleaners



Cleaner scavengers.

The molybdenum roughers, first cleaners, second cleaners and cleaner scavengers consist of 10 m3 cells, and the subsequent molybdenum cleaner circuits consist of 3 m3 cells. The flotation cells are covered and air is induced into each cell. Underflow from the copper/molly concentrate thickener is pumped to a 38 m3 agitated molybdenum rougher conditioning tank. NaHS is added to the conditioning tank to inhibit the flotation of copper, with promoter also added to enhance the flotation of molybdenum. From there the overflow flows to the ten 10 m3 molybdenum rougher flotation cells. The tailings from the rougher cells form a portion of the copper concentrate and are pumped, with the molybdenum cleaner scavenger tailings, to the copper concentrate thickener. The molybdenum rougher concentrate reports to the molybdenum 1st cleaner flotation cells. The molybdenum cleaner circuit consists of seven stages of cleaning and one bank of cleaner scavenger flotation cells. The molybdenum 1st stage cleaner consists of three 10 m3 mechanical flotation cells. From these a concentrate containing molybdenum values is recovered and flows to the molybdenum 2nd stage cleaner flotation cells for further upgrading. The tailings from the molybdenum 1st stage cleaner flow to the two 10 m3 molybdenum cleaner scavenger flotation cells. The molybdenum 2nd stage cleaner flotation consists of two 10 m3 flotation cells. From the flotation cells the concentrate reports to the molybdenum 3rd stage cleaner flotation cells for further upgrading. The tailings from the molybdenum 2nd stage cleaner flotation cells reports to the molybdenum 1st stage cleaner flotation cells. The molybdenum 3rd stage cleaner flotation consists of two 3 m3 flotation cells. From the cells, the concentrate reports to the molybdenum 4th stage cleaner flotation cells for further upgrading. The molybdenum 3rd stage cleaner tailings report to the molybdenum 2nd stage cleaner flotation cells. The molybdenum 4th stage cleaner flotation consists of two 3 m3 flotation cells. The cells produce molybdenum concentrate which reports to the molybdenum 5th stage cleaner flotation cells. The tailings from the molybdenum 4th stage cleaner cells reports to the molybdenum 3rd stage cleaner flotation cells. The molybdenum 5th stage cleaner flotation consists of a single 3 m3 flotation cell. The cell produces a molybdenum concentrate which reports to the molybdenum 6th stage cleaner flotation cell. The tailings from the molybdenum 5th stage cleaner cell reports to the molybdenum 4th stage cleaner flotation cells.

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The molybdenum 6th stage cleaner flotation consists of a single 3 m3 flotation cell. The cell produces a molybdenum concentrate which reports to the molybdenum 7th stage cleaner flotation cell. The tailings from the molybdenum 6th stage cleaner cell reports to the molybdenum 5th stage cleaner flotation cell. The molybdenum 7th stage cleaner flotation consists of a single 3 m3 flotation cell. The cell produces a molybdenum concentrate which is pumped to the molybdenum concentrate thickener. The tailings from the molybdenum 7th stage cleaner cell reports to the molybdenum 6th stage cleaner flotation cell. The molybdenum cleaner scavenger consists of two 3 m3 mechanical flotation cells. A low grade concentrate is recovered and is pumped back to the copper/molybdenum feed conditioner tank. The tailings from the molybdenum cleaner scavenger are pumped to the copper concentrate thickener. Spillage and water are removed by a sump pump in the flotation building. 18.4.2.6

Copper concentrate thickening and filtration

The tailings from the molybdenum flotation roughers and molybdenum cleaner scavengers report to an 18 m diameter copper concentrate thickener, via a static screen to remove tramp material. Flocculant is added to enhance settling. The clarified thickener overflow reports to the process water pond, to be pumped to the process water tank. The thickener underflow is removed at 60% solids via duty/standby peristaltic pumps. The thickened and de-tramped slurry is stored in two 860 m3 agitated tanks. These provide a 24 hour surge capacity, allowing filter maintenance to be conducted without affecting mill throughput. The filter feed is pumped to three vertical pressure filters with a combined maximum filtration rate of 91.2 t/h. The filters operate with a total cycle time of 12.5 min and are designed to produce a filter cake of 9.0% moisture. The filter cake is dropped onto a belt feeder and is transferred to the product conveyor. Filtrate, cloth wash, and flushing water are returned to the copper concentrate thickener. The filter uses compressed air for both drying air and pressing air. Compressed air is generated by two 31.7 Nm3/min compressors, with a maximum working pressure of 20 bar. Spillage and water are removed by a filtration building sump pump. 18.4.2.7

Concentrate storage and loadout

After filtering, the copper concentrate is transferred to a tripper conveyor and discharged onto the copper concentrate product storage stockpiles inside the copper concentrate and load-out building. The building contains a live stockpile with capacity for seven days production (12 024 t), as well as an emergency stockpile with a further seven days production capacity, if problems arise with shipment to the port. 18.4.2.8

Molybdenum concentrate thickening and filtration

The molybdenum concentrate gravitates to a 3 m diameter thickener where it is thickened to 60% solids. The thickener overflow reports to the molybdenum cleaner process water tank, where it is used as process water in the molybdenum flotation circuit. The thickened concentrate is pumped to an

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agitated 48 m3 molybdenum surge tank, which has a residence time of 24 hours and acts as surge capacity for ferric chloride leach areas. Molybdenum concentrate is pumped from the surge tank into the agitated ferric chloride leach tank where copper and lead present in the molybdenum concentrate is dissolved to reduce the copper content to less than 0.5%. The slurry is then pumped across a static screen to remove any tramp material, and fed to an agitated molybdenum filter storage tank. The slurry is then pumped to a pressure filter to produce a filter cake of approximately 15% moisture. Filter cake is conveyed by a bobcat to the molybdenum bulk bagging plant. Molybdenum concentrate is bagged into 1 m3 bulk bags. The bulk bags are loaded onto trucks using forklifts. The filtrate from the filter reports to the tailings thickener. The thickener overflow flows by gravity to the process water tank. 18.4.2.9

Tailings disposal

The copper tailings flow by gravity from the two copper rougher tailings lines, and by pumps from the cleaner scavenger tailings line, to a 94 m diameter tailings thickener, where it is thickened to 52% solids. Thickened underflow is then pumped to the tailings dam. The thickener overflow gravitates to the process water tank where it is used in the grinding and copper roughing circuits. 18.4.2.10

Reagents

The reagents area is located to the west of the grinding and flotation areas. Hydrochloric acid and promoter are delivered to the reagents area by road tankers and transferred to storage tanks for distribution to the end users. Liquid reagents such as frother and collector are held in the reagent store and relocated to the flotation floor when required. Powdered lime, flocculant, FeCl3, NaHS and ZnSO4 are delivered as solids to the reagents area by road trucks and stored in the reagents building. A packaged flocculant mixing plant prepares flocculant for storage in an agitated tank for distribution. Other powdered reagents are mixed in agitated tanks before storage in tanks for distribution. The FeCl3 mixing tank is heated by a steam heating coil. Reagents are distributed by duty/standby pumps or multiple pumps. Spillage and water are removed by four reagents area sump pumps. A lime silo and a ball mill prepare lime upstream of the lime slaker. Lime is distributed by four pumps. Spillage and water are removed by a lime slaking area sump pump. 18.4.2.11

Water services

Water services cover raw water, fire water, process water, gland seal water, potable water, safety shower water, and cooling water.

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The raw water supply is provided from bores to the process plant via a 2506 m3 fire/raw water tank, located near the copper tailings thickener. The average requirement for raw water is 365 m3/h. The bottom 50% of the tank’s volume is reserved for fire protection, with the remainder being available for other use. Water is distributed around the site from this tank. Raw water is also piped to the flotation area for use as molybdenum make-up water and for gland seal water. The fire water system consists of independent fire water pump sets and a dedicated fire water main. The fire water main is generally buried and is fitted with above-ground fire hydrants and hose reels. A jockey pump is used to maintain pressure in the fire water main. The electric fire pump is the duty pump, with the diesel fire pump as standby. The fire-water reserve provides four hours of fire fighting capability at a fire water pumping rate of 344 m3/h. The raw water supply also provides water to the potable water treatment plant located between the process plant and the accommodation camp, and is capable of producing 10 m3/h of potable water. Potable water, stored in a 528 m3 potable water tank at the process plant, is pumped to the potable water reticulation system and the emergency shower water network by duty/standby potable water pumps. Raw water provides make-up water to the cooling tower for cooling water. Cooling water is in a closed circuit cooling water system, cooled by a heat exchanger coil in the open circuit cooling tower. The closed circuit cooling water is used for cooling bearing lubrication oil, hydraulic oil, gearbox oil, and the SAG mill’s “E-house”. Closed circuit cooling water is stored in a tank and circulated by pumps. Gland seal water for process plant pumps and other seals is provided by pumps. 18.4.2.12

Air services and other miscellaneous services

Air services comprise plant air, instrument air, flotation column air, high pressure air and low pressure blower air. Plant air compressors supply the compressed air for plant air and instrument air services. Instrument air is taken from the plant air circuit before filtering and drying for instrumentation service. Dedicated high pressure air systems are used for pressing and drying of the copper and molybdenum concentrates at the high pressure filters. The high pressure air compressors are located at the filter buildings. Air receivers are provided in each compressed air circuit to lessen any air pressure drop during periods of high air usage. Low pressure air, for the copper flotation cells is generated by blowers.

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Steam is generated by an electric, low pressure, packaged boiler. Steam is used for heating the ferric chloride leach tank. 18.4.2.13

Laboratory

The laboratory facility is located within the plant site working area, close to the plant and administration area, and comprises the following areas: 

Sample receival



Wet sample preparation



Sample drying and preparation



Wet chemical and environmental laboratory



Instrument room – AAS, Sulphur Analyser



Balance room



Office



Metallurgical laboratory – grinding mill, flotation cells, cyclosiser, Malvern laser sizer



Storage for samples



Store room for reagents and consumables.

The laboratory is designed to process samples for mine grade control, exploration, process plant metallurgical accounting, metallurgical optimisation and environmental control. In addition, the laboratory will have the facility to undertake metallurgical optimisation testwork such as grinding, flotation, leaching, etc. Analysis required for the environmental samples includes water quality analysis (Hach portable spectrophotometer), NAG tests to measure acidity and pH from waste rock and tailings, and soil analysis for exchangeable cations and anions, conductivity, nutrients, total solids, etc. 18.4.2.14

Diesel storage

The main diesel storage is provided by one tank of 750 m3, containing diesel supply for two weeks. The storage facility is located on the south side of the heavy vehicle workshop. Fuel is pumped from the main storage facility to a 50 m3 day tank located south of the workshop/warehouse facility. Diesel is pumped from the day tank to the fuel metering system. This facility services both heavy (i.e. mining) and light vehicles. 18.4.3 18.4.3.1

Electrical Voltage levels

Project voltages will be aligned with existing Peruvian voltages, which follow the North American system. Voltages used include 22.9 kV, 13.8 kV, 4.16 kV and 460 V.

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18.4.3.2

Altitude derating

The site has a nominal design altitude of 4300 masl (though the process plant is at around 4100 m). Standard equipment ratings for electrical equipment typically only apply up to 1000 m. Medium Voltage (MV) design involves two correction factors to compensate for the lower pressure and density of air at high altitude: 

A voltage correction factor – 0.67 at 4300 m



A current correction factor – nominally 0.933 at 4300 m.

In practical terms, equipment is often fully substituted with higher rated equipment. For example, a 22.9 kV application may use a 36 kV rated air-insulated switchboard, and similarly a 13.8 kV application may use a 22.9 kV rated air-insulated switchboard. Note that a gas-insulated switchboard (GIS) is usually fully-rated independent of altitude, and utilises special fully-insulated plug-socket cable terminations. Low Voltage (LV) equipment typically only has current derating, with the standard phase clearances proving adequate for the lower voltages. Vendor confirmation on equipment suitability for given altitude and duty have not been verified and vendor statements have been accepted for the purposes of this study. 18.4.3.3

MV power distribution

The following 13.8 kV switchboard is proposed: 

Grinding.

The following 4.16 kV switchboards are proposed: 

Primary crusher



Grinding



Pebble crushing



Tailings



Tailings dam.

The following 22.9 kV overhead lines are proposed: 

Primary crusher / heavy workshop / pit



Camp / tailings dam / bores



Camp waste water treatment plant.

The pit transportable substation produces 7.2 kV to suit the mining equipment.

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Neutral earthing resistors (NERs) have been included on all transformers supplying MV switchboards and LV motor control centres. 18.4.3.4

Transformers

All transformers are oil-filled, outdoors, ONAN type, except for the smaller, indoors types used for small power distribution, which will be constant-voltage, dry-type. Transformer protection will be included as per the design criteria. Where practicable, the protection will trip the upstream supply circuit-breaker upon fault detection. Allowance has not been made for redundancy on transformers within the process plant area. It is recommended to purchase an additional transformer for each size as a spare. Other than in the grinding area, no allowance has been made for redundancy on transformers within the process plant area. It is recommended to purchase an additional transformer for each size as a spare. All outdoors transformers include the capability for future conversion to ONAF cooling to accommodate future expansion capacity if required. 18.4.3.5

MV switchgear

All MV switchgear within the processing plant will be fully type-tested. A preliminary fault study showed that a 16 kA fault rating could be applied for certain MV switchboards, however, a 25 kA fault rating has been proposed for all MV switchboards within the processing plant to simplify procurement, spares and maintenance. All MV switchboards will have remote operator panels, so that operators are not required to be in front of the switchboard when an open or close operation is to be performed. 18.4.3.6

Pit MV distribution

The pit area has a 22.9 kV overhead power line run to the proximity of the pit outer boundary. From there the power line extends partially around the pit periphery in both directions. A number of stab lines head towards the centre of the pit. Additional stab lines and completion of the peripheral run fully around the pit have been allowed for in the design, but those future works are not costed. The pit machinery includes 7.2 kV drills and shovels. Two transportable substations have been included, each capable of powering one shovel and two drills. Substations are mounted onto heavyduty steel-framed skids. The 22.9/7.2 kV transformer is specially designed and manufactured to ensure longevity despite rough movement. LV power for pit dewatering pumps has also been incorporated into the substation.

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18.4.3.7

LV motor control centres

LV MCCs are all 460 VAC, double-sided, indoors type. Distribution transformers have been sized to minimise spares inventory, with the preliminary selection for the largest size of 2500 kVA being chosen above the smaller 2000 kVA unit in order to minimise the quantity of MCCs required for the project. For a distribution transformer size of 2500 kVA, an MCC minimum fault rating of above 70 kA would be required. It is typical within the industry for costs to increase disproportionately when fault ratings exceed the industry threshold of 63 kA. Further assessment and discussion with potential vendors would be recommended to confirm whether it is more beneficial to reduce the large distribution transformer capacity (to say 2250 kVA) or to select the higher MCC fault rating. MCC starter field control circuit voltage is 24 VDC. Contactor control voltage is centre-tapped 110 VAC for improved safety. All LV switchgear within the process plant would be fully type-tested. Soft starters are used on the larger non-VSD 460 V drives in order to reduce peak instantaneous current on the MCC, and hence contribute to reducing the applicable arc flash category. 18.4.3.8

Switchrooms

Switchrooms within the process plant will be blockwork buildings. A transportable building is allowed for at the tailings dam. The transportable switchroom will be metalclad and manufactured from non-combustible materials, mounted a nominal 2 m above the ground. The following features have been included for the switchrooms: 

Air-conditioning, suitable for generated heat load of switchboards, MCC and VVVFs at (n+1) redundancy



Complete fit-out



Stairs and landings (as applicable)



Lighting and small power distribution boards and associated transformers



Pressurisation system to exclude dust



UPS system including distribution board and bypass switch.

A fire detection and suppression system has been included within each switchroom. 18.4.3.9

Fire detection and alarm system

Substation fire indicator panels (FIPs) would report to a site main fire indicator panel (MFIP). connection to the MFIP has been included within the new control room.

A

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The substation FIPs are connected to the MFIP via a fibre-optic connection. A basic fire detection and alarm system has been included for the process plant areas. 18.4.3.10

Variable Speed Drives (VSDs)

The variable speed drives (VSDs) are based on: ABB ACS1000 series for MV motors ABB ACS800 series for LV motors. All VSDs are mounted external to MCCs. The ABB MV VSDs cannot operate at 4.16 kV output at this altitude, and can drive a maximum 3.3 kV motor. LV VSDs have significant (x 0.67) current de-rating applied due to the altitude. 18.4.3.11

Diesel generated emergency supply

A 6 MW diesel generated emergency supply is to be provided by a turnkey power station vendor. Preliminary assessment, based upon recommendations from a potential vendor, is that four 4.16 kV diesel generating sets would be connected to a 4.16 kV power station switchboard, which feeds through a 4.16/22.9 kV transformer onto the 22.9 kV main switchboard. Generator power output de-rating is 0.66 at 4100 masl, meaning larger units are required compared to sea level. 18.4.3.12

Other power

Main power and lighting distribution has been included around the EPCM and construction contractors’ yard. 18.4.3.13

Harmonic mitigation

Harmonics primarily stem from the SAG mill cyclo-converter and electronic variable speed drives (both MV and LV). To mitigate the effect of plant harmonics, two harmonic filters have been included to account for harmonic distortion on the main switchboard and the point of common coupling to the supply authority. Various mitigation methods have been included to improve harmonic levels within all other MV and LV switchboards, including: 

High pulse conversion in all MV variable speed drives



‘Low harmonic’ type VSDs utilised for the larger LV sizes



Harmonic filtration accessories for the moderate-size LV VSDs.

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Two outdoors combined harmonic filter/power factor correction units have been included– one on each side of the main switchboard 22.9 kV bus. These units filter the harmonic content of the plant to meet typical power supply authority connection limits. As the main switchboard bus-tie can be closed, the two units incorporate a master-slave capability to allow both units to be operating simultaneously on the same bus. The Constancia site has a requirement for an overall power factor of 0.98. This would normally necessitate dedicated MV power factor correction at Constancia. The proposed Tintaya expansion includes a 60 MVAr capacitor compensation bank on the substation bus. However no information is available to determine how much of that may have been allocated to the Constancia line. Thus all required MVArs for power factor correction have been incorporated into the correction unit at Constancia. The Constancia site has a requirement for a minimum power factor of 0.95 to satisfy COES requirements. This would normally necessitate dedicated MV power factor correction at Constancia. The proposed Tintaya expansion includes a 60 MVAr capacitor compensation bank on the substation bus. However, no information is available to determine how much of that may have been allocated to the Constancia line. Thus all required MVArs for power factor correction have been incorporated into the correction unit at Constancia. A harmonics study will be undertaken by a specialist consultant, which may be the harmonics equipment vendor. These units occupy a significant combined land area, and should be assessed early in the project design phase. 18.4.3.14

Lighting and small power

MCCs supply 3 phase power to dedicated 460/400-230 V, L and SP transformers which then feed adjacent L and SP distribution boards. Three phase small power is 380 V (as opposed to 460 V for motors), with single phase at 220 V. Bulkhead lighting type is predominantly 70 W high pressure sodium, and fittings are mounted either using 2.4 m lighting poles, or attached directly to buildings and other infrastructure where possible. Emergency light fittings will have individual batteries. An assessment of a centralised UPS system may be considered in the design phase. 18.4.3.15

Earthing and lightning protection

Earthing and lightning protection includes the following hardware implementations: 

Buried equipotential bonding conductor run throughout process plant area and attached to all major structures and equipment



Earth mats at all transformer yards



Earth tails from all structure columns and concrete reinforcing onto the buried conductor system

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

Overhead earth wire on all overhead power lines, taken to earth at regular pole intervals



Surge arrestors at all overhead-to-surface power line transitions



Surge diverters on all LV MCC incomers.

18.4.4 18.4.4.1

Instrumentation and control Instrumentation - general

316SS marshalling boxes, with sloping roofs, are included in areas where a high density of instrumentation exists. Stainless steel splash/dust covers for instruments and marshalling boxes have been allowed for. Flow transmitters on services with a pipeline size of greater than 100 mm NB are remote type. Control valve position feedback has been included as for specific valves. Density transmitters as detailed in the instrument list have been allowed for. On site storage of nuclear sources will require a separate small dedicated secure building with appropriate signage. All instrumentation is on UPS power; the extent of this may be reviewed in detail design. 18.4.4.2

Process Control System (PCS)

A PLC/SCADA system has been selected as the PCS. The PLC brand proposed is Allen-Bradley, with a Citect brand SCADA system. A PLC node is placed into a tier of each MCC. All discrete MCC and starter I/O are wired directly to this node. Field instruments are wired into a separate floor-mounted cabinet, which contains a PLC node and marshalling terminals for instruments that are not bus-connected. This panel could be incorporated into the MCC if instrumentation quantities do not necessitate a separate panel. This panel (or section of the MCC) will contain the CPU for the substation/process area. Fibre-optic communications to third-party vendor packages has been allowed for the SAG mill, ball mills and regrind mill. The SCADA interface includes a minimum of two operator stations, each with two large monitors. Separate engineering workstations have also been included. The PLC network and the SCADA network will be separate and independent. The fibre cable routing and network architecture is such that any one fibre cable failure point will not cause any data loss on the network. All PLCs, SCADA, PCs and associated network communications devices will be on UPS power.

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18.4.4.3

Emergency shower and eye wash station

A green fluorescent light and up to two proximity switches (to monitor flow) or foot switches have been allowed for, connected into the site control system to raise an alarm in the control room upon the operation of the emergency shower for each emergency shower/eye wash station. 18.4.4.4

CCTV systems

Thirteen CCTV cameras have been included in strategic operational locations requiring supervision/monitoring from the control room. A monitoring station is included in the control room. 18.4.4.5

Telephone system

A field telephone distribution system and central PABX in the Administration Building is included. Communications and data are provided to site via a fibre optic connection provided as part of the transmission line. 18.4.4.6

Hazardous areas

The entire molybdenum plant has been classified as a hazardous area (i.e. potentially explosive atmosphere). Hazardous areas have also been identified in the Reagent Area. Electrical equipment and instrumentation in these areas include additional requirements for equipment selection and installation within a hazardous area. 18.4.5 18.4.5.1

Buildings Heavy and light vehicle workshop and mine office

The vehicle workshop facility will be constructed to accommodate heavy and light vehicle fleet maintenance activities. Office accommodation will be provided for workshop personnel, the mining engineers and support staff. There will also be a spare parts store and a toilet block. The main workshop (approximately 2,484 m2) includes three truck bays, and one welding bay equipped with a vehicle hoist, battery storage, tyre bay, main workshop, bulk lubrication storage, parts store, equipment store, tool store, oil/fuel laboratory and electrical workshop. A full length 40 t overhead crane, a 10 t crane and an air compressor and receiver are provided. The light vehicle workshop is included in the total area. The first floor above the main workshop includes the office area, consisting of engineering/geology superintendent’s office, maintenance superintendent’s office, mining superintendent’s office, meeting room, nine additional offices and a crib room. Male and female change rooms, and mustering areas are included with the heavy and light vehicle workshop buildings. This facility will be a steel-framed building, totally enclosed with roller doors, with translucent roof and wall panels for lighting during daylight hours.

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18.4.5.2

Heavy vehicle washbay

This facility is sized to wash a single 220 t haul truck, and consists of a concrete pad with steel platforms to access the vehicles. 18.4.5.3

Vehicle fuel distribution bowser

This facility is designed to refuel light vehicles and two 240 t haul trucks at a time. The facility will consist of canopies for personnel protection at the two refuelling locations. 18.4.5.4

Mine dispatch centre

This facility will be constructed at a location which permits mining operations to be observed and includes a computerised dispatch system. The facility will be a totally enclosed steel-framed building. 18.4.5.5

Core shed

The core shed facility (nominally 420 m2) will be constructed to store exploration samples, and to accommodate sample preparation and analysis required by the mining operation. The facility will be a totally enclosed steel-framed building, with roller doors and translucent roof and wall panels. 18.4.5.6

Process plant industrial buildings

Pebble Crusher Building This facility accommodates the two pebble crushers and associated equipment. The building consists of a steel portal framed structure, without cladding, and a full length overhead crane for pebble crusher maintenance. Copper Filtration Building This facility accommodates all the equipment associated with the copper filtration plant. The building (28x10 m) consists of a steel portal framed structure, fully clad, with 2x6000 high double-access doors and a full-length overhead crane. Molybdenum Filtration Building The molybdenum filtration plant is to be in a building, approximately 15x18 m, comprising a steel portal framed structure, fully clad, with 6000 mm high double-access doors, and an overhead crane. Concentrate Storage Building This facility accommodates the storage of filtered copper concentrate, delivered by a product stockpiling conveyor. It is a steel framed structure, fully clad to concrete bunker walls, with 4 m high double roller shutter doors to access each stockpile. Copper Flotation Blower Building This facility is constructed to accommodate nine blowers and ancillary equipment to service the copper flotation plant. The building consists of a steel portal framed structure, fully clad, with two by 6000 mm high doubleaccess doors, and a full-length overhead crane. The approximate size of the building is 18 x 10 m.

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Air Compressor House This facility accommodates three air compressors and ancillary equipment to service the process plant. It also houses the main instrument air drier. The approximate size of the building is 16 x 6 m (100 m2) Reagents and Packaging Store This facility caters for the reagents and packaging requirements. The approximate size of the building is 430 m2. 18.4.5.7

Process plant architectural buildings

Main Security and Gatehouse The main security and gatehouse building constitutes the main access to the plant site, and will screen up to 800 staff during each shift change at peak production. The main access road ends at the gatehouse, and parking is provided for private vehicles outside the premises. This building has a slab area of approximately 240 m2, extended eaves and roof over turnstiles, and features vehicle boom gates, search room, pedestrian rotating gates, weighbridge for in and out transport, and security/access badge control system. The building is of a modular prefabricated style. Main Control Room – (In Grinding Building) This facility caters for two control personnel, and is of a modular prefabricated style (floor area of approximately 100 m2). Control Room (Primary Crushing) This facility caters for two control personnel, and is of a modular prefabricated style, size: approx 18 m2) Laboratory This facility is of a modular prefabricated style (floor area of approximately 340 m2). Laboratory Store This facility is used to store goods used in metallurgical/laboratory testing. This unit is of a modular prefabricated style and forms part of the laboratory building (floor area 340 m2). Workshop/Warehouse This is a combined facility of warehouse and workshop in one building. The workshop is designed to carry out repairs and maintenance required by the plant. It also includes office accommodation for workshop staff. The toilets and crib room facilities are shared with the warehouse staff on the first floor above the warehouse. This building consists of a steel portal framed structure, fully clad, with two by 6000 mm high doubleaccess doors, and full length overhead crane and a portal crane are provided. The main workshop building, together with the warehouse, is nominally 2022 m2 with 500 m2 of office space above.

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The warehouse, is designed to accommodate goods receiving, storage and issue in support of plant requirements. It also includes offices for warehouse staff, amenities and an external storage compound (nominally 11 500 m2), fenced with two sets of double gates. The warehouse includes a goods receivable area, goods issue counter, and storage racking for materials and spare parts requiring undercover storage. Offices, toilets, and crib rooms are included on the first floor above the warehouse. The building is steel portal framed structure, fully clad, with two by 6000 mm high double-access doors and full length overhead crane. The approximate size of the building is 800 m2. Administration Building The administration building will accommodate management, supervisory and administrative staff. It includes offices, change facilities, first aid and crib rooms. There are undercover areas for personnel mustering and parking for mine rescue vehicles and a fire engine. The administration building is located adjacent to the main gatehouse. The total floor area is 1464 m2, and includes 25 individual offices, open work areas (58 workstations), printing room, server room, two meeting rooms, kitchen facilities, toilet facilities and a reception area. The approximate size of the building is 23x53 m. The changing facilities are on the ground floor occupying 1464 m2. They consist of male and female change rooms, lockers, toilets, laundry and crib room. The first aid area is on the ground floor and has offices for nurses, treatment rooms and a patient waiting room, including toilets. There is a separate ground floor area for fire offices which contains firemen's room and fire extinguisher room. Undercover parking is provided for a fire engine, mine rescue vehicle and ambulance, adjacent to the first aid facility. Kitchen and Mess The Kitchen and mess hall is designed to accommodate the maximum number of personnel (432) during the day shift. The building is located between the workshop/store and administration buildings. The building has a total floor area of 1034 m2, and includes a general storeroom, freezer room, cool room, food preparation area, kitchen and a buffet area. The approximate size of the building is 23x44 m. Crib Room No 1 - located in Grinding Building This facility is constructed adjacent to the grinding building control room to provide crib facilities. The approximate size of the building is 135 m2. Emergency Power Station This facility houses the four emergency diesel generators, switchgear and associated equipment. The building is a purpose designed building, installed at grade.

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18.4.5.8

Electrical buildings

Switchyard and Main Substation No 1 This facility houses substation switchgear and associated equipment for the incoming HV transmission line. The building is a two-level blockwork building, installed at grade. Substation No 2 Primary Crushing This facility houses switchgear and associated equipment for the primary crushing and heavy vehicle areas. The building is a blockwork building, installed at grade (floor area of approximately 70 m2). Substation No 3 Grinding This facility houses switchgear and associated equipment for the grinding area. The substation is a blockwork building, and installed at grade within and partially outside the grinding building (floor area of approximately 810 m2). Substation No 4 Pebble Crusher This facility houses switchgear and associated equipment for the pebble crusher area. The building is a blockwork building, installed at grade (floor area of approximately 70 m2). Substation No 5 Molybdenum Plant This facility houses switchgear and associated equipment for the molybdenum plant area. The building is a blockwork building, installed at grade (floor area of approximately 177 m2). Substation No 6 Tailings This facility houses switchgear and associated equipment for the TMF area. blockwork building, installed at grade (floor area of approximately 143 m2). 18.5

INFRASTRUCTURE

18.5.1

Access road

18.5.1.1

This building is a

Background

Road access to the mine is required for the construction and operational phases of the project. From a construction phase perspective, road access will be utilised for the transportation of the plant, equipment and personnel required during the construction period, with the majority of imported equipment transported from Matarani. From an operational perspective, road access is required to service the mine with the required consumables, reagents, fuels, mining plant and equipment, and mining and support staff. In addition, the road is required for the transportation of concentrate product from the mine to the port of Matarani. SIGT S.A. Ingenieros Consultores (SIGT), a Peruvian engineering consultant has undertaken a study of the road improvements necessary to provide the appropriate level of service for the Project. The study investigated the road improvement requirements from the city of Yauri to the site (Figure 18.24). The road has been divided into two discrete sections, the first being the 70 km section from Yauri to the Livitaca turnoff, and the second a 12.5 km section from the Livitaca turnoff to the process plant site.

Page 314

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.24 Constancia Project Location and Access Road Constancia

14Project .0

ABANCAY

Dv.Livitaca

70

CUSCO

.0 Yauri

0 80.

AREQUIPA 69.0 Imata

0 1. 10

Yura

25 .0

AREQUIPA .0 45 0 . Repartición (48) 5 1 57 .0

Dv.Mollendo

OCEANO PACIFICO Matarani

18.5.1.2

Existing road access

Location 

Section 1 begins at the city of Yauri, and passes through the villages of Coporaque, Pumahuasi and Finaya. This section of the road is unsealed. Two bridges at Quero and Upamayo will require improvements.



Section 2 begins at the Livitaca turnoff and follows the existing unsealed road up to the Chilloroya River. A new intersection will be constructed on this public road, from where a private road will be

Page 315


constructed to service the mine. This private road crosses the Chilloroya River over a new bridge and terminates at the entrance to the process plant. Topography The topography is typical of high plateaus along the summit of the Andes mountain range, with small uphill and downhill sections ranging in elevation between 3900 and 4684 masl. Existing Condition The section of road between Yauri and the project forms part of the National class 3 road system (less than 400 vehicles per day). The road condition generally consists of a pavement varying between 4 and 6 m wide. Road surfacing consists of a layer of compacted wearing course approximately 0.25 m thick, obtained from local borrow pits, and comprises either glacial tills or weathered rock. 18.5.1.3

Basis of design

Design Criteria The design criteria and specifications used are in accordance with the Ministry of Transportation and Communications or MTC” in Peru, and include the following design parameters: Governing speed

40 and 50 km/h

Minimum radius

50 m

Maximum gradient

9%

Roadway width

6m

Shoulder width

0.50 m (on each side)

Total width at grade level

7m

Maximum camber

8%

Triangle ditches

0.50 x 1.00 m

A typical cross-section of the proposed upgraded road is shown in Figure 18.25. Figure 18.25 Typical Section of Upgraded Access road TYPICAL SECTIONS

BERM

S = 2.5 %

S = 2.5 %

BERM

Road Considerations Design considerations included using the existing road alignment as much as possible, although some changes and improvements of the existing road alignments and turn radii are required.

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A traffic study was undertaken to confirm current vehicle movements, to which was added projected movements to support Constantia Project operations. Based on the proposed mine production, there will be an estimated daily increase in truck traffic in the order of 50-60 trucks exporting concentrates and importing supplies. Geotechnical Investigation A field work program and laboratory testwork was completed to support development of the design parameters. Test pits were dug to depths of 1.5 m at one km intervals along the route, and samples submitted for analysis. Pavement types and depths were designed based on the characteristics of existing materials, traffic forecasts, the road bearing capacity and environmental conditions, so that the road surfacing has an acceptable service “index” during its estimated life. In addition the investigation makes the following key findings: 

Low bearing capacity and expansive soils located along the route will require removal.



Groundwater problems exist predominantly in Section 1. Longitudinal sub-drains are designed in order to divert water sources below sub-grade, deplete the water table and eliminate seepage water to prevent moisture from affecting the performance of the road pavement.



The existing drainage systems require augmentation due to the high rainfall in the area.

Local sources of construction materials and suitable water supplies were identified as part of the survey. Hydrological Study An investigation was undertaken to determine the hydrological characteristics of the rivers and creeks, in order to determine the design requirements for the bridge and drainage structures. Bridge design and management of potentially flooded areas were based on this study. Return periods of 500 years were used for major bridges, reducing to 200 years for potentially flooded areas and down to 10 years for ditches (Table 18.21). Table 18.21 Design Return Periods for Hydraulic Structures Return Period - Year

Hydraulic Structure

10

Ditches

50

Culverts

100

Water crossings

100

Pontoon Bridges

200 and 500

Bridges

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18.5.1.4

Construction works

The road improvement works once complete will provide an all-weather road allowing the safe transportation of product, day and night. Road surfacing will use local materials such as glacial tills and weathered bedrock, although bridges will have asphalt seals. The route generally follows the existing horizontal alignment and, as such, no resumptions of land are required. An exception is in the vicinity of bridge approaches where there are minor departures from the current alignment that require land resumption and/or land owner compensation. Earthworks cut and fill operations are required to improve both the horizontal and the vertical alignments to meet the design requirements. Pavement works for both sections consist of the complete replacement/overlay of the existing pavement with a suitable pavement material compliant with the design requirements. Drainage is improved through cross-drainage culvert replacement and longitudinal drainage. Signage improvements are required over the full route to meet the design requirements. This includes guard rail, signage (regulatory, preventative and informative), delineating poles, etc. The design of all bridge work has been undertaken in accordance with the updated Peruvian design standards, AASHTO LRFD HL93, which considers live loads of up to 48 t. The capital costs for the improvements and new bridge construction have been considered in the final cost estimate. 18.5.1.5

Access road operation and maintenance

A program of maintenance will be required to ensure that the road remains in its optimal condition. The operating costs for the access road (other than the section of road on the Constancia site) will not be assumed by Norsemont as the road will form part of the National road system, and the governing MTC will be responsible for all future operating costs. 18.5.1.6

Access road environmental assessment

An environmental assessment for the road works has been prepared with reference to typical guidelines both from the Peruvian Ministry of Transport and Communications and the World Bank for road specific projects. This comprised: 

Field Reconnaissance - to collect information on baseline environmental conditions, road features, status of the road condition, and location of communities along the route. In addition, a review of ancillary facilities such as material sources, water sources, probable sites for excess material disposal, etc. were carried out.

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

Information was also collected from local authorities and citizens, including municipal departments, to obtain a better understanding of the natural and social environment that could be affected by project activities.



Population centres in the vicinity of the road were identified in order to analyse the dynamics and social and productive relations, problems and expectations that are common to Andean communities, either gathered or isolated. The assessment also included an analysis of the historical and archaeological components.

No major environmental constraints were identified, but attention is required to avoid impacting archeological remnants or other historical buildings, and to minimise the impact of noise, dust and increased likelihood of accidents. 18.5.1.7

Capital cost estimate

The capital cost estimate (Table 18.22) includes the total cost of construction of the road improvements. The Direct costs include all costs incurred by a construction contractor in executing the works, inclusive of non–management resources, plant and equipment, materials, consumables etc. The Indirect costs include the construction contractor’s overheads, inclusive of management and temporary works, the engineering consultancy costs for project management, engineering and supervision. The capital cost estimate is based on feasibility level design of the road improvements, bridge works and other drainage system requirements. Material volumes have been developed based on the designs. Unit rates have been derived from first principles and then applied to the quantities to develop item costs. The methodology is consistent with producing an estimate with an accuracy of ±15% as required for the DFS. The base date of the estimate is October 2008. Costs are inclusive of all payments associated with permits and approvals for all the works. Diesel costs of $0.67/L have been used, and a contractor profit of 10% of Direct costs added. The capital cost estimate for this component of the road and bridge work is estimated to be $20.2 M, excluding IGV.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.22 Access Road Summary of Capital Costs DESCRIPTION Preliminary works

TOTAL $ 265 701

Section 1: Yauri – Livitaca turnoff

11 738 391

Road

10 873 904

Quero Bridge

478 025

Upumayo Bridge

386 463

Section 2: Livitaca turnoff – Constancia (access to plant and camp)

2 117 116

Road

1 633 090

New Bridge at km 10+900 Total Direct Cost (preliminary works Sections 1 and 2) Fixed General Expenditures

484 027 14 121 209 190 576

Variable General Expenditures

1 435 864

Profit (10%)

1 412 121

Overhead Costs

3 038 561

Total (W/O IGV tax) Estimated cost of engineering studies (5% of works Direct Cost)

17 159 770 773 218

Estimated cost of work supervision (8% of work Direct Cost)

1 237 150

Temporary diversions and drainage structure propping

1 000 000

Total Capital – Access Road

20 170 138

The following are excluded from this capital cost estimate: 

Any land acquisition or compensation costs associated with “right of way ”agreements



Cost of archaeological assessments and rescues



Owner’s Costs.

18.5.2

Water supply

Total raw water requirements have been estimated at 365 m3/h for process plant, mine and camp operations. Field investigations and numerical modelling indicate that there is sufficient groundwater supply from pit dewatering operations. Raw water to the process will be supplied from the pit dewatering wells, in pit sumps, drainage from the PAG WRF and reclaim water from the TMF as described in Section 1.9.1. The water will be temporarily stored in a raw water tank. The raw water is of sufficiently high quality to be used directly in the process plant for hose-down, gland water, cooling water and reagent make-up. Up to 11 m3/h of raw water will be diverted to the potable water treatment plant where it will be upgraded by reverse osmosis and ultraviolet treatment, making it fit for human consumption.

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Process water will be supplied from water captured by in-pit sumps, the PAG WRF and rain falling on the plant site. Additional requirements will be met by returning tailings dam supernatant water to the plant process water pond. 18.5.3

Accommodation camp

Construction and operation workforces are housed in a purpose-built accommodation camp located adjacent to the process plant. The camp is supplied and installed by a single construction contractor, with another contractor operating the camp during the construction period. The camp will be constructed from modular units to minimise cost and expedite delivery. Management of the camp will be transferred to Norsemont once the construction phase is complete. The camp is self-sufficient and is to have full sleeping, bathing, dining, laundering and recreational facilities. The nearest town is located two hours away (by vehicle). The construction manning demand has been developed based on the current project schedule and labour estimates, and indicates that a maximum accommodation demand of approximately 1800 persons will occur. The operations workforce is expected to peak at approximately 800 persons. Operations recruitment strategy is to maximise the use of locals from the immediate area. As such, it is assumed that a minimum of 50% of the operations workforce will be sourced locally and commute to and from work each day. Accordingly, the maximum demand for the camp during the operations phase is 400 persons. The additional accommodation units not required after the construction phase will be demobilised, and the oversized facilities will be reduced in size to meet the operation’s requirements. Consistent with in-country industry standards the camp comprises three different types of accommodation units: 

Single accommodation with dedicated bathroom – 20% of total



Double accommodation with a bathroom shared between two double rooms – 15% of total



4 person accommodation with communal bathrooms – 65% of total.

All buildings are to be supplied prefabricated, and fitted out internally (including services) for their final intended use with air conditioning units fitted. The camp design includes camp administration offices, internal access roads, potable water treatment, storage and distribution, medical requirements, a recreation hall, waste management, sewage treatment, laundry, landscaping, IT and communications infrastructure, and site security. The capital cost estimate of the 1800 bed accommodation camp is $30.27 M, based on budgetary quotes from specialist Peruvian contractors, and has an accuracy of ±15%. This equates to approximately $16 819 per bed. The operating cost estimate is approximately $15 per person per day, again based on budgetary quotes from Peruvian contractors.

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18.5.4 18.5.4.1

Power supply Electricity market study

Cesel Ingenieros S.A. (Cesel) was engaged by Norsemont to undertake a DFS-level appraisal for the power supply. The electricity market study investigated the forecast demand for power in the region over a time period of ten years (2008 to 2018), with the year 2007 adopted as the base. The study covered the South Sierra region of Peru including the areas of Cusco, Arequipa, Apurimac, Moquegua, Tacna and Puno. In developing this study, CESEL has utilised information provided by energy distributors that have concessions in the region, together with information in the public domain as published by Osinergmin. Forecast changes in power demand and availability of supply in the region are heavily dependent upon Constancia and other proposed large mining projects in the area, which will ultimately determine the regional requirements for new generation and transmission infrastructure. 18.5.4.2

Power supply option

Option studies were undertaken based upon connection into the Peruvian National Interconnected Electric System (SEIN), and terms of reference for the investigation included technical, economic, environmental, land ownership, cultural and heritage considerations. The plant maximum demand is estimated to be 90 MW with an average load of 75 MW, remaining fairly constant over the operating life. The scope of the study included the required facilities at the origin of supply, the transmission line, switchyard and main switchroom at the Constancia mine site. The study identified the preferred option of initially securing supply at 138 kV from Tintaya substation, with transmission by means of a single circuit supported by lattice steel towers over a route length of 70 km to the Constancia mine site, designed for future operation at 220 kV. It was assumed that 220 kV supply will become available from the Tintaya substation in 2012. 18.5.4.3

Power line design

The transmission line from Tintaya substation to the Constancia mine site has been designed in accordance with relevant North American and European codes and standards, and in recognition of all environmental, geological, social and cultural considerations related to the land and airspace easement along the proposed route of the line. Topographic and geological surveys were completed along the proposed route and at the Tintaya and Constancia substation sites. An archeological survey was completed along the proposed powerline route, from which it was concluded that there would be no impact upon or compromise of any kind to archaeological remains in the vicinity of the easement along the full length of the proposed transmission line. However, a

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“Certification of the Non-Existence of Archaeological Remains” (CIRA) will be required from National Cultural Institute of Peru, prior to construction. Earth resistivity measurements were conducted along the proposed transmission line, to support the design of the line-grounding systems. The transmission line is designed for ultimate operation at 220 kV, will have a design capacity of 150 MW and traverse a route length of approximately 70 km. The proposed line is a single circuit of 600 mm2 AAAC conductor in triangular disposition suspended from four-legged galvanised steel lattice towers. The transmission line will also include an OPGW or optical fibre composite overhead ground wire serving the multiple functions of earth bonding for the towers, lightning protection shielding for the power conductors and communications via the enclosed 24 optical fibres. The fibre optic communications will serve protection relaying and control functions for the transmission line and data links from Constancia to the Peruvian communications network. 18.5.4.4

Constancia substation 138/220 kV

The proposed Constancia substation switchyard has been designed with overhead busbar systems, switchgear, metering and protection equipment to control the ultimate installation of two by 100 MVA primary transformers to provide the plant with 100% redundancy at this level in the main power supply. The main switchyard will be designed for a capacity in excess of 200 MVA, and the layout provides for the incoming bay for the transmission line from Tintaya and space for a future bay to accommodate extension of the line, or an alternative connection to the Peruvian grid. Initially, one 100 MVA primary transformer has been allowed for in the estimate with all associated switchgear, metering and protection equipment to supply the estimated plant maximum demand of 90 MW or 95 MVA at the minimum required COES power factor of 0.95. Switchyard switchgear current and short-circuit ratings will be standardised at 1250 A and 25 kA respectively. Insulation levels and surge protection equipment will be standardised with IEC values and reflect the 4000 masl elevation of the Constancia Substation and the level of iso-keraunic activity in the region. The proposed switchyard design and layout is based upon the use of air-insulated switchgear manufactured in accordance with IEC standards. The switchyard design will accommodate the installation of a 40 MVAR power factor correction and harmonic filtering unit for control of the plant power factor and harmonics on the 22.9 kV main substation bus. The primary transformer will be constructed with winding tappings of 220 kV/36.5-22.9 kV to facilitate the transfer from 138 kV to 220 kV in 2012.

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The main 22.9 kV metal clad switchboard is configured with two outdoor main circuit breakers connected by means of dual bus duct to the split main busbar system fitted with a normally open bustie. The switchboard provides for 29 supply feeders to the plant and the overhead distribution systems. The main switchboard is housed within a double-storey blockwork switchroom. The ground floor comprises a cabling chamber, battery, station transformer and auxiliaries room, and the upper floor houses the main switchboard, alarm, control and communications panels. 18.5.4.5

Tintaya substation 138 kV expansion

For the initial supply to Constancia at 138 kV, an additional switchyard bay at Tintaya substation has been allowed to accommodate the extension of the existing 138 kV double busbar, 45 MVAR of capacitive compensation equipment and all associated switchgear, transformer, metering and protection equipment for the transmission line. The additional bay proposed for Tintaya substation to serve the transmission line to Constancia will be designed for a capacity in excess of 200 MVA. Capacitive compensation for the transmission line will operate at 22.9 kV, connecting to the 138 kV bus via a 60 MVA, 138/22.9 kV transformer. Switchgear current and short-circuit ratings will be standardised at 3150 A and 40 kA respectively, in line with existing substation equipment and in consideration of forecast projections to 2016. Insulation levels and surge protection equipment will be standardised with IEC values and reflect the elevation of Tintaya substation and the level of iso-keraunic activity in the region. Transmission line reactive power compensation will be provided by a 45 MVAR unit complete with associated transformer, switchgear, control and protection equipment. 18.5.4.6

Tintaya substation 220 kV augmentation

The cost of the proposed 220 kV augmentation of Tintaya substation in 2012 has not been included in the estimate for the DFS. 18.5.4.7

Power supply transfer from 138 kV to 220 kV

The transmission line and Constancia substation have been designed for ultimate operation at 220 kV in 2012. The main power transformer and associated metering and protection transformers have been specified with winding tappings to enable operation at either 138 kV or 220 kV. Following initial operation at 138 kV, the transfer to 220 kV in 2012 will require a shut-down in order to changeover the power, metering and protection transformer tapping connections.

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18.5.4.8

Control and communications

Tintaya substation is operated by Red de Energía del Perú (REP) from its regional control centre in Arequipa, which, in turn, is supervised by the main control centre in Lima. The new transmission line from Tintaya to Constancia will be integrated into the REP SCADA system for purposes of monitoring, control, energy metering and load management. Tintaya substation has existing microwave links to Callalli, Ayaviri and Combapata substations, and there is also a telephone network linking all these substations with the REP control centres in Arequipa, Lima and the Peruvian telephone network. The microwave links serve the requirements for network tele-protection, data and telephony. The OPGW conductor forming part of the Tintaya-Constancia transmission line will also include 24 fibre-optic cores for tele-protection, data and telephony services. A local area network will form the platform for monitoring and control of the transmission line and the Constancia substation. An operator workstation in the Constancia substation will provide access to monitoring and control functions, system protection, energy measurement and alarms. 18.5.4.9

Power supply capital cost estimate

The capital cost for power supply is estimated at $24.45 M as shown in Table 18.23. Table 18.23 Power Supply Capital Cost Estimate Description

Capital $

Transmission line

8 521 734

Tintaya substation expansion

2 454 639

New Constancia substation

10 636 133

Telecommunications

18.5.4.10

369 042

Indirect estimate (EPCM)

2 470 907

Total

24 452 455

Operation and maintenance

The power transmission system, from Tinataya substation to the point of supply at the 22.9 kV terminals of the main transformer at Constancia substation, will be constructed, owned, operated and maintained by Red de Energía del Perú (REP). 18.5.4.11

Energy consumption

On the basis of an overall average demand of 90 MW at 138 kV and a peak/off-peak ratio of 8/16, the total annual energy consumption is estimated to be 615 GWh. From tariff information provided by OSINERGMIN, and an exchange rate of $1=3.2 Soles, this equates to an average cost of $48.90/MWh or ¢4.89/KWh.

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The tariff data used for this calculation is based upon OSINERGMIN Resolutions Nos.577-2008-OS-CD and 017-2009-OS-CD from their archive file PBA04022009.pdf dated 4 February 2009. 18.6

WATER MANAGEMENT

18.6.1

Introduction

Water management within the project area is divided into process and non-process water. Process water is water that will be used in the plant or will be conveyed around the site via pipelines. On an annual basis, the process water circuit will be in a surplus condition and to avoid continuous and ongoing accruals of water the excess amounts will be released at a controlled rate each year during the wet season. Releases will be from the TMF into the Chilloroya River and will only occur upon the achievement of adequate quality. The wet season discharges will take place when the Chilloroya River has a large assimilative flow. The water will be treated if necessary before release. Non-process water is surface drainage and runoff water that will require management at the site. It will be comprised of two streams: contact and non-contact water. Contact water is water that will come into contact with facilities such as the pit, stockpiles, the plant site, PAG WRF, etc. This water will be collected in designated ponds for treatment, as necessary, prior to its release. This water will be used in the process to the maximum extent possible. Non-contact water will be directed around the facilities via diversion channels and discharged into natural drainages upon the achievement of sediment removal to acceptable levels. Figure 18.26 presents a schematic of the process and non-process water circuits, and identifies the contact and non-contact water components. Hydrology and hydraulic analyses, including a detailed process water balance, have been completed to determine the types, locations and sizes of the water management structures as well as the times and quantities of discharges.

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Figure 18.26 Water Management Schematic

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The following sub-sections discuss water management for the project which includes process and nonprocess water components as well as details of the project process water balance. 18.6.2

Process water

Water for the mill will be supplied from the following sources: groundwater from dewatering wells located behind the pit walls, in-pit surface water collected in the pit sumps, drainage from the PAG WRF that will be collected in the containment pond just below the PAG WRF, and reclaimed water from the surface water pond in the TMF. A temporary water reservoir will be constructed in 2011 to collect and impound water to support the start of operations in early 2013. The temporary reservoir, located within the eastern valley of the TMF, will be constructed as part of the TMF Stage 1 construction. The reservoir will have the capacity to store up to five million cubic meters of water and will be filled with water collected from precipitation and spring discharges from the upper reaches of the TMF basin. If necessary, water will be pumped from the Chilloroya River into the reservoir during the wet season of 2012. Raw water for sustaining mill operations will be obtained, in part, from pit dewatering wells installed around the pit perimeter. Based on the current hydrogeologic model, a total of 18 dewatering wells will be installed in and around the pit. Water from these wells will be pumped to a collection box immediately upstream of the plant and then conveyed by gravity to a dedicated tank in the plant. Water from the PAG WRF and in pit sumps is expected to be acidic and will be neutralised with lime at the process plant before use in the process. When the volume of water available to the mill from the pit, TMF and PAG WRF exceeds the volume required the excess amount will be removed from the pit groundwater component, and a gravity-flow pipeline will convey it from the pit dewatering well system to the environment upon confirmation of acceptable water quality. The priority will be to recycle as much water as possible from the PAG WRF, in-pit sumps and TMF. The water balance indicates that during infrequent and short extreme precipitation events water in excess of the amount recycled to the process will report from the PAG WRF to the containment pond. Under such circumstances the excess water will be pumped to the upper reaches of the PAG WRF and sprayed onto the top for re-circulation through the pile. The resulting attenuation of flow will provide temporary storage for the excess water. On average, one such event can be expected to occur each year over the life of the mine. Water from the TMF surface water pond will be removed by dedicated pumps on a floating barge and will be pumped to a head tank above and immediately southeast of the ultimate limits of the TMF. From there the water will be pumped to the process water pond at the mill through a combination of HDPE and steel pipes. During dry periods, if the amount of water from the in-pit sumps and the PAG WRF containment pond is insufficient for the mill, additional water will be drawn out of the TMF. During periods of excess water discharge from the TMF surface water pond, the water will be pumped using a second set of dedicated barge mounted pumps to the TMF buffer pond located immediately

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north of, and below, the eastern side of the TMF embankment, and then from this pond by gravity flow to the Chilloroya River. Such discharges, when required, will only take place during the wet season, which is between December and April. Results of the current water quality analyses and water balance modelling indicate that during those periods, when run-off flows into the TMF are higher, no water treatment is expected to be necessary. The excess water removal pumps will be provided with instrumentation to monitor water levels and water quality. Water quality monitoring instrumentation will also be located in the buffer pond. The TMF buffer pond will provide six hours of storage capacity to allow time for mitigating actions in the unlikely event that water of inadequate quality is pumped from the TMF. In such a case, the water will be pumped back from the pond to the TMF. The following summarises the main water systems for the project: 

Pit surface water to plant treatment area and then to the process water pond



PAG WRF containment pond to plant treatment area and then the process water pond



PAG WRF water recirculation for temporary storage during short-term extreme events



TMF surface water pond to plant process pond



TMF surface water pond to TMF buffer pond before discharge to the Chilloroya River or pump back to TMF if dictated by water quality



TMF seepage collection sumps to TMF



Pit dewatering wells to plant raw water tank, with excess to TMF.

Sizing of elements associated with process water management was completed incorporating the results of the process water balance as described in the following section. 18.6.3

Water balance

A site wide process water balance was developed by Knight Piésold using its in-house computer model. This model evaluates the amount of stored water in the various components of the mine, as well as water inflows and outflows between the mine facilities on a monthly basis throughout the life of mine. A schematic of the process water balance is presented in Figure 18.27.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.27 General Water Balance Schematic

The main objectives of the process water balance are to provide operating guidelines, design flows, and information inputs for the design of various water management facilities. The results also provided input to the environmental management plans for the project. Specific key objectives and results of the site wide water balance are: 

Evaluate the range of normal operating volumes for the TMF surface water pond. This information is used to determine the pond configurations throughout the life of the facility. The predicted normal operating volumes for the TMF surface water pond vary between 1.8 and 3.3 Mm3.



Evaluate the maximum potential volume for the TMF pond above its normal operating volume to reflect conditions of extreme precipitation; this information is used to establish the minimum required freeboard heights for the TMF embankment at any time. The maximum operational volume for the TMF surface water pond may reach approximately 3.7 Mm3 after a 100 year/24 hr precipitation event.



Evaluate the amount of water to be discharged from the TMF surface water pond to the Chilloroya River via the TMF Buffer Pond. This information provides the design criteria for the discharge facilities as well as for the analysis of any downstream environmental impacts and mitigation plans. Under extreme precipitation conditions the amounts of water discharged via the TMF Buffer Pond

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may reach up to 3000 m3/hr. These discharges will occur only during the wet season and after adequate water quality has been confirmed. 

Evaluate flows reporting to the PAG WRF containment pond to determine the required capacity of the containment pond and the pumping recirculation rates under conditions of extreme precipitation. The capacity of the containment pond is 600 000 m3 below a freeboard of 4 m, or 800 000 m3 at the pond crest. Pumping requirements for recirculation under extreme conditions may reach 1100 m3/hr.



Evaluate approximate flows from the in pit sumps and PAG WRF containment pond to determine the needs for treatment at the process plant, as well as the pumping requirements. Normal operating flows from the in pit sumps are expected to be 0 to 400 m3/hr and from the PAG WRF containment pond are expected to be 100 to 800 m3/hr, due to seasonal fluctuations. Maximum operating flows are expected to be between 1200 m3/hr and 1800 m3/hr respectively.



Evaluate the amounts of fresh water from the pit dewatering wells to the operating system. Operating flows expected from the dewatering wells are expected to be at least 330 m3/hr.

18.6.4

Non-process water

Contact water Non-process contact water includes water that has been in contact with areas that generate sediment loads and/or requires treatment before being released. Diversion channels will be constructed to convey contact water to designated ponds or treatment areas. Contact water with sediment loads will be directed to sediment ponds before it is released to natural drainages; these ponds include the following: 

Pit sediment pond: drainage of the pit area during initial development is expected to occur mostly towards the south. Water in contact with areas of excavation in the pit will be directed to the pit sediment pond located immediately south of the pit and adjacent to the haul road. Due to the ground slope characteristics, water from the southwest portion of the pit area will be directed via contact channels to the main sediment pond.



Main sediment pond: this will be located immediately downstream of unsuitable material/topsoil stockpile No. 1. This sediment pond will collect water from several areas within the mine including haul roads and, to some extent, from areas upstream of the pit and from the final pit development. Contact channels will be constructed along designated routes to direct the flows to the main sediment pond.



Plant construction sediment pond: since the plant will be located in a different watershed from the main sediment pond, a temporary sediment pond will be located immediately downstream of the plant area to receive sediment laden water generated during construction of the plant.



Crusher sediment pond: as with the plant, the crusher will be located in a separate watershed, and a temporary sediment pond will be located downstream of it, connected via contact channels, to receive sediment laden water generated during construction of the crusher.

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Contact water that will require additional treatment will be managed as follows: 

Plant contact water: during operations, water in contact with the plant area will be directed to a dedicated plant contact pond. From this pond, water will be directed for treatment before it is released to a natural drainage or, alternatively, it will be directed to the mill process pond. Contact and non-contact channels will be constructed around the plant to reduce the amount of water directed to the plant contact pond.



PAG WRF contact water: water in contact with the PAG WRF will be directed to the PAG WRF containment pond and will not be released to the environment. To reduce erosion and sediment loads, temporary diversion channels will be constructed as the PAG WRF is expanded.

Non-contact water Non-contact water will be directed around the mining facilities via non-contact channels from and to natural drainages. Two non-contact water channels are currently contemplated as follows: 

Non-contact channel upstream of the PAG WRF: this diversion channel will direct water from a relatively large watershed upstream of the PAG WRF (Quebrada Cunahuiri) to the southern valley (Quebrada Telaracaca) below the PAG WRF.



Non-contact channel upstream of the pit: water from the drainage to the north of the pit will be diverted to the west valley. This diversion channel will be constructed in stages as the pit is expanded.

All channels will be lined for erosion protection with HDPE geomembrane (temporary channels) or riprap (permanent channels). The channels have been sized based on average surface run-off volumes and flows, as well as the 100 year storm event. Sufficient freeboard will be included in the permanent channels to contain flows associated with a 500 year storm event. 18.7

WASTE MANAGEMENT

Major waste management facilities within the project area include a PAG WRF and the TMF. Unsuitable material and topsoil generated during construction will be disposed in dedicated structures or in the major waste management facilities. The major waste management facilities are described in the following sections. 18.7.1

Waste rock facility

A PAG WRF will be developed to accommodate PAG waste rock mined from the San José and Constancia Pits. Approximately 295 Mt of waste have been characterised as having the potential to generate acid and will be placed in the PAG WRF. Approximately 55 Mt of waste have been characterised as non-acid generating and will be used as material for the construction of the TMF embankment, haul roads, construction roads and access roads. The PAG WRF will be located immediately southeast of the Constancia pit. Figure 18.28 presents the general layout of the facility and its associated structures.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.28 PAG Waste Rock Facility Stage 5 (Ultimate) Loading Plan

The PAG WRF will be located in the Cunahuiri valley and has been designed with an ultimate storage capacity of 300 Mt of waste rock. This will result in a maximum elevation of the facility at 4335 masl, with a maximum vertical height of 200 m. Development of the PAG WRF is planned in five stages with the objective of reducing haul distances during the initial years of mining. Loading will start at the northern limits of the facility, closest to the pit, and will progress southwest towards the ultimate toe of the facility. Foundation preparation of each stage will involve stripping of topsoil and excavating unsuitable material within critical sectors at the downstream toe of each stage to improve overall slope stability. The hydrogeologic studies and modelling show that the natural groundwater levels and gradients beneath the PAG WRF will provide hydraulic containment for any seepage and direct it to a single reporting point below the southwest toe of the facility. This containment precludes the need for a liner. The design does, however, include a robust underdrain system to collect the seepage from the base of the waste rock and direct it to the reporting point. Seepage from the PAG WRF and groundwater under the facility will be directed to a containment pond constructed downstream of the PAG WRF and retention pond. This pond will also contain surface runoff from the PAG WRF. A 28 m high earthfill embankment will provide approximately 600 000 m3 of water storage. The design consists of a cross valley zoned embankment with a grouted curtain which spans nearly the entire length of the embankment and extends to depths of approximately 40 m into

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rock. Water stored in the pond will be used as process water for the mill after treatment. During infrequent, extreme wet periods, excess water may accumulate in the containment pond and be pumped to the upper surface of the PAG WRF for recirculation through the dump as a means of adding additional temporary storage. Water from the PAG WRF is not intended to be released; however, a valve controlled outlet pipe has been incorporated into the design to be used only in the event of an emergency in which water levels within the pond need to be rapidly lowered. A retention pond between the toe of the PAG WRF and containment pond will be constructed to contain any rocks falling down over the slope of the WRF and to provide for energy dissipation of the drainage flows. A 5 m high, flow-through, rockfill embankment will be constructed at the downstream end of the retention pond, which will provide an approximate capacity of 24 000 m3. Overall slopes for the PAG WRF have been selected to provide stability and allow for easier progressive re-vegetation. A 3H:1V slope was adopted, and will be achieved with the construction of intermediate 32 m wide benches separating 1.4H:1V inter-bench slopes and 20 m lifts. Analyses of the slope of the PAG WRF were completed to confirm adequate slope stability and acceptable deformations under static and seismic loading, respectively. 18.7.2

Tailings management facility

The TMF will be developed behind an embankment dam crossing two broad, gently sloped, south to north valleys above the south side of the Chilloroya River, as shown on Figure 18.29. The site is some 5.2 km southwest of the mine pit and 3.7 km south of the process plant and was selected from an extensive alternatives assessment and ranking study that contemplated 7 alternatives. The TMF has been designed with an overall storage capacity of 277 million dry tonnes of tailings assuming an average in storage dry density of 1.5 t/m3. The embankment will have an ultimate height of 130 m and length of 2300 m, although additional storage can be developed by raising the TMF above this elevation. Instrumentation will be installed in the PAG WRF for performance monitoring. Such instrumentation will include: 

Piezometers in the waste pile and underlying drain



Settlement and deformation monuments on the outer faces of the waste pile



Flow rate and volume meters on the pumped discharge from the PAG WRF containment pond to the process pond.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.29 Tailings Management Facility Final Configuration Plan

The embankment will be aligned in an east-west direction across the outlets of the two valleys and the ridge separating them. It will terminate on the natural ridges to the east and west that provide lateral containment for the impoundment. Initially, the embankment will consist of two separate structures crossing each of the east and west valleys, but after the first year of operations these will be raised over the separating ridge to form one continuous embankment. The TMF embankment will consist of a zoned earthfill structure constructed in stages out of local borrow materials and selected non-PAG mine waste. The design process considered two main alternatives, downstream and modified centreline. After trade-off analyses the modified centreline embankment was selected, based on environmental and economic considerations. A modified centreline embankment involves constructing staged raises so that the centreline of each raise is located slightly upstream of the centreline of the raise below. This is a variation on the centreline approach where the position of the centreline does not change with each raise. By moving the raises slightly to the upstream the quantity of fill in the embankment is reduced while still maintaining a high level of overall stability. Centreline and modified centreline embankments rely on the strength of the adjacent upstream tailings only for local support of the toe of each raise, but do not rely on the strength of the tailings for overall support of the embankment since the large majority of the fill in each raise is placed on underlying fill.

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Notwithstanding the above, during the first two years of operations the embankment will be constructed following a downstream configuration, since the tailings beach will not be immediately developed against the upstream face of the embankment. This will happen over the initial two year period and then from Year 3 onward the modified centreline approach will be adopted. Figure 18.30 presents a cross-section of the embankment showing the staged raising. Figure 18.30 Tailings Management Facility – Typical Staged Embankment Sections

Tailings will be deposited from designated off-take points on a distribution pipeline located along the upstream crest of the embankment, and will be delivered through drop-bar pipes running down the upstream face of the embankment into the TMF. The points of active deposition will be frequently rotated to form a thin layered, drained and well consolidated beach that will slope away from the embankment towards the south side of the TMF basin. Initially, the surface water pond will be located against the embankment in the east valley but it will be displaced progressively upward and to the south by the development of the tailings beach, such that within the first two years of operations the beach is expected to become well developed against the embankment. The surface water pond will vary in size throughout the life of mine depending on the season, precipitation, and operational requirements. The embankment zones will consist of (from upstream to downstream) a compacted, low permeability core, a filter/drain against the downstream side of the core, a wide structural fill zone and a downstream face of erosion protection rockfill. The initial downstream embankment will be covered with a geomembrane on its upstream face, such that, together with the core, a composite liner effect will be created for containment of the initial surface water pond that forms against the embankment. The initial embankment (downstream embankment) will be constructed with upstream and downstream slopes of 2.0H:1V and 2.25H:1V, respectively. The modified centreline embankment raises will be constructed with an overall tailings to embankment fill interface slope of 1H:3V on the upstream side and a downstream slope of 2.25H:1V. The local upstream slope forming the toe of each embankment raise on the tailings beach will be 3H:1V, although this slope may be flattened if necessary. The width of the final embankment crest will be 10 m; however, staged crest widths will be wider to provide for construction efficiencies.

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The east valley embankment will be constructed first to provide initial, temporary storage for water to support the start-up of operations. Random fill for construction of the starter dam will be obtained from in-situ borrow; after that random fill will be obtained from non-PAG waste rock mined from the pit. Core material will be obtained from glacial till located within the upper area of the PAG WRF. Filter/drain material will be obtained from crushing/screening operations within the TMF area or from screening operations from alluvial material in the Chilloroya River. Tailings deposited in the facility will consist of rougher tailings (RT) and cleaner scavenger tailings (CST). These streams will be combined at the plant at an approximate ratio of 4 to 1 RT to CST, prior to transportation and deposition into the TMF. Although the CST will contain significant sulphide minerals that can make it potentially acid generating, the combined stream is predicted to have excess alkalinity from the mill such that its pH will be initially in the order of 8 to 9. Geochemical analyses indicate that an exposure period of six months to a year will be necessary until this alkalinity is consumed and the tailings deposition plan calls for a fresh layer of tailings to be placed over each previously deposited layer well within this period to reduce the potential for acidic conditions developing. At least one and half or two years prior to closure, the CST tailings will be routed through a tertiary flotation stage to de-sulphurise the stream so that the combined RT and CST tailings remain non-acid generating. These de-sulphurised tailings will be deposited into the TMF via the previously used distribution and deposition system to form a covering layer of benign tailings that will serve as a foundation for the permanent closure cover. The small volume of concentrated sulphide tailings removed from the CST will be deposited in a dedicated facility built just prior to this. The concentrated sulphide tailings facility may be constructed within the TMF or near the mill and is expected to comprise a small geomembrane lined basin. Once all of the concentrated sulphide tailings have been deposited in the facility, a geomembrane will be installed over this facility and welded to the lower geomembrane to fully encapsulate the concentrated sulphide tailings. The facility will then be covered by soil and erosion protection layers and contoured to shed run-off. The TMF impoundment includes a geomembrane liner over the base of the entire eastern valley and most of the western valley to provide geomembrane containment in areas where the surface water pond is in contact with the impoundment base at any time over the life of mine. A 50 m wide tailings underdrain is placed on top of the geomembrane against the upstream toe of the embankment to assist in depressing pore pressures in the tailings against the embankment and to minimise the heads on this part of the geomembrane liner. The tailings underdrain comprises a system of perforated pipes installed in drainage layer material and overlain by non-woven geotextile to provide filtration between the tailings and drainage aggregate. The geomembrane liner has not been extended into the upper, southern reaches of the western valley since the surface water pond will never be located there. However, to assist in intercepting and collecting any small amounts of seepage that may pass beneath the western side of the embankment in the absence of the liner, an intercept trench and drain is constructed under the embankment across this valley. Groundwater drains will be installed under the geomembrane to intercept groundwater seeps and keep them separate from the geomembrane. These drains will consist of perforated pipes installed in trenches, backfilled with drainage aggregate and encapsulated within a non-woven geotextile wrap.

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A separate foundation drain system will be installed under the embankment and will consist of perforated pipes installed within the horizontal drainage blankets of the east and west valley bottoms. This foundation drain system will collect water seeping through the embankment, which may include direct precipitation on the embankment and/or small amounts of seepage passing through it, as well as localised groundwater seeps in the embankment foundation. Water collected by the drain systems will be conveyed to sumps located immediately downstream of the embankment in each of the east and west valleys. The basin groundwater drain and tailings underdrain outlet pipes in each valley will be installed together in a common reinforced concrete encasement under the embankment. At the sumps, monitoring and control systems will allow for automated water quality and flow rate determinations to be made for release either to the Chilloroya River or for pump back to the TMF based on the water quality. Two sumps will be constructed, one in each of the valleys. Construction of the TMF will require the removal of wet soils composed of organic materials and peat (bog) in the bottom of each of the two valleys. The estimated quantities of these materials are approximately 1.3 and 0.8 Mm3, in the east and west valleys, respectively. The removal process is expected to consist of water drainage, sediment control and excavation of the bog materials. Following excavation, the materials will be hauled and placed into a dedicated storage facility located approximately 1.5 km north of the TMF. During the initial stage of construction before the embankment is in place, sediment control measures will be taken to contain the excavation works and limit the potential for organic materials to migrate downstream into the Chilloroya River. Instrumentation will be installed in the TMF to monitor the structure throughout the life of the facility. Settlement and deformation monuments will be installed on the embankment with each major stage of construction, at an approximate spacing of 250 m along the embankment crest, to monitor any displacements and confirm that they do not exceed acceptable tolerances. Additionally, vibrating wire piezometers will be installed within the following materials: 

Deposited tailings



Tailings underdrain system



Embankment fill



Horizontal drainage blanket



Bedrock foundation underlying the embankment.

Regular monitoring of the piezometers will be performed to gain an understanding of the overall performance of the embankment and underdrain systems over the life of the facility. Pore pressure data will also be used to monitor the tailings behaviour and rate of consolidation as the facility is raised. Periodic geotechnical investigations using cone penetration techniques will also be conducted in the tailings to confirm the strength, pore pressure and seismic liquefaction potential characteristics of the deposit. Flow rate and flow volume totaliser measurements will be made of the various seepage or drainage flows, as well as flows from the surface water pond to the mill or TMF buffer pond and Chilloroya River.

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Tailings beach and surface water pond level measurements will also be made at regular intervals in the TMF. 18.7.3 Topsoil and unsuitable material stockpiles Topsoil removed during initial construction of project facilities will be stored, as needed, for future reclamation works. The need for storage is not anticipated until after the first two to three years of production based on a topsoil mass balance calculation. Two main stockpiles for storage of unsuitable materials and topsoil are planned for the project. Design of these structures included a combined stockpile (Unsuitable Material/Topsoil Stockpile No. 1), located downstream of the PAG WRF containment pond, which will provide storage capacity for bog, topsoil, and unsuitable materials. A second stockpile (Topsoil Stockpile No. 2) strictly for topsoil will be located directly north of the PAG WRF. The design of the combined stockpile consists of a cross-valley constructed embankment that will provide approximately 2.65 Mm3 of total storage capacity. During initial construction of project facilities, bog, topsoil, and unsuitable materials will be stored together in the combined stockpile. Following this, at higher elevations, topsoil will be stored for future use in reclamation activities at the end of the project life. It is estimated that of the 2.65 Mm3 of total storage capacity, approximately 0.8 Mm3 will be available for reclamation as uncombined topsoil material. Additional storage capacity will be available if more topsoil material is to be saved. The topsoil stockpile directly north of the PAG WRF was designed predominantly for storage of topsoil removed from the pits and PAG WRF. The design of the facility utilises natural topography and the northern face of the PAG WRF for containment. The capacity of the stockpile is approximately 0.4 Mm3, which, together with the topsoil of the combined stockpile, accounts for approximately 1.2 Mm3 of total topsoil that will be available for reclamation. 18.8

PORT AND TRANSPORT

18.8.1

Introduction

Constancia concentrates will be transported by truck from site to the port of Matarani, which is the closest port with established facilities, and which currently handles mineral concentrates from other mines. Hopper-type trucks with a closed-cover system will be used to transport the concentrates. Each truck would have a hauling capacity of 35 t. The travel distance from Constancia to Matarani port is 475 km by road, made up of 232 km gravel surfacing and 243 km asphalt. Travel time per truck is estimated at two days, returning empty to Constancia. There is, therefore, a need to load up to 32 trucks per day to transport the maximum monthly production of approximately 33 800 t per month at peak production in Year 3. Average LOM monthly production of 21 000 t will require loading of 20 trucks per day or a trucking fleet of about 45, allowing for maintenance.

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The concentrate will be stored at Matarani in a facility provided by International South Terminal S.A. (TISUR), in a roofed storage area, for export to its final destination in Asia. 18.8.2

Management of concentrates

Concentrates are stockpiled in a purpose-built building located adjacent to the concentrate filtration plant. Additional outdoor storage is available to manage out-of-specification concentrate. The concentrate stockpiles are handled by FEL. A load cell totaliser is required on the FEL to ensure that the trucks are loaded within reasonable transport limits before going to the truck scale. Concentrate at the mine site is loaded onto a truck using a FEL. The trucks would be capable of carrying up to a maximum of 35 t of product. Concentrate spilled during truck loading operations is contained within a concentrate paved area and all trucks are washed with high-pressure sprays prior to being despatched. This procedure minimises the possibility of environmental contamination and concentrate losses. 18.8.3

Transportation

The transportation of concentrate from the mine to Matarani would be undertaken by a specialist haulage contractor, under a competitively bid haulage contract. The trucks will be a hopper type with a closed-cover system. A convoy of four to six trucks will be escorted to the Matarani port. The transportation company will provide a communications system that includes cellular radio, satellite communications and GPS which will allow for continuous control monitoring of the truck fleet through the entire transportation circuit. Speed, time and distance would be monitored for each truck unit during a 24 hour period. At peak production in Year 3, the truck fleet must guarantee transport of 9000 t of concentrate per week, operating 7 days a week. This represents the need for a maximum of approximately 64 trucks operating on the route, assuming that each truck would transport up to 35 t load, with a LOM average of 40 trucks. 18.8.4

Port management

The concentrate will be stored in the Matarani storage warehouse owned and operated by TISUR, a private organisation that has held the port operations contract since 1999. Its contact term is 30 years with an option to extend for another 30 years. TISUR maintains the following international certifications: 

ISO 9001 (Quality)



ISO 14001 (Environment)



OSAHS 18000 (Security)



BASC (the USA Custom)



ISPS Code (IMO).

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TISUR is presently providing concentrate storage services for Phelps Dodge’s Cerro Verde and Xstrata’s Tintaya mining operations. Reception and unloading at the warehouse consist of receiving the concentrates on a hopper, weighing it, and unloading in a manner that ensures that technical and environmental conditions are met. The concentrate is deposited in a roofed storage area which can be stacked to a maximum height of 7 m. The storage facility will be sized to equal the monthly production for the Constancia project (average 19 000 tonnes of concentrate). Load-out from the concentrate storage area is controlled by a 1500 t/hr loading belt. The concentrate is delivered via the loading belt to the ship loader, where it is deposited at the bottom of the ship holding cell. Loading is conducted 18 hours per day, 365 days a year. The storage and ship loading of concentrate would be contracted to TISUR. Correspondence between TISUR and Norsemont indicates that port handling charges of $7.50/t (wet) will apply. A “take-or-pay” contract will be applied based on minimum annual tonnage. Norsemont may elect to enter into a nominal 10 year contract with TISUR. After this they may negotiate a new contract to obtain more favourable terms to reflect the lower tonnages of concentrate produced in the final years of mine-life. 18.9

PROJECT IMPLEMENTATION PLAN

18.9.1

Approach and strategy

This part of the Technical Report outlines the proposed project execution methodology from the time of Norsemont board approval of the project through to completion of commissioning. A preliminary project delivery model is provided in Figure 18.31. An Owner’s team will be formed to deliver the project through the engagement of an EPCM contractor and specialist engineering consultants, suppliers and Peruvian construction contractors. The Owner’s team will consist of specialist delivery personnel sourced locally and using expatriate resources where appropriate. The Owner’s team will develop policy and ensure its implementation and compliance through the consultancy, supplier and contractor’s systems in the following areas: 

Safety



Environment



Human Relations



Industrial Relations



Legal



Security.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.31 Delivery Model

Page 342


The Owner’s team will deliver the project through a high level and incentivised contracting strategy, thereby reducing interface management and minimising duplication of roles. The team will have integrated systems and procedures with a specialist EPCM provider responsible for the delivery of the process plant and associated infrastructure. The EPCM provider will provide the underlying framework for systems and procedures for the Owner’s team. The Peruvian construction industry has sufficient capability and capacity to carry out the construction works. Foreign contractors would only be used in either high risk or specialist work areas, for example mill construction, lining, etc. The contracting strategy aligns with in-country contractors’ capabilities and local industry practices. The strategy will allow a competitive tendering environment, whilst providing sufficient flexibility to maintain control of the project. Fixed price contracts are preferred; reimbursable contracts, where used, will be structured with incentive clauses to encourage performance. Both practices support a reduced level of performance management by the Owner. The Owners team will deliver the mining development work, and, ultimately, this team will transfer through to the operations team. The Owners team will cover all aspects of the mining development which will include areas such as mine design, mine fleet selection and procurement, assembly of the mining fleet, and operations, etc. All the tailings, waste rock, topsoil and water management facilities will be engineered by a single consultant to minimise interface management. Bulk earthworks across the site will be delivered by a civil construction fleet purchased and managed by the Owner. This fleet will commence construction early in the project and be retained through to operations to continue with subsequent construction stages of the tailings management facilities. Construction of the balance of the works will be contracted to local specialist contractors. The access road will be completed in two phases, the first stage involving detailed design, including government agreement and approvals, and the second stage construction. The detailed design will be provided by a specialist Peruvian engineering consulting firm, and the construction by a specialist civil construction contractor. Power supply will be delivered in two phases, the first stage including design, government agreement and approvals, land and acquisition; and the second stage being construction and commissioning. The detailed design will be provided by a specialist Peruvian engineering consultancy firm and the construction by a specialist construction contractor. The accommodation camp will be delivered through a single supply and install contract, with the management under a separate management contract. 18.9.2

Quality assurance

A project Quality Plan will be developed for the Project. This Plan will form an integral part of the overall Project Execution Plan and detail how quality assurance will be achieved during engineering, procurement, construction and commissioning.

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18.9.2.1

Engineering

All detailed engineering documentation following the FEED package will be undertaken in Spanish. Engineering consultancies will undertake engineering in accordance with ISO 9001 and compliant with the Owner’s Project Quality Plan. Engineering systems will inherently provide design review and verification. HAZID and / or HAZOP studies will be undertaken to ensure the operational integrity and safety of the process plant systems. 18.9.2.2

Procurement, construction and commissioning

Construction contractors and suppliers will be contractually responsible for the quality of their products and will be required to implement accredited quality systems in support of their quality control and assurance program. Each supplier or contractor will be required to develop project specific Quality Plans responding to the Owner’s Project Quality Plan. Vendors will be required to have quality systems accredited to ISO 9001 or otherwise demonstrate that their design, manufacturing and construction activities comply. The level of individual vendor inspection and audits will then be determined on a case by case basis. The EPCM provider will ensure the contractors’ compliance to their Quality Plans through an auditing process. Supervisors will monitor construction quality on a daily basis in the field. Independent testing and verification will be undertaken on a case by case basis. Quality assurance and control of the works undertaken by the Owner’s civil construction fleet will be undertaken independently by the design consultant. The Owner, however, will retain the construction risk. 18.9.3

Project construction fleet

Significant construction works are required to prepare the plant site, TMF, WRF, water management structures and other site infrastructure. Norsemont has decided to purchase the fleet, part of which is required for on-going development of the TMF throughout the operating life, and for closure. Knight Piésold has estimated construction fleet requirements based on a conventional owner-operated earthworks construction project utilizing hydraulic excavators and off-highway trucks. The construction design criteria are based on designs developed by Knight Piésold and the materials movement has been calculated accordingly. Pre-construction earthworks will last for approximately two years with the tailings embankment construction continuing for an additional fourteen years. Estimates of the horizontal and vertical haul distances for transporting both construction and waste materials have been established for each construction activity. This quantification is important since hauling constitutes the largest single component of construction costs, in terms of both capital and operating costs. For the purpose of this analysis, one size of truck was evaluated for the material haul over the duration of the project. A review of haul speeds was conducted, and therefore reasonable haul times can be estimated, using:

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

33 km/hour average truck speed



Specific haul capacities were calculated for the larger earthwork components



An average haul capacity of 4,000 m3 per truck per day for smaller earthwork projects

Trucks are expected to operate for approximately 5270 hours per year per truck. The equipment is scheduled to operate 15.6 hours per day, 7 days per week, 340 days per year. In addition to the primary equipment, a number of support units has been added which will be a necessary component of the construction activities. These include FELs, dozers, graders, compactors, water trucks, service trucks, and the like. Many of these items will be pre-purchased by the mine and will service both the mine and construction fleets. Most of the construction fleet will be salvaged after three years with the remaining equipment used on the construction of the tailings pond embankment and for plant field maintenance operations. The mine truck fleet will be used to provide the embankment material after the construction fleet has been demobilised. In addition, a separate sub-contractor will be responsible to construct the underdrains, geosynthetics, pipework, and small peripheral construction activities during the first three years and smaller earthworks at the end of the mine life. In addition to primary construction activities, aggregate material is provided for tailing dam construction, and possibly surfacing rock for mine haul roads. This material will be drilled on a tight pattern and blasted hard to produce finer grained aggregate material. The blasted material is coarse-screened through a 0.5 m grizzly, loaded and hauled to a designated final dumping point. Much of the required tonnage is required in the two pre-production years prior to start-up of the main mining activity. The proposed construction fleet is shown in Table 18.24.

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Table 18.24 Construction Fleet Requirements Type

Hydraulic Excavator

Equipment

Fleet

Class

Units

382 kW

1

Operating

Operating

Unit Purchase

Hours

Costs

Price

(Hrs/yr)

($/hr)

($M)

5304

149.61

0.827

Wheel Loader with Rock Bucket

597 kW

1

5304

135.18

1.952

LPG Dozer with VR Blade and Ripper

192 kW

1

5304

47.16

0.478

Dozer with CU Blade and Ripper

482 kW

1

5304

92.23

1.227

Soil Compactor

299 kW

2

5304

43.86

0.644

Vibratory Soil Compactor

130 kW

1

5304

43.86

0.113

Articulated Truck

30 T

1

5304

30.85

0.413

Dump Truck

60 T

8

5304

149.02

1.004

Light Duty Dump Truck

19T

1

2652

73.01

0.070

Fuel and Lubrication Truck

43 T

1

1326

8.12

0.585

Water Truck

60 T

2

5304

149.02

1.166

Service Truck

1

1326

8.72

0.175

Tire Truck

1

1326

10.77

0.175

1

5304

87.79

2.083

Motor Grader

397 kW

Utility Backhoe – 4WD

102 kW

1

1856

50.50

0.156

¾ ton Pickup Truck, 4WD

6

5304

5.27

0.027

Onan Light Plant

6

2122

1.68

0.027

Hydraulic Excavator

200 kW

2

5304

131.40

0.445

Carry Dozer with Ripper

698 kW

2

5304

118.85

1.753

Wheel Loader

260 kW

1

5304

39.02

0.395

Motor Grader

221 kW

1

5304

37.72

0.706

18.9.4

Project implementation schedule

A project implementation schedule is summarised in Figure 18.32. The schedule shows a total project duration of approximately 39 months, including detailed design, procurement, construction and commissioning.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.32 Project Implementation Schedule Year Quarter

2009 Q3

2010 Q4

Q1

Q2

2011 Q3

Q4

Q1

Q2

2012 Q3

Q4

Q1

Q2

Q3

Q4

APROVALS Finance Approval ESIA Construction Permit MINE Planning Tender and Award Fleet Supply Assembly /Training Strip and First Ore CIVIL CONST. FLEET Planning Tender and Award Fleet Supply PLANT Design Long Lead Procurement Construction Commissioning TAILINGS STORAGE Design Mine Access Road and Camp Process Plant Pad Mine Area Pad Tailing Management Facility – Initial Stage Tailing Management Facility – Reservoir Fill Tailing Management Facility – Embankment Other Operational Facilities Waste Rock Dump ACCESS ROAD Detailed Design Tender and Award Construction ACCOMMODATION Design and Tender Procurement Complete 500 rooms Complete 1000 rooms Complete 1500 rooms Complete 1800 rooms

X X X

POWER SUPPLY Design Tender and Award Long Lead Procurement Construction

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Key dates for the project are as follows: 

Project Execution commencement – Q3 2009



Environmental approval – Q3 2010



Construction permit issued and commencement of construction – Q4 2010



Completion of Process Plant construction – Q2 2012



Completion of Process Plant commissioning – Q4 2012.

18.9.4.1

Key schedule assumptions

The schedule is developed based on the following key assumptions: 

No significant familiarisation period is required for key consultancies or contractors and there are no delays in their commencement.



Approvals and permits can be achieved in the allocated durations.



Critical members of the Owner’s team are in place.

18.9.4.2

Critical items

The following activities are early critical activities that need to commence upon Norsemont Board approval, to provide schedule assurance and mitigate project risk: 

Commence negotiations to secure a power supply agreement or option.



Commence the detailed design of the improvement works for the access road upgrade and investigate the acceleration of its delivery.



Commence the Front End Engineering Design, process design, flowsheet verification and optimisation, plant layout, and long lead procurement.



Commence the detailed design and planning for the bulk earthworks and purchase of the Owners civil construction fleet.



Commence the recruitment of key Owners team.



Commence the development of the accommodation camp contract, and investigate availability of second hand camps.



Investigate the availability of new long lead equipment, i.e. cancelled orders, etc.



Commence development of project systems. This includes OHS&E requirements and standards, equipment numbering, asset numbering, document numbering, cost control and reporting systems, document control, and procurement documentation and systems.

18.9.5

Permits and licences

A draft list of permits applicable to the construction and operation of the Project has been developed by Norsemont and is included in Appendix 1. This list is a draft and is subject to change.

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18.10

PROJECT OCCUPATIONAL HEALTH, SAFETY, ENVIRONMENT AND SECURITY

18.10.1 Occupational health and safety The Owner has overall responsibility for the development of Occupational Health, Safety & Environment (OHS&E) and Security policy and its implementation. The Owner’s team will produce a Project OHS&E plan detailing the requirements for OHS&E in full compliance with the laws and regulations of Peru. All contractors, sub-contactors, suppliers, vendors and consultants will comply with these requirements. The Owner’s team policy will ensure the project is delivered to world’s best practice. The EPCM Contractor will actively pursue and promote zero health and safety accidents and incidents throughout all stages of the project. Each contractor will be responsible for the development and implementation of its own safety program developed in response to the Project OHSE Plan, and will be required to designate a senior person who will be responsible to implement the safety program. Key elements of the program will include new employee safety induction / site orientation and weekly safety (toolbox) meetings, etc. As an integral part of Project execution, the EPCM Contractor will prepare a Health, Safety and Environmental Management Plan to which the EPCM Contractor and all its contractors and suppliers will adhere. This Plan will include the following: 

Adherence to all Statutory regulations



Adherence to all Client’s regulations



Arrangements for the management of health and safety of the construction work



Standards for the transportation and storage of chemicals, fuels and lubricants



Standards for the protection of local and regional waterways



Standards for the protection of local flora and fauna



Standards for the disposal of waste materials



Emergency response plan



Monitoring systems to verify compliance with the Health, Safety and Environment plan.

Workers will be provided with a safe place to work. Proper inspection, guidance and strict adherence to safety policy will ensure that safe working conditions are maintained. The EPCM Contractor will employ a full-time site safety coordinator and will ensure, through regular audits that each contractor implements its own safety program in accordance with safety procedures established for the Project. 18.10.2 Security The Owner will manage site-wide security. Security services will be contracted for the duration of construction. Approximately 50 persons will be required to provide coverage of the site 24 hours per day, seven days per week.

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The accommodation camp, process plant and mine workshop area will be enclosed with security fencing. Lighting and security cameras will assist with providing a high level of security across the site. Mobile security patrols will maintain security for remote construction areas. The project will have a single guarded entry point at the process plant site. The guard house located at this location will control access 24 hours per day, seven days per week. This guard house will also control access between the accommodation camp and the process plant and other construction sites. 18.11

PROJECT OPERATIONAL PLAN

The Operations Plan reflects the broad principles on which the plant would operate. These relate to the mining, maintenance and supply strategies, transport of concentrates, manning requirements and employment philosophy. In general, the project will operate in accordance with the Equator Principles. The Owner will require its operations to meet the prescribed standards of environmental care and interaction with local peoples and communities. The Owners company arrangements will include an office in Lima and corporate headquarters in Canada. The mine and plant operations will be fully managed from the site. Site operations will generally be run by the Owners staff. Mining will be done by an Owner-operated mining fleet mining, and the plant operated and maintained by the Owner. Although the plant will have fully equipped maintenance workshops, where possible, maintenance will be performed off-site by suitable contractors (e.g. motor re-winds). The maintenance strategy reflects capital cost reduction measures adopted in the DFS; specifically these are based around limited provision of standby equipment and the use of mobile cranes in place of gantry cranes for maintenance throughout the concentrator. In general it is thought that better efficiencies can be obtained by outsourcing services wherever possible. This approach will be applied to supply contracts for fuel and lime in particular, as these commodities will be used in bulk. Transport of concentrates will be contracted to a freight company that would supply and operate the trucking fleet. Storage and ship-loading of copper concentrate will be contracted to the port operator (TISUR) at Matarani through a long term “take-or-pay” contract. Manning levels will be determined by the functional requirements of the operation. Additional positions for site labour would be filled to satisfy Peruvian labour laws and local agreements with the community. The manning structure will be made up of management and administration staff, technical staff, operators, trades and general labour. The labour numbers will be adjusted to satisfy local requirements, but it is estimated that total numbers employed on the site will approach 800. The employment philosophy is based on preferential hiring of local people to fill general labour and operating positions. Where possible, skilled labour will be employed from local villages and surrounding townships. First preference would be given to Peru Nationals to fill managerial and technical positions, although it is anticipated that some of these may have to be filled by expatriate staff initially.

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Training is considered to be a key activity before plant start-up, with the need to prepare the local workforce in areas of plant operation, maintenance, operating procedures and workplace health and safety. Where possible contract services will be used to develop training materials and deliver training. Community relations and land management were considered to be highly significant to the operation of the site. Norsemont has provided input to this area of the DFS and has had an active development plan in place since commencement of the project. Community relations were taken into account in the employment strategy, which aimed to utilise and train local people wherever practical. An Environmental Management Plan will be implemented as part of the overall Operation Plan. This is currently under development by Norsemont, and forms part of the ESIA. 18.12

ENVIRONMENTAL CONSIDERATIONS

18.12.1 Legal framework The objective of current mining legislation in Peru is to ensure that new mining operations incorporate measures to prevent or mitigate foreseeable impacts to the surrounding environment, and that discharges meet the applicable effluent and emissions limits. The project has been designed to comply with all relevant legislation applicable to the development of mining projects in Peru including mines, roads, port and transmission lines. Additional legislation that has been considered includes legislation and regulations regarding archaeological areas of significance, endangered and protected species as well as community relations and public disclosure programs. The ESIA is being developed in accordance with the following legislation and standards: 

Peruvian Political Constitution, 1993



Environmental General Law, 2005



Private Investment Growth Law, 1991



Environmental Impact for Works and Activities Evaluation Law, 1997



Sustainable Utilisation of Natural Resources Organic Law, 1997



Environmental Impact Assessment National System Law, 2001



Water Resources Law, 2009



Water Quality National Standards, 2008



Air Quality National Standards, 2008



Noise Quality National Standards, 2003



Health Law, 1997



Environmental Management National System Law, 2004



Cultural Heritage Protection Law, 1985



Solid Waste Law (2004) and modifying (2008)



Forestry and Wildlife Protection Law, 2001

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

Mines Closure Law, 2003



Public Participation Regulations, 2008



Regulation for Environmental Protection in Mining - Mineral Processing Activities, 1993



Maximum Permissible Levels for Liquid Effluents from Mining – Mineral Processing Activities, 1996



IFC/World Bank Social and Environmental Performance Standards.

The Ministry of Energy and Mines (MINEM) is the principal regulatory agency responsible for permitting mining projects in Peru. Changes are underway in Peruvian environmental legislation with the recent creation of the Ministry of Environment. In the future, this new entity will be responsible for monitoring, controlling and permitting all mining projects in Peru, however, this process is still in the developmental and review stages. 18.12.2 Environmental impact assessment 18.12.2.1

Introduction

Knight Piésold was contracted to manage the Environmental Impact Assessment and overall completion of the ESIA for the Constancia Project. Social Capital Group was contracted to complete the Social Impact Assessment, stakeholder mapping and land ownership surveys. The ESIA is in its final stages of completion and the baseline information collected will be used to assess potential impacts and to develop the appropriate mitigation and management plans. The ESIA will consist of the following chapters: 

Executive summary



Background information



Legal framework



Environmental description (environmental baseline)



Socioeconomic baseline



Project description



Alternatives analysis



Impact identification and analysis



Environmental Management Plan



Environmental Monitoring Plan



Resettlement Action Plan



Community Relations Plan



Communications and Public Participation Plan



Conceptual Mine Closure Plan



Cost-Benefit Analysis.

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18.12.2.2

Environmental description

The area has a mining history that dates back 50 years to when Mitsui operated a copper-zinc mine in Katanga near the community of Uchucarco. Mitsui sold the mine to a Peruvian company which abandoned the mine in the late 70s, leaving a legacy of environmental contamination. Tailings were discharged into the river basin, and the open pits and waste dumps were never reclaimed. People from the community of Uchucarco continue to mine (illegally) the abandoned mine workings at Katanga. Near the community of Chilloroya, the local population is engaged in the illegal mining of superficial gold. There is documented evidence that the community employs the use of child labour, and levels of mercury found in sediments near a rudimentary gold processing area exceed international standards for the protection of aquatic life. Studies completed as part of the ESIA have documented elevated blood levels of mercury in local children. Environmental baseline studies were initiated in 2007 by Vector Perú S.A., and updated by Knight Piésold during 2008-2009. All baseline study fieldwork has been completed (wet and dry season data). The area of study included the mine, transmission line right of way and the concentrate transportation route. Power for the project will be supplied by the construction of an 82 km long high tension line (138 kV) from the Tintaya sub station to Constancia. Concentrate will be transported by road from Constancia to the Matarani Port via Livitaca, Tintaya, Imata and Yura. Upgrades and improvements will be completed to 70 km of the existing road from Constancia to Tintaya. Physical environment Location, topography and physiography Constancia is located in the Central Andes of Peru, 634 km southeast of Lima and 112 km south of the city of Cusco, at an altitude that varies between 4000 and 4500 masl. The Chilloroya basin, where the Project is located, was formed by glacial activity, and contains creeks and valleys with typical U-shaped sections. In general terms, the creeks flow into the Chilloroya river, that discharges into the Apurimac river, which flows toward the Atlantic Ocean. Climate and meteorology The climatological database includes information collected from the Constancia meteorological station during 2007 and 2008. The climate of the area is cold (median annual temperature of 6.2°C), with frequent frosts and wide daily thermal variation. Annual rainfall varies from 530.4 mm (Imata Station) to 1 118.7 mm (Livitaca station). There are two distinct seasons, namely dry winters (May through September) and wet summers (October through April). The predominant wind direction is from the NW with an average annual velocity of 3.1 m/s (Figure 18.33). Further details are provided in Section 5.2.

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.33 Wind Rose

Air quality Air quality (SPM, PM10, PM2,5, CO, NO2, SO2, O3, H2S, and metals in PM10) was monitored in the area of the Project and in the communities of Chilloroya and Uchucarco (monitoring completed in April and July 2007; July and November 2008; and January 2009). Results indicate a maximum particulate concentration (PM10) of 103 g/m3, well within applicable Peruvian standards. Air quality within the project site and the surrounding communities is generally good. Noise and vibration Baseline levels for noise and vibration were measured in the communities of Uchucarco, Chilloroya, Coporaque and Yauri. Results from the baseline study indicate that the levels of noise and vibrations from static sources are within national standards. Seismicity The Project site is classified as an area with medium seismicity, according to the Peruvian Seismic Research Center. There have been no local events registered in the Project area during Norsemont’s time.

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Soils The majority of the soils in the study area are of irregular morphology, variable texture and contain considerable cleared gravels. On the hills and slopes, the soils contain colluvial–alluvial and residual materials. In general the erosion potential of the soil in the area surrounding the project is low to medium. The topography of the mine site consists of rolling hills containing native grasses (ichu) that are typically found in the sierra of Peru. During the construction phase, erosion and sediment control techniques will be applied. Hydrology The Project is located in the Chilloroya river basin, a tributary of the Velille river which empties into the Apurimac river, and flows to the Atlantic Ocean. The Chilloroya river basin is approximately 301.84 km2. Major uploads to the river basin occur during the wet season between January and April. The average annual flows in the Chilloroya River are estimated between 6.52 and 5.10 m3/s, as recorded in Section 5. Water quality The following sub-basins were included in the study area: 

Chilloroya river



Pincullune creek



Sacrane creek



Huayllachane creek



Casasuma creek



Arocoyo creek



Canrayoc creek



Chonta – Orcochiri creek



Velille river



Ccatunhuaycco creek.

A total of 53 surface water quality stations were located in the Chilloroya river basin. Surface water quality sampling stations were also established in the Velille and Huancane river basins (Figure 18.34).

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Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.34 Location of Water Quality Sampling Sites

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A total of 12 surface water and sediment quality stations were established within the project footprint (pit, waste rock and tailings storage facilities) as well as in five wetlands in the area. In general, results indicate that the water quality within the project area has neutral to alkaline pH and in some locations exceeds the national standards for iron, manganese, copper, lead and zinc in the Chilloroya river; iron and microbiological parameters in “Hacienda Fortunia”; coliform in Pincullune – Condormarca creek (Jappococha); copper, manganese, zinc, and lead in Sacrane creek; lead and coliform in Huayllachane and Casanuma – Pumacocha creeks; zinc in Arocollo and Canrayoc creeks; and arsenic, iron, and lead in the Velille river and Ccatunhuaycco creek. Biological environment The vegetation in the area consists primarily of grasslands, and wetlands. Wildlife is represented by fox, deer, rodents, lizards, frogs and snakes. Domestic animals are primarily represented by milk producing cows, sheep and alpaca. The high diversity of bird life is associated with the presence of wetlands, rivers and lakes in the area. Flora and vegetation A total of 494 different species of plants and 10 vegetative formations were identified within the project area. While 18 species of flora are classified as endangered, these are not located within the direct area of influence. Terrestrial fauna The diversity of the fauna was characterised through evaluations of habitat and the identification of specific vegetative formations. The evaluations were completed using transects and fixed observation points, and include direct (field observation and traps) and indirect (tracks, remains, excrements) reconnaissance methods. Important habitats for species of interest (national or international protective status) were documented, geo-referenced and photographed. A total of 96 different species of birds, 19 species of mammals, four species of reptiles, and four species of amphibians were identified within the project area. Of these, five species of birds, three species of mammals and one amphibian are listed as protected species. Aquatic life A detailed hydrobiological evaluation of aquatic environments located within the project area, was conducted. Three species of trout and two species of catfish (bagre and challhua) were found within these aquatic environments. Challhua was found only in wetland areas. One species of fish belongs to the IUCN Red List of Threatened Species. Human interest environment Landscape Landscape is defined as the human perception of nature that is observed in a given moment. It consists of the visible features of an area of land, including physical elements such as landforms, flora and fauna, lighting, weather conditions, human activities (archaeological and cultural) and constructed environments.

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The following elements of the landscape will be evaluated in the impact analysis: geological, geomorphological and hydrological elements and processes, relevant to the landscape; biological and ecological processes and elements; and anthropological elements. In general, the Sacrane, Misayoc, Telaracaca and Chillorya river basins are considered to contain low values of visual quality as the landscape elements are common to the surrounding areas. The Japoccocha, Ccuesoccocha-Yanaccocha, Urasana, Cunahuire – Huayllachane, Pumacocha – Casanuma and Huarmicocha – Condorcocha sectors are considered to contain medium values of visual quality, due to the presence of lakes and the dominant mountain landscapes in these sectors. Archaeological Heritage An archaeological survey was completed in the areas directly affected by future Project facilities and activities. A total of 46 archaeological sites were identified in the area of the future mine site. The process to obtain a certificate of non existence of archaeological remains of significance from the INC (National Cultural Institute) has been initiated. Environmental liabilities from Previous and Current Mining Activities An inventory of mining related environmental liabilities was conducted in all areas where previous mining activities occurred and where current artisanal mining is taking place. The inventory was completed by geo-referencing the locations of these environmental liabilities. Samples were collected and field measurements were taken of mining-related effluents, and material from abandoned waste rock dumps and tailings storage facilities was inspected for evidence of acid generation. Five zones where mining related environmental liabilities exist on the Norsemont mining concessions were identified and include: Sacsa Orcco, Katanga, San José, Chilloroya, and Yanaccocha lake. Traffic Copper concentrate will be transported by truck from the mine to the port at Matarani. As part of the evaluation of impacts a traffic study will be completed to evaluate the impacts from the transportation of Constancia concentrate on current traffic levels. Current levels of traffic will be assessed from a survey that will identify vehicle types, the daily volume of traffic and the kinds of loads being transported along the existing road networks. This study will be completed by SIGT. 18.12.2.3

Stakeholder mapping

Taking into account the political and social context, the key stakeholders that have been identified in the two communities are: 

The Community Assembly, which represents all of the registered dwellers (143 families in Chilloroya, and approximately 527 families in Uchucarco) in the two communities. This is the most powerful group of the Peasant Community, and where important local decisions are taken.



The Governing Board (“Junta Directiva”), which represents the community in front of external organisations and institutions such as the State, NGO´s, other private and public agencies. Its purpose is to review and approve activities that promote community development, settle boarder and land disputes, and enforce community order. It is composed of a President, Vice-President,

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Secretary, Prosecutor, Treasurer, and two other members, who are elected democratically by the general assembly for a two-year term. 18.12.2.4

Status of land ownership in the communities of Uchucarco and Chilloroya

The land administration system in Uchucarco and Chilloroya is consistent with the community system of land possession. This system, which is recognised by the State and National legislation, acknowledges that the peasant communities possess different social dynamics from the rest of society. Despite the 4 changes in the law on this regard , communities at the present time have the right to decide upon their territories. These decisions must be made and undertaken by consensus, through a General Assembly of the community. Community organisations have implemented several means to distribute and use their lands, most commonly through the division of the land into plots among the dwellers (as in the case of Chilloroya and Uchucarco). The plots of land are defined and the boundaries marked in the presence of members of the governing board. The plot of land is then given to the community member. These plots of land 5 are registered with the community and the dwellers become holders or (“posesionarios”) of these lands. The dwellers do not have official title to the lands, but can enjoy the benefits from the land (growing crops, grazing of livestock, construction of a house or dwellings), etc. The dweller is not permitted to make any transaction or exchange of the land. Approximately 20% of households in Chilloroya and 50% in Uchucarco hold a certificate issued by the community organisation that demonstrates the allocation of these lands to the specific household. Eleven percent of households in Chilloroya and 22% in Uchucarco do not possess lands allocated by the community, but do live in the communities. Existing disputes between land holders regarding the size and extent of their lands has been recorded. The project has purchased 4097 ha of private lands to date and has plans to purchase additional lands belonging to the two communities. As a result of the project development, approximately 35 families from the community of Chilloroya will need to be relocated. Norsemont is currently preparing a Resettlement Action Plan (RAP) in accordance with IFC standards on Involuntary Resettlement and Peruvian National legislation. In 2009, Norsemont completed a land ownership survey of households and lands in both communities that will assist in determining the numbers of people that would be affected by the Project development and the

4

In 1993 and 1995 two important changes were made in connection with the community land system. The first was

introduced with the entry into force of the 1993 Political Constitution, which only recognised the unlimited nature of the community land, but at the same time the autonomy of peasant communities was extended giving them the freedom to dispose of and use their lands as deemed convenient. In 1995, Act Nº 26505 (Act on the Private Investment on the Development of Economic Activities in the Land on the National Territory and in the Lands of Peasant and Native Communities “Ley de la Inversión Privada en el Desarrollo de las Actividades Económicas en las Tierras del Territorio Nacional y de las Comunidades Campesinas y Nativas”) was passed. 5

Legally recognised form of land holder.

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potential impacts from the resettlement. displacement. 18.12.2.5

The RAP will address both physical and economic

Impact identification and evaluation

Positive and negative impacts related to the project development phases will be identified and evaluated. Mitigation measures will be proposed and evaluated for their relative level of impact and significance. The environmental impact assessment will identify both quantitative and qualitative changes to the environment as a result of the Project, and will evaluate the significance of these changes. The proposed impacts will be analysed before and after mitigation measures are adopted and the final results will be based on the analysis of residual impacts. The process for the evaluation of impacts will be guided by national and international standards, using impact assessment matrices and predictive models. The evaluation of impacts will consider direct, indirect and cumulative impacts. The following predictive tools will be used in evaluating the impacts: Noise and vibration modelling Modelling the impacts from noise and vibrations will be completed by including information from the baseline study such as topography, meteorology and baseline noise levels. The propagation of noise and vibrations will be simulated from the sources to the receptors (populated areas and sensitive faunal habitats). Estimated levels will be compared to national environmental quality standards. Air dispersion modelling This model uses numeric and mathematical techniques to simulate physical and chemical processes that affect the dispersion of particulates and reactions with the atmosphere. The model is designed to characterise the transport and dispersion of particulates using meteorological, topographical and emission source (particulate and gaseous emissions) information. Air dispersion models are widely used by environmental protection agencies in the United States, Canada and Australia, to control environmental discharges through the identification of sources that contribute to air pollution and assists in designing strategies for control and mitigation of these sources. Air quality models are used to verify that the generation of new emission sources will not lead to non-compliance with air quality standards. The Constancia dispersion model will quantify the effects from the construction and operation phases on the surrounding air quality, with specific emphasis on sensitive receptors (populated areas). Modelling of Visual Impacts Visual impacts will be characterised by developing visual basins. The visual basin analysis graphically represents the extent to which the project’s footprint will impact the visual environment. It includes an analysis of natural barriers that affect visual accessibility to the project area of influence. Viewshed 3-D will be used to develop the visual basins. Rays or bands are used to cover the project area of influence and each ray will identify points that are visible and those that are not. Specific points of view include: 

Points located within the visual basins of the Project site

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

Points close to the edge of specific infrastructure



Representative observation points that are free of visual barriers



Viewpoints of scenic importance close to the project area that are not necessarily within the visual basins.

After identifying the observation points, the visual accessibility will be simulated, and a series of criteria evaluated: 

Scenic importance of the point



Remoteness of the observation point with respect to project infrastructure



Perception of the infrastructure as a function of remoteness, context, accessibility, etc.

Water quality modelling The potential impacts to surface water and groundwater resources in the project area of influence will be controlled and mitigated by specific strategies to ensure minimal impact to the quantity and quality of these resources. A mixing zone water quality model will be used to simulate the impacts of discharges on these water resources. Peruvian water quality objectives as well as guidelines from the USEPA (United States Environmental Protection Agency), the CCME (Canadian Council of Ministers of the Environment) and the WHO (World Health Organisation) and predicted water quality limits from project waste water streams will be incorporated as inputs to the water quality modelling. Socio-economic Impacts The probable social effects resulting from the project in the region include: 

migration of people (influx of workers and migration of farmer peasants)



increase in the demand for housing and an increase in vehicle traffic



improvement in education and health centres



new job opportunities to construct housing, improved infrastructure and social services



changes in land use



restrictions on livestock movement



limited physical and economic displacement of project-affected people within the area of direct influence.

Impact evaluation methodology The impact evaluation methodology that will be used is based on the “Methodological Guide for Environmental Impact Assessment” (Conesa Fernández-Vítora et al., 2003). Knight Piésold has made some modifications to this methodology. The various steps included in the methodology are described below. Verification matrix: Location of effects will be determined by the interaction of project activities and environmental baseline study components.

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Significance of Effects: Specific effects of each project activity on each receptor will be determined. Specific parameters such as character, magnitude, extension, duration, reversibility, accumulation and periodicity will be used to determine the significance of the effects. Significance of Environmental Receptors: The determination of the significance or importance of the environmental receptor in the area of influence is a function of its uniqueness, functional importance and/or conservation status. Significance of Impacts: In this step the significance of the affect on the environmental receptor is evaluated in order to obtain the significance of the impact from the Project activity. Area of influence: The final step in impact evaluation is determining the project’s area of influence. The determination of the area of influence will consider effects from Project activities on the environmental receptors located in the study area. 18.12.2.6

Environmental management plan

The Environmental Management Plan (EMP) forms part of the ESIA, and is a management tool used to ensure that impacts from the development stages of the Project are controlled and mitigated. The EMP for the Project will be designed to: 

Identify mitigation and management strategies



Set objectives and targets



Define performance indicators



Document time frames to achieve targets



Allocate responsibilities and identify the necessary resources for the implementation of the plan



Establish mechanisms to monitor, evaluate and report on progress.

EMPs are important tools for ensuring that the management actions arising from the ESIA are clearly defined and implemented throughout all phases of the Project life cycle. 18.12.2.7

Community relations plan

The Community Relations Plan will be prepared taking into account the following requirements: 

Final determination of the Direct and Indirect Areas of Influence of the project, based on the outcomes of the Social Impact Analysis.



The needs in the construction and operation phases, as determined in the Social Impact Analysis.



The State’s requirements, as expressed in the Community Relations Guidelines of the MINEM and the Prior Commitment Act (DS 042-2003-EM).



The requirements of international financial institutions, taking into account the Equator Principles, IFC Performance Standards, APELL for Mining, and social management standards, such as SA8000 and AA1000SES.

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The Community Relations Plan will include the policies, mission and vision of Norsemont. Based on these principles, specific social programs will be designed for the mitigation and prevention of identified impacts. Generally, these programs address the following key topics: 

Communication and Consultation



Participative Monitoring



Claims and Dispute Resolution



Local Employment



Social Investment



Land Acquisition and Resettlement



Code of Conduct for Workers

18.12.2.8

Cost-Benefit analysis

Norsemont is in the process of completing a cost benefit analysis (CBA) for the Project. A CBA is a relatively simple and widely used technique for deciding whether to make a change. It is completed by adding up the value of the benefits of a course of action, and subtracting the costs associated with it. Costs may be either one-off or ongoing. Benefits are most often received over time. This effect of time is built into the analysis by calculating a payback period. This is the time it takes for the benefits of a change to repay its costs. 18.12.3 HSE management and monitoring plan Norsemont will develop a comprehensive HSE management plan for the Constancia Project. The HSE management system will be designed to: 

ensure HSE compliance



demonstrate that all hazards are appropriately managed



achieve continuous improvement in HSE performance.

Periodic and regular monitoring constitutes a principal component of the HSE plan for the construction, operation, closure and post closure phases of the project. The plan includes direct monitoring of air and water resources, and indirect monitoring of flora and fauna. The monitoring program will provide information for evaluating actual project impacts and the effectiveness of the mitigation measures in place. This will allow for dynamic adjustments to the mitigation plan. Six air quality monitoring stations have been proposed four adjacent to the open pit, and one each in Uchucarco and Chilloroya. Strict measures to maintain air quality will be implemented. This will involve, for example, spraying water on access roads for dust control to ensure compliance with legislation regarding airborne particulates.

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Twelve surface water quality monitoring stations have been proposed in the area. The TMF, PAG WRF, topsoil and unsuitable material stockpiles and all associated ponds will be instrumented for performance monitoring. This will include pore pressures in the tailings and mine waste, as well as in the drain zones and embankment structural zone and foundations. Water flow rates and totalised volumes will also be measured as will the water and tailings levels. Monitoring of slope movements and materials settlement will also be made. Hazardous waste will be separated from common domestic waste. Domestic waste will be recycled whenever feasible. Norsemont will construct a sanitary landfill for the disposal of domestic wastes. Hazardous waste will be stored temporarily in secondary confinement areas prior to removal to designated facilities for resale, recycling or definitive storage in accordance with Peruvian regulations. In all cases, storage facilities for fuel and chemical substances will be designed with secondary impermeable containment. These lined and bermed containment areas will be designed to hold 110% of the capacity of the largest tank to avoid spillage of contaminants. No underground storage facilities for fuel are planned. Sewage treatment facilities will be constructed servicing all project components. Operating procedures, including monitoring discharges, will comply with the corresponding Peruvian standards. Health and safety procedures will be developed in accordance with Peruvian legislation and will be strictly enforced. A restoration program will be developed to re-establish a landscape that is environmentally and aesthetically compatible with the surrounding countryside. All personnel and contractors will be required to comply with the standards and procedures contained in the ESIA for all Project stages. Internal and external audits will be performed periodically to verify compliance. 18.12.4 Closure plan 18.12.4.1

Introduction

In accordance with Peruvian National Regulations for the mining sector, a Conceptual Social Closure Plan must be developed and included in the ESIA. The conceptual plan includes the principal impacts from project closure to the communities in the area of influence and identifies measures to mitigate these impacts. Reclamation and closure of the project will be conducted in accordance with international best practices. The objective of the mine reclamation and closure program is to return mined lands to conditions capable of supporting prior land use or uses that are equal to or better than prior land use to the extent practical and feasible. In addition, long-term stability and safety issues must be addressed as a priority.

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The area to be disturbed and reclaimed encompasses approximately 796 ha. Reclamation and closure activities are to be conducted concurrently with mining operations, to the extent practical, to reduce the final reclamation and closure costs and minimise long term environmental liabilities. The key goals of reclamation and closure are to ensure the physical and chemical stability of the TMF and the PAG WRF. 18.12.4.2

Surface reclamation and revegetation plan

During construction of the TMF, pit and PAG WRF, topsoil and subsoils that provide suitable growth media will be salvaged from the foundations during development of these facilities. The salvaged soils will be stockpiled for later use and seeded to control invasive weeds and erosion. 18.12.4.3

Reclamation and closure by facility

Specific plans for closure and reclamation of each portion of the project area are detailed below. Table 18.25 summarises the surface area disturbance associated with each facility. Table 18.25 Disturbance Areas Area

Perimeter

(Ha)

(m)

Pits, Adits, and Trenches

136.5

3329

Waste Rock Facility

197.7

Tailing Storage Facilities

392.8

Roads

32.9

Ponds

28.2

Structures and Building Areas

7.9

Description

Total Disturbance

796.1

Open pit Reclamation of the pit includes re-contouring, re-grading and re-vegetating of an approximately 30 m wide area surrounding the open pit. For costing purposes it is assumed that this structure will be built during operations and maintained at closure. After closure a pit lake will form. It has been assumed for closure costing that the pit lake water will need to be treated in perpetuity. During the first years of filling, seepage from the PAG WRF will be pumped to the pit and in situ pit water treatment with lime addition will be conducted. The current model estimates filling to occur in 10-15 years, when the pit reaches the spill point (4217 masl). The current closure planning includes long term water treatment with lime and discharges into the Chilloroya only during the wet season when there is sufficient flow in the river for dilution. Waste rock storage facility (WRF) The waste characterisation indicates that much of the waste rock may have a potential for acid generation (PAG) and solute leaching. Closure of the PAG WRF will include covering the facility with an approximately 1.8 m cap, which includes a 1 m thick low permeability layer, a 0.5 m thick drain layer

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and 0.3 m of topsoil/growth media, to inhibit oxidation of waste material and reduce infiltration. The soil cover will be scarified and seeded. The upstream diversion channel of the PAG WRF will remain in place at closure. This facility has been designed to pass the 500 year/24 hour storm event. The current closure concept is to pump the seepage and run-off from the WRF to the pit for in-pit treatment, as discussed above. Tailing storage facility (TMF) The mixed tailing that will be deposited in the first 10 to 12 years of operations will have a high acid generation potential. Analysis of the neutralizing potential suggests that without a significant cover, the time to onset of acid generation of the tailing is approximately one year. The current closure plan includes de-sulfurizing the tailing during the last one and half to two years of operations into an inert tailing with low sulphide sulphur content (<0.1 wt. %). The pyrite float material will be stored in a lined area of the TSF. A spillway will be constructed in the southeast side of the TMF that has been designed to pass the PMP. A soil cover will be placed on top of the tailing beach and the facility and the dam face will be revegetated. Roads Roads within the project, including remnants of old access and/or exploration roads, will be reclaimed. Road reclamation is to include pushing safety berms down and over roads, removing culverts, backfilling ditches, and re-grading areas to re-establish the natural drainage system. Fill will be brought in where necessary. After re-grading is completed, road areas will be scarified and seeded. Infrastructure Infrastructure items will be removed from the project area when they are no longer needed to support mining or process activities. The closure cost assumes that the demolition costs would be covered by the salvage costs of the equipment. Concrete foundations will be buried in place. Scrap metal will be removed from the project area. The mill will be decontaminated. Associated yard areas will be ripped to eliminate compacted soils and regraded. After regrading is completed, previously disturbed areas will be scarified and seeded. Production water wells will be abandoned in accordance with local regulations or transferred to support an approved post-mining land use. Monitoring wells are to be abandoned properly once regulatory officials decide they are no longer necessary for monitoring purposes. Water lines, utility poles, power lines, fuel tanks, generators, transformers and other items remaining in the project area after mine operations cease will be removed from the site and disposed of properly unless they can be used by the communities. The non-hazardous sanitary land fill will be closed by placing an inert cap over the facility and removing any infrastructure (fences, platforms, etc.). 18.12.4.4

Monitoring and reporting

Each year, annual reports will be prepared to document the closure and reclamation activities. Revegetation efforts will be monitored biannually by a range specialist to record vegetation success, monitor erosion, and modify reclamation plans if necessary. Groundwater wells and surface water sites will be sampled quarterly to record post-mining water quality. At the conclusion of reclamation activities, as-built diagrams of the reclaimed features will be prepared for future reference purposes.

Page 366


18.12.4.5

Reclamation schedule and cost

Closure and reclamation activities are anticipated to take place over a five year period. Table 18.26 presents the reclamation cost estimate, including a 25% contingency. The total estimated cost is approximately $38.2 M. Reclamation costs for earthworks were supplied by GRD Minproc assuming owner operated fleets. Closure costs for water treatment are based on similar projects in South America. Water treatment costs assume in-pit lime treatment and does not account for assurances of meeting in-stream standards. Water will be discharged only during the wet season in order to use dilution as a means to meet in-stream standards. The closure costs for water treatment presented in Table 18.26 assume one year of water treatment costs based on a net present value using a 7% interest rate. This fund will be maintained in perpetuity in the annual bond accrual for water treatment. Demolition and removal is assumed to be covered in salvage costs - net cost is therefore zero. Contingency is included in the overall project contingency. Table 18.26 Reclamation Cost Estimate Summary Item Earthwork/recontouring

23 010 000

Water treatment

7 000 000

Revegetation Well abandonment and chemical handling Closure monitoring Demolition and removal Closure studies

18.13

Total Cost ($)

660 000 620 000 3 548 000 317 000 1 250 000

Engineering design and construction plan - 5%

1 821 000

Total

38 230 000

CAPITAL COST

18.13.1 Summary The Constancia Project capital estimate by facility is summarised in Table 18.27. including sustaining capital amounts to $993.6 M of which $846.0 M is initial capital.

Total capital,

Page 367

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.27 Capital Cost Estimate Summary Initial

Sustaining

Total

Capital $

Capital $

Capital $

Mining

120 510 597

45 422 345

165 932 942

Process plant and associated Infrastructure

367 183 057

0

367 183 057

Waste and water management infrastructure

84 164 464

46 739 667

130 904 131

Offsite Infrastructure

87 646 211

0

87 646 211

Access Road

20 170 138

0

20 170 138

Accommodation Camp

43 023 618

0

43 023 618

HV Power Supply

Description

24 452 455

0

24,452,455

Owners Civil Construction Fleet

69 392 661

0

69 392 661

Owners Costs

42 962 242

55 474 281

98 436 523

Project Contingency

74 127 070

0

74 127 070

Total

845 986 302

147 636 291

993 622 592

The capital cost estimate for the Constancia Project has a level of accuracy of ±15% and is expressed in Q1 2009 US dollars. The estimates have been developed by the following organisations: 

Mining – GRD Minproc



Process plant & associated infrastructure – GRD Minproc



Waste and water management infrastructure – Knight Piésold



Access road – SIGT



Accommodation camp – GRD Minproc / budget price



HV power supply – Cesel



Owners civil construction – Knight Piésold



Project contingency – GRD Minproc / Knight Piésold / Norsemont



Owners Costs - GRD Minproc / Knight Piésold / Norsemont.

The sustaining capital estimate in Table 18.27 includes: 

Replacement of mining equipment



Staged construction of the TMF



Staged construction of the PAG WRF



Staged construction of the water conveyance system



General Site equipment replacement



Closure and reclamation

Page 368


The above estimate is exclusive of escalation and IGV tax. 18.13.2 Mining capital cost estimate The mining capital estimate was developed by GRD Minproc and is summarised in Table 18.28. Table 18.28 Mining Capital Cost Estimate Summary Initial

Sustaining

Total

Capital $

Capital $

Capital $

Mine fleet and support

101 508 181

45 422 345

146 930 526

Pre-stripping

19 002 416

0

19 002 416

Total

120 510 597

45 422 345

165 932 942

Description

Further details regarding the equipment and unit costs are provided in Section 18.1. 18.13.3 Process plant and associated infrastructure capital cost estimate The process plant and associated infrastructure estimate was developed by GRD Minproc and is summarised in Table 18.29. Table 18.29 Plant Capital Cost Estimate Summary Description

Initial

Sustaining

Total

Capital $

Capital $

Capital $

Process Plant & Assoc. Infra. (incl. Accuracy Provision)

321 768 037

0

321 768 037

Indirect Cost Estimate (EPCM)

45 415 020

0

45 415 020

Total

367 183 057

0

367 183 057

The process plant and associated Infrastructure capital cost estimate covers the engineering, administration, procurement services, construction, pre-commissioning, and commissioning of the process plant and associated on-site facilities. The exception is the electrical, instrumentation and controls component of the estimate which includes for the distribution of power and communications from the plant site to other facilities site-wide. The estimate comprises equipment supply inclusive of import duties, fabrication, installation and construction costs for the permanent and temporary works. Equipments costs are based on budget pricing sourced from international suppliers with the appropriate currency conversions applied. Fabrication, installation and construction costs are based on budget pricing obtained from Peruvian contractors. The rates provided by the contractors include allowance for inherent risks, including productivity impacts due to the high elevation of the site. Piping has been factored as a percentage of the mechanical equipment supply and installation price, based on GRD Minproc experience. An accuracy provision allowance has been provided in the cost estimate for quantity growth. This allowance accounts for the increase in quantities in developing the design from the current DFS-level to

Page 369


the final constructed product. The quantum of this allowance is based on GRD Minproc experience and is consistent with accepted industry practice. A summary estimate by area and discipline for the process plant and associated infrastructure is included in Table 18.30. 18.13.4 Indirect cost estimate (EPCM) The engineering cost estimate has been developed based on a factor of 6% of the process plant and associated infrastructure costs including accuracy provisions. The management, procurement and construction management cost estimate has been developed from first principles. Typical South American EPCM labour rates have been used in the development of the EPCM estimate. The labour rates account for the cost differential between sources, i.e. professional Peruvian resources are lower cost than professional expatriate resources.

Page 370

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.30 Process Plant Area Summary Capital Cost Total Cost Summary Constancia Cu-Mo-Ag Definitive Feasibility Study A.C.N. 008 992 694

Client: Job No: Job Currency:

Norsemont Mining Inc 60021 United States Dollar

Type of Estimate: DFS +/-15% Estimate Revision Date: Q1 2009

Area

Estimate Option No: Estimate Revision No:

01

02

03

04

05

06

07

08

09

10-18

Earthworks

Civil Works

Structural Steel

Platework

Equipment

Piping

Electrical

Instruments

Buildings

Miscellaneous

B 2

Total

0000

General

4,268,011

1,779,563

10,822,705

0

530,050

2,172,611

71,158,123

1,724,202

10,478,551

0

0010

Crushing

0

8,293,369

713,806

2,042,609

13,962,054

588,067

0

0

323,400

0

102,933,816 25,923,305

0020

Grinding

0

8,607,689

4,253,100

842,014

68,206,102

10,276,165

0

0

1,090,518

0

93,275,588 28,663,327

0030

Copper Flotation and Regrind

0

6,840,713

534,362

760,469

17,079,703

3,304,451

0

0

143,629

0

0032

Moly Flotation and Regrind

0

422,063

2,365

133,034

4,978,766

985,345

0

0

0

0

6,521,573

0040

Copper Concentrate Thickening and Filtration

0

666,024

913,610

1,060,358

13,863,875

2,303,796

0

0

119,691

0

18,927,354

0042

Moly Concentrate Thickening and Filtration

0

341,458

660,602

356,855

1,168,558

335,925

0

0

0

0

2,863,398

0060

Tailings Thickener

0

2,908,703

0

282,743

3,062,037

647,885

0

0

53,196

0

6,954,564

0061

Tailings Water Recovery

0

0

0

0

0

0

0

0

0

0

0

0070

Reagents

0

812,663

721,014

1,319,855

3,469,684

1,580,780

0

0

0

0

7,903,996

0080

Concentrate Storage and Loadout

0

440,607

1,370,965

0

368,168

16,230

0

0

0

0

2,195,970

0090

Water Services - Potable, Raw, Fire etc

0

485,209

0

601,225

580,659

1,524,271

0

0

0

0

3,191,364

0100

Air Services - Supply and Distribution

0

137,748

0

0

1,498,595

744,332

0

0

0

0

2,380,675

0110

Diesel Storage

0

189,294

0

261,626

50,289

7,820

0

0

0

0

509,029

1000

Temporary Facilities & Works

0

625,185

312,180

0

0

0

419,893

0

0

8,220,343

9,577,601

1002

Mobile Equipment (Permanent)

0

0

0

0

0

0

0

0

0

0

0

1003

Vendor Representitives

0

0

0

0

0

0

0

0

0

2,489,850

2,489,850

1004

Spares and First Fills

0

0

0

0

0

0

0

0

0

6,600,000

6,600,000

1005

Commissioning and Startup

0

0

0

0

0

0

0

0

0

856,627

856,627

4,268,011

32,550,288

20,304,709

7,660,788

128,818,540

24,487,678

71,578,016

1,724,202

18,166,820

321,768,037

Total

12,208,985

Total EPCM

45,415,020

Grand Total

367,183,057

Page 371


18.13.5 Waste management facilities and water infrastructure capital cost estimate The waste management facilities and water infrastructure capital cost estimate was developed by Knight Piésold, and is summarised in Table 18.31. Table 18.31 Waste and Water Infrastructure Capital Cost Estimate Summary Initial

Sustaining

Total

Capital $

Capital $

Capital $

Tailings delivery pipeline

18 249 343

2 759 938

21 009 281

Tailings management facility

29 954 931

29 964 564

59 919 495

353 714

0

353 714

Description

Internal access roads Topsoil and unsuitable material stockpiles

1 617 642

0

1 617 642

PAG WRF

1 961 869

1 233 534

3 195 403

Landfill and solid waste management facilities Process water pipeline

110 198

83 040

193 238

16 724 132

2 580 588

19 304 720

Site-wide water management

5 343 607

7 592 492

12 936,099


9 849 029

2 525 511

12 374 540

84 164 464

46 739 667

130 904 131

Total

The capital cost associated with the construction of the waste management facilities and water infrastructure has been estimated in two portions. The first portion, tabled above, has been developed based on budget pricing obtained from suppliers for equipment supply. The material volumes were calculated based on the design and unit rates derived from first principles were then applied to the quantities. This work has been estimated on the basis that it will be carried out by specialist contractors. The second portion of this estimate is the bulk earthworks associated with these facilities. The work is to be carried out by the Owners Civil Construction Fleet and the cost is included in Section 18.13.7. 18.13.6 Infrastructure capital cost estimate 18.13.6.1

Access road upgrade

The access road upgrade estimate was developed by SIGT and is summarised in Table 18.32. Table 18.32 Access Road Capital Cost Estimate Summary Initial

Sustaining

Total

Capital $

Capital $

Capital $

Section 1 – Yauri to Livitaca

14 537 764

0

14 537 764

Section 2 – Livitaca to mine site

2 622 006

0

2 622 006

Allowance for temporary diversions and structure propping

1 000 000

0

1 000 000


2 010 368

0

2 010 368

Total

20 170 138

0

20 170 138

Description

Page 372


The capital cost estimate is based on the feasibility design of the road improvement, bridge works and other drainage system requirements. Material volumes have been calculated based on the design. Unit rates have been derived from first principles and then applied to the quantities to develop item costs. The methodology is consistent with producing an estimate with an accuracy range of ±15%, as required for the DFS. 18.13.6.2

Accommodation camp

The accommodation camp estimate was developed GRD Minproc and is summarised in Table 18.33. Table 18.33 Accommodation Camp Capital Cost Estimate Summary Initial

Sustaining

Total

Capital $

Capital $

Capital $

Permanent & Construction Camp construction

26 471 784

0

26 471 784

Potable & Wastewater Treatment Plants

3 801 834

0

3 801 834

12 750 000

0

12 750 000

43 023 618

0

43 023 618

Description

Permanent & Construction Camp operating costs during construction Total

The estimate is based on budgetary quotes from specialist Peruvian contractors specific to the location. The estimate methodology has an accuracy range ±15%, consistent with the DFS requirement. The capital cost equates to a cost of approximately $16 819 per bed. 18.13.6.3

HV power supply

The HV Power Supply estimate was developed Cesel, and is summarised in Table 18.34. Table 18.34 HV Power Supply Capital Cost Estimate Summary Initial

Sustaining

Total

Capital $

Capital $

Capital $

Transmission line

8 521 734

0

8 521 734

Tintaya substation expansion

2 630 794

0

2 630 794

New Constancia substation

12 119 542

Description

Telecommunications

12 119 542

369 042

0

369 042


2 470 907

0

2 470 907

Total

24 452 455

0

24 452 455

The estimate was developed based on the option to source a 138 kV service from the expanded Tintaya substation, and then upgrade the substation to accommodate the 220 kV service once available.

Page 373


The Directs cost estimate includes the expansion of the Tintaya substation (but not the future upgrade to 220 kV), supply and installation of a new substation at Constancia and the construction of the 70 km HV power line and fibre optic communication line from Tintaya to Constancia. The Indirect cost estimate includes the engineering and construction management associated with the supply and installation of the abovementioned work. Equipment supply costs are based on budget pricing obtained from suppliers. Material volumes have been calculated based on the design and unit rates derived from first principles and then applied to the quantities to develop item costs. The methodology is consistent with producing an estimate with an accuracy range of ± 15%, as required for the DFS. It is noted that the power contractor will not utilise the construction camp located on the mine site. The cost of providing a site camp for the construction and supervision is included within the Indirect cost. 18.13.7 Owners civil construction costs The Owners civil construction fleet capital cost estimate was developed by Knight Piésold with input provided by GRD Minproc on some equipment pricing, and by Norsemont on diesel prices and labour rates. These capital costs are summarised in Table 18.35. Table 18.35 Owners Civil Construction Fleet Capital Cost Estimate Summary Initial

Sustaining

Total

Capital $

Capital $

Capital $

Civil Fleet labour (during construction)

6 105 989

0

6 105 989

Civil Fleet equipment purchase

26 213 266

0

26 213 266

Civil Fleet fuel & Maintenance (during construction)

37 073 407

0

37 073 407

Total

69 392 661

0

69 392 661

Description

Costs for all major equipment include the FOB price at an off-shore port, ocean freight and inland freight. For those machines not specifically identified by a vendor, a similar approach was followed in order to arrive at an estimated cost at the site for a fully functional machine. Capital costs for crushing and screening have been included in the cost estimate. Most of the construction fleet will be salvaged after three years with the remaining equipment used on the construction of the tailings pond embankment and for plant field maintenance operations. Operating costs during the construction period have also been included, comprising equipment operating costs, operating and maintenance labour. Costs for operating the crusher and screens are included. Norsemont provided labour and overheads rates. It is expected that there will be no equipment replacement during the mine life simply because the operating hours will be within the expected equipment life in most cases.

Page 374


18.13.8 Owners cost estimate The Owners cost estimate for the Project has been developed by Norsemont. It includes both site and Lima corporate office costs, and is summarised in Table 18.36. Table 18.36 Owners Cost Estimate Summary Description Labour - Owners Delivery Team

Initial

Sustaining

Total

Capital $

Capital $

Capital $

11 491 250

0

11 491 250

Disbursements

2 300 000

0

2 300 000

Contributions - Enviro/Socio

1 500 000

0

1 500 000

Insurance - Total Project

5 500 000

0

5 500 000

General Site Equip - Purchase

6 897 713

0

6 897 713

General Site Equip – Replacement

0

17 244 281

17 244 281

HR (recruitment/hire/training costs)

1 500 000

0

1 500 000

834 375

0

834 375

0

0

0

Land Acquisition & Administration

6 039 121

0

6 039 121

Strategic Spares

6 899 784

0

6 899 784

Labour - Operations Team Marketing / Sales

Taxes (excluded)

0

0

0

Escalation

0

0

0

Closure costs - not incl. perpetual water treatment

0

38 230 000

38 230 000

Scope growth allowance

0

0

0

42 962 242

55 474 281

98 436 523

Total

18.13.9 Contingency The Contingency estimate has been developed collaboratively with input from GRD Minproc for the mining, process plant and associated infrastructure and accommodation camp; Knight Piésold for the tailings, waste and water management; and input from Norsemont for the access road, HV power supply and Owners costs. The project contingency has been developed using a risk based approach rather than a traditional approach of applying a simple percentage contingency allowance. The uncertainty of the key inputs such as rates, quantities, major equipment pricing, and external risk and opportunities has been analysed by examining the likelihood of occurrence together with a range of cost implications. The analysis has been used to develop a probabilistic model which generates a probability profile for the project that includes inherent and contingent risks (Figure 18.35).

Page 375

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.35 Capital Cost Estimate Probability Profile

The Constancia Project has chosen a P80 to support its Contingency Estimate. The P80 has an 80% probability that the final expenditure will not exceed the estimate, and a 20% probability that the final expenditure will exceed the estimate. Adopting a P80 figure results in a contingency of $74.13 M or 9.6% of the capital estimate. 18.14

OPERATING COSTS

18.14.1 Summary The Constancia Project operating cost estimate is summarised in Table 18.37. The operating cost estimates associated with the various facilities have been developed by the following organisations: 

Mining – GRD Minproc



Process plant and associated infrastructure – GRD Minproc



General and administration - Norsemont



Off-site– Norsemont and GRD Minproc



Royalty – Norsemont

The operating cost estimates were developed based on materials cost and unit rates from supplier quotations as well as historical experience on similar projects.

Page 376

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.37 Operating Cost Estimate Summary Area

LOM Average

LOM Average

LOM Average

Annual Cost $

$/t Milled

$/lb Cu Metal

Mining

48 405 954

2.62

0.32


71 216 612

3.85

0.48


11 144 200

0.60

0.07

Off-site costs (refining, smelting, transport)

53 533 883

2.89

0.36

Civil Construction Fleet

3 565 905

0.19

0.02

Royalty

7 980 470

0.43

0.05

195 847 024

10.59

1.31

Total

The operating cost estimate for the Constancia Project has a level of accuracy of ±15% and is expressed in Q1 2009 US dollars. All relevant TMF operating costs are included in the process plant and associated infrastructure area, except the costs for future TMF uplifts, which are included in the sustaining capital estimate. On a LOM Project basis, the unit operating cost for the Constancia Project is estimated to be $10.59/t ore, or $1.31/lb of copper produced excluding by-product credits. After inclusion of credits for sale of molybdenum, silver and gold the unit cost reduces to $0.92/lb of payable copper. 18.14.2 Mining - operating cost estimate The mining operating cost estimate was calculated based on material quantities, unit costs supplied from existing operations or derived from various external sources, and testwork data available, including: 

Internal database and experience from similar projects



Comparisons with Peruvian operations in 2008/09



Quotations for supply of goods and services from Q1 2009.

The estimate is summarised in Table 18.38 and was built up based on the following major items: 

Load



Haul



Drill



Blast



Support Equipment



Labour



Other.

Page 377

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.38 Mining Operation Cost Estimate Area

LOM Average

LOM Average

LOM Average

Annual Cost $

$/t Milled

$/lb Cu Metal

Load

3 616 008

0.20

0.02

Haul

19 420 109

1.05

0.13

Drill

3 142 204

0.17

0.02

Blast

10 143 674

0.55

0.07

Support Equipment

7 770 704

0.42

0.05

Labour

3 518 724

0.19

0.02

Other

794 532

0.04

0.01

Total

48 405 954

2.62

0.32

The overall mining unit cost over the 15 year life of mine is $2.62/t of ore processed, excluding prestripping costs. 18.14.3 Process plant and associated infrastructure - operating cost estimate The process plant and associated infrastructure operating cost estimate, including the costs to operate the TMF, PAG WRF and water management systems, was calculated based on material quantities, unit costs supplied from existing operations or derived from various external sources, and testwork data available at the time of writing including: 

Internal database and experience from similar projects



Comparisons with Peruvian operations in 2008/09



Quotations for supply of goods and services from Q1 2009

The process plant and associated infrastructure operating cost estimate was developed by GRD Minproc, with input from Knight Piésold regarding the waste and water management infrastructure, and is a build-up based on the following major items: 

Fuel and Miscellaneous



Labour



Maintenance materials



Reagents



Power



Crushing and grinding consumables

Unit operating costs were developed to reflect the full design throughput operating conditions at 91.3% availability. The operating cost estimate is summarised in Table 18.39.

Page 378

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.39 Process Plant Operating Cost Estimate Area

LOM Average

LOM Average

LOM Average

Annual Cost $

$/t Milled

$/lb Cu Metal

Fuel and Miscellaneous

1 424 332

0.07

0.01

Labour

2 136 498

0.12

0.02

Maintenance Materials

4 985 163

0.27

0.03

Reagents

9 258 160

0.50

0.06

Power

25 637 980

1.39

0.17

Crushing and Grinding Consumables

27 774 479

1.50

0.19

Total

71 216 612

3.85

0.48

Over a fifteen year mine life, Constancia is expected to have an average processing cost of 48 ¢/lb copper metal produced. The major plant costs relate to grinding consumables and power within the grinding circuit, representing 39% and 36% of the processing cost, respectively. These costs are associated with the hardness of the Hypogene ore, which is dominant in the latter part of the mine life. Figure 18.36 shows the variation in operating costs over the life of the project. Average plant and offsite costs range as follows: 

31 ¢/lb copper metal produced in the first five years, when softer, higher copper grade ore is being processed



53 ¢/lb copper metal produced in Years 6 to 10, as the amount of harder, lower grade Hypogene ore treated increases, resulting in increased grinding costs and reduced throughput and copper metal production



70 ¢/lb copper metal produced in the last five years as harder Hypogene ores are treated and byproduct credits decline.

Page 379


Process and Post Production Cost ($/lb Cu Metal

Year by Year Metallurgical Costs

1.4 1.2 1.0 0.8 0.6 0.4 0.2 0.0 1

2

3

4

5

6

7

8

9

10

11

12

13

14

15

Year of OperationPost Production $/lb

Processing $/lb

The overall Process plant unit cost over the 15 year life of mine is $3.85/t milled, increasing over time as shown in Figure 18.37. Figure 18.37 Unit Process Operating Costs Per Tonne Milled

Unit Processing Cost ($/t ore milled)

5.0 4.5 4.0 3.5 3.0 2.5 2.0 1.5 1.0 0.5 0.0 1

2

3

4

5

6

7

8

9

10

11

12

13

14

15

Year of Operation

Page 380


18.14.4 General and administration - operating cost estimate The General and Administration (G&A) operating cost estimate includes both site administration and Lima corporate office costs and was developed based on following approach: 

Detailed labour build up using in-country rates as detailed in Price Waterhouse Coopers (PWC) Salary Survey Report – July 2008.



Percentage of capital annualised for maintenance



Lump sum allowances based on historical information.

The G&A operating cost estimate is summarised in Table 18.40 and details are contained in Table 18.41. Table 18.40 General and Administration Operating Cost Estimate Summary Area

LOM Average

LOM Average

LOM Average

Annual Cost $

$/t Milled

$/lb Cu Metal

Labour

4 538 000

0.24

0.03

Administration

2 229 200

0.12

0.01

Other Costs

4 377 000

0.24

0.03

Total

11 144 200

0.60

0.07

The G&A cost estimate covers Labour, Administration and Other Costs required to support site operations. The Labour cost estimate includes: 

Management



Administration



Community relations



Environment and safety



Medical



Security



Accounting



Procurement and logistics



Camp



General and miscellaneous (cleaners, drivers etc).

The Administration cost estimate includes: 

Flights (FIFO)



Buses (BIBO)

Page 381




Camp



Recruitment



Training



Fuel for general administration vehicles and plant.

The “Other Costs” cost estimate includes: 

Insurance



Telephone



Couriers/post



Permits, water licences etc.



Legal and other fees



Government charges



Conferences



Directors fees



Community relations



Road maintenance



HV power line maintenance



Building maintenance (including camp)



Vehicle maintenance



Office supplies (corporate and site)



Consultants



Software



Travel expenses (other than FIFO & BIBO).

Page 382


Table 18.41 General and Administration Operating Cost Estimate Item

Unit

Rate

US$ Cost/year

Labour Corporate Office Labour Costs Site G&A Labour Costs Unskilled General labour

Allowance

920 000

Salary build-up

3 068 000

Allowance

550 000

Administration Flights (FIFO)

$ per return flight

330

79 200

Buses (BIBO)

$ per return trip

500

60 000

Camp Costs

$/man day

15

1 650 000

Recruitment

20% turnover

1 000

160 000

Training Costs

% labour costs

3

120 000

Fuel costs for general plant

$US/litre

0.66

136 000

Fuel costs for admin light vehicles

$US/litre

0.66

24 000

Allowance/mth

15 000

180 000

0.5

1 200 000

Other Costs Lima Office rental Insurance - Capital and Mining only

% Equipment Capital Cost

Telephone

Allowance

100 000

100 000

Logistics/Freight/Couriers

Allowance

200 000

200 000

Permits, Water Licences etc.

Allowance

100 000

100 000

Legal and other Fees

Allowance

150 000

150 000

Conferences

Allowance

40 000

40 000

Directors Fees

Allowance

400 000

400 000

Community Relations

Allowance

250,000

250 000

Communications e.g. website, adverts etc

Allowance

150 000

150 000

25,000$/month

0

0

500 000

500 000

Allowance

120 000

120 000

% of Capital Cost

217 000

217 000

10,000$/month

120 000

120 000

PPE and Medical supplies

Allowance

100 000

100 000

Consultants

Allowance

350 000

350 000

Software

Allowance

100 000

100 000

Travel Expenses (other than FIFO & BIBO)

Allowance

100 000

100 000

Road Maintenance HV Powerlines Maintenance Building Maintenance (incl Camp) Vehicle Maintenance Office Supplies (Corporate & Site)

% of Installed Capital Cost

Total

11 144 200

The overall G&A unit cost for the operation over the 15 year life of mine is $0.60/t.

Page 383


18.14.5 Off-site operating cost estimate Input regarding transport and port charges, smelter costs and penalties (Table 18.42) was supplied by Norsemont. GRD Minproc reviewed smelter costs and penalties and found them to be generally in line with current costs. Table 18.42 Transport and Off-site Treatment Charges Parameter

Unit

Estimate

Transport - Cu concentrates - truck to port

$/wmt conc

32.30

Transport - Mo concentrates - truck to port

$/wmt conc

77.57

Port charges concentrate

$/wmt conc

7.50

Shipping - Cu concentrates

$/wmt conc

35.00

Smelting charges - Cu concentrate

$/dmt Cu conc

65.00

Refining charges - Payable Cu

$/lb Cu

0.07

Refining charges - Payable Ag

$/oz Ag

0.40

Royalties

% of NSR

3.5%

State Royalty

% of NSR

1.0% - 3.0%

Minera Livitaca and Katanga

% of NSR (capped at $10 M)

0.5%

The off-site operating cost estimate is summarised in Table 18.43 and was developed by GRD Minproc as a build-up from the above costs. Table 18.43 Off-site Operating Cost Estimate Area

LOM Average

LOM Average

LOM Average

Annual Cost $

$/t Milled

$/lb Cu Metal

Shipping (inc Port Charges)

11 777 454

0.64

0.08

Transport

9 100 760

0.49

0.06

Smelting & Refining

31 584 991

1.71

0.21

Penalties

1 070 678

0.06

0.01

Total

53 533 883

2.89

0.36

The overall off-site unit cost for the operation over the 15 year life of mine is $2.89/t. The off-site costs for transport, shipping and smelting charges are a significant item, partly reflecting the semi-isolated location and high altitude of Constancia. Periods of lower grade concentrates result in elevated transport charges, while periods of higher zinc in the concentrate result in elevated smelting charges. 18.14.6 Royalty – cost estimate The royalty is made up of both a State Royalty and the Minera Livitaca and Katanga Royalty. A summary of the royalty operating cost estimate is contained in Table 18.44.

Page 384

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.44 Royalty Cost Estimate Summary Area State royalty Minera Livitaca & Katanga royalty Total

LOM Average

LOM Average

LOM Average

Annual Cost $

$/t Milled

$/lb Cu Metal

7 300 401

0.39

0.05

680 069

0.04

0.00

7 980 470

0.43

0.05

The State Royalty cost for the Constancia Project has been calculated as a percentage of the NSR, as follows: < $60 M = 1% >$60 M & <$120 M = 2% >$120 M = 3% The Minera Livitaca and Katanga Royalty cost has been calculated at 0.5% of the NSR, with a total capped cost of $10 M. 18.15

MARKETING, TREATMENT CHARGES AND PRODUCT PRICING

Norsemont reviewed a variety of market reports and studies for copper pricing, copper concentrate treatment and refining charges and penalties. Based on the cyclical nature of metal prices and due to the unpredictable nature of commodities markets, particularly over recent times, it is impossible to obtain solid metal price quotations. Norsemont has received copper forecasting reports from Brook Hunt (BH) and the following comments summarise the current market conditions: 

Chinese purchasing of copper during the first half of 2009 has partly compensated for the reduction in demand elsewhere. However, there is uncertainty whether China can maintain this pace in purchasing.



Base case mine production capability is expected to decline from 2013 onward as production losses, due to mine closures or diminishing head grade exceed committed incremental expansion or committed new project capacity. Base case mine production capability is forecast to decline by 2.6% p.a. between 2011 and 2020 (inclusive) with mine output falling from a peak of ~18.3 Mt in 2012 to ~14.6 Mt in 2020. At the same time, demand for additional mine production to meet increased market requirements for refined metal will create an implied shortfall in mine output of ~0.2 Mt/a in 2013, 4.8 Mt/a in 2017 and 7.1 Mt/a by 2020.



The long term forecast is that global industrial production (IP) growth will average 3% over the period 2008-2020. Mature economy IP is expected to grow by an average of 1.1% over the period, while developing country IP is forecast to increase by 6.7% p.a.

The long term copper pricing from the BH report is shown in (Figure 17.38) for LME Month of the Day copper. The copper spot price for the year to June 2009 is presented in Table 18.39.

Page 385

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.38 LME Long Term Copper Price Predictions (BH)

Figure 18.39 Copper Spot Price, July 2008-June 2009

Page 386


The BH long term pricing for copper in 2009 is well below the current spot price. Brook Hunt in June 2009 reported a moderate demand for copper from the Chinese which coincides with other reports and a 2009 forecast for copper averaging $4179/t ($1.90/lb). The copper spot price as of 3 September 2009 is $2.83/lb, which is a 30% change in price above the forecast 2009 of $1.90/lb. This uncertainty is also reflected in the treatment and refining charges (TC/RC). Recent agreements between Japanese copper smelters and BHP-Billiton show settlement for mid-year processing fees at a 33% discount, equating to $50/t for smelting and 5 c/lb for refining for the year that began July 1 2009. That compares with $75/t and 7.5 cents for calendar 2009 fees. The assumptions used in the DFS financial analysis take into consideration the market forecasting and look at balancing out the prices which best reflect the existing prices used in the market for similar projects. These are summarised in Table 18.45. Table 18.45 Prices and Charges Used in Constancia DFS Price Cu ($/lb)

$2.00

Au ($/oz)

$800

Ag ($/oz)

$12.00

Mo ($/lb)

$13.00 TC/RC (Copper)

Treatment Charge $/t

65

Refining Charge c/lb

7.0

Refining Charges (per oz) Au

$1.20

Ag

$0.40 Payable Metal

Cu Pay min deduction

96.50% 1%

Au Pay min deduction g/t

98% 1

Ag Pay min deduction

90% 30

Mo Pay Treatment Charge (% Mo price)

100% 12.75%

Page 387


18.16

PROJECT FINANCIAL ANALYSIS

18.16.1 Summary Post-tax analysis of predicted project cash flows for three metal price scenarios is summarised in Table 18.46. Table 18.46 Project After Tax Analysis Commodity Price Scenarios

Post Tax Analysis

Case 1

Case 2

Case 3

Cu ($/lb)

2.00

2.25

2.75

NPV (5%)

$496.8 M

$731.9 M

$1277.5 M

NPV (8%)

$303.6 M

$494.2 M

$931.8 M

IRR

15.5%

19.4%

26.9%

Payback

4 yrs

3 yrs

3 yrs

* NPV is quoted after taxes, royalties and profit sharing, and sunk costs.

Case 1 (Base Case): for NI-43-101 reporting purposes, Norsemont has elected to use the following long-term commodity prices: $2.00/lb Cu, $13.00/lb Mo, $12.00/oz Ag and $800/oz Au. Case 2: $2.25/lb Cu represents the approximate mean analysts’ long-term Cu price assumptions. Other metals remain constant at $13.00/lb Mo, $12.00/oz Ag, and $800/oz Au. Case 3: $2.75/lb Cu represents the mean Cu price forward curve through to mid-2011, which historically has been the most accurate indicator of long-term copper prices. Other metals are based on recent metal prices of $18.00/lb Mo, $14.00/oz Ag and $950/oz Au. GRD Minproc provided the capital costs, operating costs and production plan inputs for the financial analysis and prepared the pre-tax cashflow, but expresses no opinion on the post-tax cashflow and financial analysis prepared by Norsemont. The following financial analysis and discussion is based on the base case metal price assumptions outlined above. 18.16.2 Key project assumptions The financial analysis was based on the production schedule provide in Table 18.47. The production schedule targets higher grade ores early in the project life and stockpiles a total of 49.4 Mt of low grade material.

Page 388

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.47 Production Schedule Units Mine Production Ore

Value

2012 -1

2013 1

2014 2

2015 3

2016 4

2017 5

2018 6

2019 7

2020 8

2021 9

2022 10

2023 11

2024 12

2025 13

2026 14

2027 15

m dmt

326.79

0.82

21.95

22.97

23.14

24.59

25.17

24.62

16.89

22.69

23.21

18.79

19.34

24.96

30.57

22.65

4.43

Cu Zn Mo Ag Au Pb Sulphur

% % g/t g/t g/t % %

0.39% 0.11% 107.75 3.46 0.04 0.04% 2.40%

0.37% 0.07% 42.85 3.77 0.08 0.05% 0.53%

0.58% 0.10% 89.79 4.23 0.07 0.04% 2.28%

0.58% 0.11% 116.42 4.43 0.06 0.04% 3.01%

0.63% 0.15% 152.41 5.07 0.05 0.04% 3.40%

0.46% 0.10% 98.35 4.04 0.04 0.06% 2.42%

0.42% 0.10% 88.24 3.58 0.06 0.04% 2.84%

0.39% 0.14% 96.89 3.63 0.05 0.04% 2.86%

0.34% 0.19% 106.79 3.75 0.04 0.06% 2.55%

0.32% 0.12% 110.23 3.15 0.03 0.04% 2.65%

0.33% 0.11% 158.69 3.17 0.03 0.04% 2.29%

0.34% 0.12% 154.38 2.99 0.03 0.03% 2.07%

0.26% 0.05% 90.30 2.51 0.03 0.05% 1.60%

0.27% 0.09% 100.57 3.13 0.04 0.07% 2.05%

0.28% 0.09% 78.43 2.69 0.05 0.04% 2.10%

0.31% 0.06% 86.38 2.31 0.04 0.03% 1.70%

0.41% 0.03% 108.44 2.72 0.06 0.03% 0.86%

Waste TMM SR

m dmt m dmt

300.03 626.82

12.76 13.58

23.60 45.55

21.80 44.78

21.79 44.93

20.33 44.92

19.67 44.84

20.21 44.83

28.02 44.91

22.10 44.79

21.53 44.74

25.99 44.78

25.52 44.86

19.70 44.65

12.34 42.92

4.67 27.32

4.43

0.92

15.53

1.07

0.95

0.94

0.83

0.78

0.82

1.66

0.97

0.93

1.38

1.32

0.79

0.40

0.21

Stockpile

m dmt

1.17

8.91

7.04

12.76

23.32

43.27

17.74

16.27

21.76

27.87

30.02

33.24

41.56

55.10

Process Plant Feed

m dmt

277.39

-

21.64

21.89

21.45

21.05

19.64

Cu Zn Mo Ag Au Pb

% % g/t g/t g/t %

0.43% 0.11% 116.67 3.69 0.05 0.04%

0.00% 0.00% 0.00%

0.59% 0.10% 90.08 4.24 0.07 0.04%

0.60% 0.11% 119.81 4.50 0.06 0.04%

0.67% 0.15% 162.14 5.25 0.06 0.04%

0.50% 0.10% 108.16 4.35 0.05 0.06%

0.49% 0.10% 98.70 4.02 0.07 0.04%


m m m m m

163.55 71.03 14.59 4.73 23.49

-

0.72 18.18 0.04 1.07 1.62

4.40 13.36 0.89 1.50 1.74

5.24 10.18 3.58 0.55 1.90

6.54 11.28 1.36 0.87 1.00

9.58 6.69 0.88 0.70 1.79

19.86 0.44% 0.15% 107.54 3.96 0.06 0.04% 8.81 6.26 2.11 0.01 2.67

18.35 0.32% 0.18% 102.79 3.72 0.04 0.05% 11.53 2.56 0.92 3.34

17.20 0.37% 0.12% 125.13 3.44 0.04 0.04% 13.81 1.11 0.63 0.02 1.63

17.11 0.38% 0.12% 187.34 3.56 0.03 0.04% 13.98 1.04 0.53 0.00 1.56

16.64 0.36% 0.11% 164.90 3.16 0.03 0.03% 14.65 0.02 0.77 1.19

16.11 0.28% 0.04% 95.89 2.49 0.03 0.04% 15.58 0.14 0.38

16.64 0.31% 0.09% 120.18 3.47 0.05 0.07% 14.58 0.21 0.16 1.69

17.04 0.35% 0.10% 94.08 3.25 0.06 0.06% 13.88 0.13 1.10 1.93

% % % % %

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

91.40% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

% % % % %

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

357,410

409,705

450,871

332,280

304,351

28.86% 3.55% 205.30 2.45 0.57%

28.77% 2.66% 192.40 1.83 0.49%

28.54% 2.72% 199.90 1.60 0.52%

28.57% 2.79% 220.27 1.77 0.91%

28.26% 2.64% 207.54 2.58 0.63%

287,649 27.18% 3.64% 218.82 2.30 0.75%

210,240 25.63% 5.11% 259.52 1.99 1.17%

211,969 27.02% 3.14% 223.20 1.75 0.82%

215,531 27.39% 2.99% 226.17 1.40 0.75%

196,298 28.12% 2.82% 214.12 1.38 0.57%

Process Plant Recoveries Cu Hypogene Supergene Skarn 1 Skarn 2 High Zn Mo


Concentrate Cu Tonnes Cu Zn Ag Au Pb

Mo

Tonnes Mo

Payable Metal Cu Cu Ag Au

Mo

Mo

t:t

dmt dmt dmt dmt dmt

dmt % % g/t g/t %

3,766,311

dmt

44,245

28.09% 2.93% 217.21 2.10 0.79%

%

m m m m

lb's oz's oz's lb's

2,237.79 22.55 0.13 39.02

0.00% 0.00% 0.00%

2028 16

2029 17

2030 18

2031 19

2032 20

-

-

-

-

-

0.00% 0.00% 0.00% 0.00%

0.00% 0.00% 0.00% 0.00%

0.00% 0.00% 0.00% 0.00%

0.00% 0.00% 0.00% 0.00%

0.00% 0.00% 0.00% 0.00%

-

-

-

-

-

-

-

-

-

-

-

60.89

49.40

49.40

49.40

49.40

49.40

49.40

16.86 0.36% 0.07% 96.09 2.52 0.05 0.03% 14.33 1.47 1.05

15.92 0.25% 0.04% 75.97 2.13 0.04 0.03% 15.92 -

0.00% 0.00% 0.00% -

0.00% 0.00% 0.00% -

0.00% 0.00% 0.00% -

0.00% 0.00% 0.00% -

0.00% 0.00% 0.00% -

87.00% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

87.00% 89.00% 89.00% 89.00% 89.52%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

55.00% 55.00% 55.00% 55.00% 55.00%

134,200 28.99% 1.48% 239.48 2.41 1.34%

163,779 27.58% 2.71% 281.94 2.79 1.83%

190,048 27.65% 2.85% 233.40 3.12 1.24%

183,100 28.63% 2.02% 185.69 2.58 0.62%

118,882 29.17% 1.59% 228.14 2.84 0.96%

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

0.00% 0.00% 0.00%

-

2,424

3,606

4,782

3,131

2,666

0.00%

40.00%

40.00%

40.00%

40.00%

40.00%

2,937 40.00%

2,594 40.00%

2,960 40.00%

4,406 40.00%

3,772 40.00%

2,124 40.00%

2,749 40.00%

2,204 40.00%

2,227 40.00%

1,663 40.00%

0.00%

0.00%

0.00%

0.00%

0.00%

-

218.36 2.00 0.02 2.14

249.53 2.13 0.01 3.18

272.37 2.45 0.01 4.22

200.90 2.02 0.01 2.76

181.99 1.73 0.02 2.35

165.20 1.74 0.01 2.59

113.58 1.54 0.01 2.29

121.00 1.31 0.01 2.61

124.77 1.35 0.00 3.89

116.77 1.16 0.00 3.33

82.34 0.90 0.01 1.87

95.51 1.32 0.01 2.42

111.11 1.24 0.01 1.94

110.98 0.91 0.01 1.96

73.39 0.75 0.01 1.47

-

-

-

-

-

Page 389


The following key assumptions and parameters were used in preparing the project cashflow projections: 

Capital costs as described in this study



Operating costs as described in this study and summarised in Table 18.48. Mining royalties are based on percent of NSR (as advised by Norsemont) and G&A, transport, shipping, treatment and refining charges provided by Norsemont



Production ramp-up recovery as illustrated in Figure 18.40



Metals price assumptions were provided by Norsemont and are summarised in Table 18.45



Working capital assumes two months debtors (as advised by Norsemont), representing the estimated time for shipping of concentrates and delivery of payment



Ore treatment at approximately 21 Mt/a for the first 4 years (@ ~0.6% Cu and 120 g/t Mo), tapering down to approximately 17 Mt/a by Year 14, generally at decreasing grades (Figure 18.41), producing a maximum of approximately 120 Kt payable copper metal in Year 3 and declining quite significantly after Year 6 (Figure 18.42). Table 18.48 Operating Cost Summary Units Average Mining Costs

Processing Costs


$/t mined $/t ore processed

1.18 2.62

$/t ore processed $/t ore processed $/t ore processed $/t ore processed $/t ore processed

4.31 3.26 2.84 3.13 3.24

Mining Royalties

<$60M >$60M & <$120M >$120M Minera Livitaca & Katanga cap (max payment)

%NSR %NSR %NSR %NSR $M

1.00% 2.00% 3.00% 0.50% 10.00

Transport & Shipping Charges

Copper

Road Transport Port Charges Shipping Costs Insurance Transport & Shipping Losses

$/wmt $/wmt $/wmt $/wmt %

32.30 7.50 35.00 1.78 0.50%

Molybdenum

Road Transport Other Charges Insurance Transport & Shipping Losses

$/wmt $/wmt $/wmt %

77.57 5.86 1.78 0.00%

Copper

Treatment Charge Price Participation Upper Escalator Lower De-escalator Refining Charges Cu Ag Au Treatment Charge

$/dmt

65.00

Fixed

$M/a

Treatment & Refining Charges

State

Value

Molybdenum G&A

$/lb % $/lb % $/lb $/oz $/oz $/dmt

1.20 0.90 0.07 0.40 1.20 1,461.65 11.14

Page 390


Figure 18.40 Ramp up Recovery Assumptions 120%

Percent of Design Recovery

100%

Cu

Mo

80%

60%

40%

20%

0% 1

2

3

4

5

6

7

8

9

10

11

12

Month

Figure 18.41 Plant Feed Grade Mo

Ag

1.00%

200

0.90%

180

0.80%

160

0.70%

140

0.60%

120

0.50%

100

0.40%

80

0.30%

60

0.20%

40

0.10%

20

0.00%

Mo/Ag/Au Grade (g/t)

Cu Grade (%)

Cu

‐ 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031 2032

Page 391

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.42 Payable Metal Cu

Mo

120000

2400

100000

2000

80000

1600

60000

1200

40000

800

20000

400

Cu Metal (t)

2800

Mo (t)

140000

0

0 2012 2013 2014 2015 2016 2017 2018 2019 2020 2021 2022 2023 2024 2025 2026 2027 2028 2029 2030 2031 2032

A number of criteria were considered in determining the viability of the project, including: 

Pre and post-tax net present value (NPV), assuming a real-terms discount rate of 8% pa



Cash cost per pound of payable copper



Cash breakeven price



Economic breakeven price



Price and cost sensitivity analysis.

18.16.3 Pre-tax analysis 18.16.3.1

Key cashflow assumptions (pre-tax)

The cashflow projections and cashflow evaluations based on those projections have been prepared considering the following schedule: 

Capital works commences Jan 2010 and completed by Jan 2013



Ore treatment commences Jan 2013 and completed by Dec 2027



Working capital is provided for as set out in Section 18.13 (as per Norsemont advice).

The analysis has been conducted on a pre-tax, 100% equity basis. Costs from Jan 2010 are considered project costs for the purpose of the evaluation, and cashflows are discounted back to that date.

Page 392


Escalation and inflation have been excluded. Corporate income tax and statutory employee profit sharing has been excluded (these are included in the calculation of post-tax cashflows undertaken by Norsemont, Section 18.16.4). Financing and the effects of debt have not been considered or assessed, nor has hedging been considered. Sensitivity analysis were undertaken to demonstrate the effect of variations in key parameters on project economics. GRD Minproc has reviewed financial returns at a project level only. The funding structure of the project has not been considered and therefore equity level returns have not been reviewed. 18.16.3.2

Summary of results (pre-tax)

Based on the assumptions set out above, the project pre-tax NPV at 8% is $529.3M and pre-tax IRR is 19.5%. The project is expected to pay back the initial capital after 3 years (pre-tax, undiscounted basis) of production. The cash breakeven copper price (the price at which operating surplus plus sustaining and deferred capital equals zero) is $0.98/lb (Figure 18.43). The economic breakeven copper price (the price at which NPV at 8% equals zero) is $1.55/lb.

Page 393

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.43 Estimated Cash Costs

$/lb Payable Transport Costs Shipping Costs TC's RC's Penalties Marketing Royalties Mining Opex Mining Civil Fleet Processing Opex G&A Other Rehab Rehandle Silver Credit Gold Credit Mo Credit Total (before …

(0.92)

Sustaining & … Total (after … 0.40

0.20

‐

(0.20)

(0.98) (0.40)

(0.60)

(0.80)

(1.00)

(1.20)

$/lb Payable

Over the life of the project a total of $2015 M is anticipated to be paid in on-site operating costs. Onsite operating costs over the life of the project are anticipated to average $0.90/lb of payable copper, while total operating costs (including mining royalty, transportation, shipping, treatment and refining costs) are anticipated to average $1.31/lb of payable copper. After silver, gold and molybdenum credits equivalent to $0.40/lb of payable copper, the cash cost is estimated at $0.92/lb. Capital development costs are estimated to be $846.0 M, while deferred and sustaining capital, including closure costs, are estimated to be $147.6 M. At constant prices over the life of the project of $2.00/lb of copper, $12.00/oz of silver, $800/oz of gold and $13.00/lb of molybdenum, it is anticipated that $4557 M would be realised in NSR, generating $1429 M in pre-tax cashflow (before employee profit distributions). Table 18.49 provides a summary of the key project metrics.

Page 394

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.49 Key Project Metrics Units Project Economics

Pre-tax

$M % Yr Yrs

529.3 19.5% 31-Dec-15 3.00

Breakeven Prices

Cash Breakeven Copper Price Economic Breakeven Copper Price

$/lb $/lb

0.98 1.55

Operating Costs

Mining

$M

726.1

Mining Civil Fleet

$M

53.5

$M $M $M $M $M $M

1,068.2 704.7 231.3 41.4 14.8 76.0

Processing

Total Hypogene Supergene Skarn 1 Skarn 2 High Zn

Tails Dam Water Management Pit Dewatering Monitoring

$M $M $M $M

-

Rehabilitation

$M

-

Rehandle

$M

-

G&A

$M

167.2

Total

Mining Process Plant & Associated Infrastructure Waste Management Facilities & Water Infrastructure Infrastructure Project Contingency Owners Cost

$M $M $M $M $M $M $M

846.0 120.5 367.2 84.2 87.6 74.1 43.0

Sustaining Capital

Total Mining Process Plant & Associated Infrastructure Waste Management Facilities & Water Infrastructure Infrastructure Project Contingency Owners Cost

$M $M $M $M $M $M $M

147.6 45.4 46.7 55.5

Project Cashflows

NSR Royalties Operating Costs Operating Surplus Capital Expenditure Pre-tax Cashflow

$M $M $M $M $M $M

Cost per lb payable Copper

$/lb

Development Capital

Unit Costs

18.16.3.3

NPV @8% IRR Undiscounted Payback

Value

-

4,556.9 119.7 2,015.0 2,422.2 993.6 1,428.6 0.92

Sensitivity analysis (pre-tax)

Table 18.50 illustrates the sensitivity of the project economics to copper price, silver price, gold price, molybdenum price, mining costs, processing costs and capital development costs on a pre-tax basis.

Page 395

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.50 Project Sensitivity Analysis (pre-tax) Pre-tax NPV@8% $M

% Change

Copper

Silver

Gold

Moly

Mining Costs

Processing Costs

Development Costs

Base = $ -20% -10% 0% 10% 20%

2.00 /lb

Base = $ -20% -10% 0% 10% 20%

12.00 /oz

Base = $ -20% -10% 0% 10% 20%

800.00 /oz

Base = $ -20% -10% 0% 10% 20%

13.00 /lb

Pre-tax IRR


Cash Costs $/lb

60 294 529 764 999

9.6% 14.9% 19.5% 23.5% 27.3%

5.00 4.00 3.00 3.00 3.00

0.97 0.98 0.98 0.99 1.00

501 515 529 543 557

19.0% 19.2% 19.5% 19.7% 20.0%

3.00 3.00 3.00 3.00 3.00

1.01 1.00 0.98 0.97 0.96

519 524 529 535 540

19.3% 19.4% 19.5% 19.6% 19.7%

3.00 3.00 3.00 3.00 3.00

0.99 0.99 0.98 0.98 0.97

480 505 529 554 579

18.6% 19.0% 19.5% 19.9% 20.3%

3.00 3.00 3.00 3.00 3.00

1.03 1.01 0.98 0.96 0.94

Base = $ variable -20% -10% 0% 10% 20%

601 565 529 493 457

20.6% 20.1% 19.5% 18.9% 18.2%

3.00 3.00 3.00 3.00 3.00

0.92 0.95 0.98 1.02 1.05

Base = $ variable -20% -10% 0% 10% 20%

635 582 529 477 424

21.2% 20.3% 19.5% 18.6% 17.7%

3.00 3.00 3.00 3.00 4.00

0.89 0.94 0.98 1.03 1.08

685 607 529 452 374

25.4% 22.2% 19.5% 17.1% 15.1%

3.00 3.00 3.00 4.00 4.00

0.98 0.98 0.98 0.98 0.98

Base = $ -20% -10% 0% 10% 20%

845.99 M

The pre-tax NPV and IRR are most sensitive to copper price, which accounts for approximately 84% of payable metal value (Figure 18.44 and Figure 18.45). A 10% change in copper price from the base case of $2.00/lb results in a change in pre-tax NPV at 8% of ±$235 M.

Page 396

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.44 Pre-tax NPV Sensitivity (@8%)

1200

Copper Spot Price

1000

Silver Spot Price

NPV ($M)

800

Gold Spot Price

600

Moly Spot Price

400

Mining Cost

200

Processing Costs

0 ‐25%

‐20%

‐15%

‐10%

‐5%

0%

5%

10%

15%

20%

25%

Development Costs

‐200

Figure 18.45 Pre-tax IRR Sensitivity

35%

Copper Spot Price

30%

Silver Spot Price

IRR (%)

25%

Gold Spot Price

20% Moly Spot Price 15% Mining Cost 10% Processing Costs 5% Development Costs

0% ‐25%

‐20%

‐15%

‐10%

‐5%

0%

5%

10%

15%

20%

25%

The relative sensitivities for copper price, silver price, gold price, molybdenum price, mining costs, processing costs and capital development costs are shown in Figure 18.46, illustrating the impact of a 10% change in base assumptions on pre-tax NPV at 8%.

Page 397

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Figure 18.46 Pre-tax NPV Sensitivity (@8%) Copper Spot Price

Silver Spot Price

Gold Spot Price

Moly Spot Price

Mining Cost

Processing Costs

Development Costs ‐300

‐200

‐100

0

100

200

300

Change in NPV ($M)

18.16.4 Post-tax Analysis 18.16.4.1

Key cashflow assumptions (post-tax)

Norsemont commissioned a Peruvian tax specialist to assist in determining the impact of tax on the project post-tax cashflows and financial evaluations. Tax and depreciation rates provided are summarised in Table 18.51. Table 18.51 Tax and Depreciation Assumptions Units

Value

Corporate Tax Rate

%

30.0%

Profit Distribution to Employees (pre-tax profits)

%

8.0%

Tax Depreciation Rates

Owned

Fixed Plant Buildings

%pa %pa

20% 5%

Leased

Movable Plant Buildings

%pa %pa

50% 20%

Pre-operating Costs 100% deductible in one year

Appendix 2 contains the tax, depreciation and amortisation advice provided by Peruvian tax specialist Picon & Asociados.

Page 398


18.16.4.2

Summary of results (post-tax)

Based on the assumptions set out above, the project post-tax NPV at 8% is $303.6 M and post-tax IRR is 15.5%. The project is expected to pay back the initial capital after 4 years (post-tax, undiscounted basis) of production. Table 18.52 provides a summary of the key post-tax project metrics. Table 18.52 Key Post-tax Project Metrics Units Project Economics

18.16.4.3

Post-tax

NPV @8% IRR Undiscounted Payback

Value

$M % Yr Yrs

31-Dec-16 4.00

1.60

Breakeven Prices

Economic Breakeven Copper Price

$/lb

Project Cashflows

Employee Profit Distribution Sales of Remaining Assets Tax Post-tax Cashflow

$M $M $M $M

303.6 15.5%

-113 37 -370 983

Sensitivity analysis (post-tax)

Table 18.53 illustrates the sensitivity of the project economics to copper price, silver price, gold price, molybdenum price, mining costs, processing costs and capital development costs on a post-tax basis.

Page 399

Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report – Rev0 Table 18.53 Post Tax Project Sensitivity Analysis Post-tax NPV@8% Post-tax IRR $M

% Change

Copper

Silver

Gold

Molybdenum

Mining Costs

Processing Costs

Development Costs

Base = $ -20% -10% 0% 10% 20%

2.00 /lb

Base = $ -20% -10% 0% 10% 20%

12.00 /oz

Base = $ -20% -10% 0% 10% 20%

800.00 /oz

Base = $ -20% -10% 0% 10% 20%

13.00 /lb


2 150 304 456 608

8.0% 11.9% 15.5% 18.7% 21.6%

5.00 4.00 4.00 3.00 3.00

286 295 304 313 322

15.1% 15.3% 15.5% 15.7% 15.9%

4.00 4.00 4.00 4.00 4.00

297 300 304 307 311

15.3% 15.4% 15.5% 15.5% 15.6%

4.00 4.00 4.00 4.00 4.00

272 288 304 320 336

14.8% 15.1% 15.5% 15.8% 16.1%

4.00 4.00 4.00 4.00 4.00

Base = $ variable -20% -10% 0% 10% 20%

351 327 304 280 257

16.4% 15.9% 15.5% 15.0% 14.5%

4.00 4.00 4.00 4.00 4.00

Base = $ variable -20% -10% 0% 10% 20%

372 338 304 269 235

16.8% 16.2% 15.5% 14.7% 14.0%

3.00 4.00 4.00 4.00 4.00

Base = $ -20% -10% 0% 10% 20%

417 360 304 247 189

20.2% 17.6% 15.5% 13.6% 12.0%

3.00 3.00 4.00 4.00 4.00

845.99 M

Page 400


As with the pre-tax analysis, the post-tax NPV and IRR are most sensitive to copper price (Figure 18.47 and Figure 18.48). A 10% change in copper price from the base case of $2/lb results in a change in post-tax NPV at 8% of ±$153 M. Figure 18.47 Post-tax NPV Sensitivity (@8%) 800 700

Copper Spot Price

600

Silver Spot Price

500 Gold Spot Price

NPV$(M)

400 Moly Spot Price

300 200

Mining Cost

100

Processing Costs

0 ‐25%

‐20%

‐15%

‐10%

‐5%

0%

5%

10%

15%

20%

25%

‐100

Development Costs

‐200

Figure 18.48 Post-tax IRR Sensitivity 25% Copper Spot Price 20%

Silver Spot Price Gold Spot Price

IRR (%)

15%

Moly Spot Price 10%

Mining Cost Processing Costs

5% Development Costs 0% ‐25%

‐20%

‐15%

‐10%

‐5%

0%

5%

10%

15%

20%

25%

Page 401


The relative sensitivities for copper price, silver price, gold price, molybdenum price, mining costs, processing costs and capital development costs are shown in Figure 18.49, illustrating the impact of a 10% change in base assumptions on post-tax NPV at 8%. Figure 18.49 Post-tax NPV Sensitivity (@8%) Copper Spot Price Silver Spot Price Gold Spot Price Moly Spot Price Mining Cost Processing Costs Development Costs ‐200

‐150

‐100

‐50

0

50

100

150

200

Change in NPV ($M)

18.17

RISK ASSESSMENT

This part of the report defines the project risk profile from a qualitative perspective. Section 17 - Capital Cost Estimate contains a quantitative assessment of the project risks as part of the development of the project contingency. Generally, study participants have individually undertaken qualitative assessments and analysis of risks and opportunities specific to the facility within their scope of works. Whole of project risks and opportunities, i.e. risks and opportunities that affect the delivery or the project outcome of all facilities have been developed through a risk workshop attended by key study participants. The most significant risks to delivering the project to the budget and achieving the stated viability are summarised as follows; 

Budget over-run – the primary contributors to budget over-run are poor management resulting in scope creep and schedule prolongation. Any over-run will ultimately affect the overall project viability.



Power supply - there is potential that all available capacity at the current supply point of Tintaya substation is secured by other parties. The likelihood is judged to be low, but, should this occur then Norsemont would need to fund the additional capital cost of obtaining a supply from the more distant Cotaruse Substation.

Page 402




Implementation timing - in a regional context the current global financial crisis has resulted in the deferment of project expenditure. Should the activity in the market increase then escalation will increase commensurately with market activity. Opportunity exists to capitalise on the current market by the immediate execution of the project, failure to do so is likely to expose the project to greater competition and hence price escalation.



Power Price - the electrical charge rate paid by Norsemont during operations will be subject to market conditions, hence competition for power may result in the cost of power being higher than the current assumption within the DFS.



Escalation - the estimate is based on pricing lower than what was experienced in late 2008 at the peak of the market. Should commodity prices increase and market competition intensify then it is likely that the project will be exposed to positive escalation.



Political - the upcoming elections will coincide with the execution of the project. This has potential to affect both the project delivery and the project outcome through social conflict and civil unrest.



Pit geotechnical - geotechnical parameters are key criteria in the mine design, in particular the assessment for overall pit slope stability. Additional operational costs would be incurred if the pit slope stability adopted proves to be poorer than anticipated.



Norsemont as Developer/Operator - the implementation strategy is based on Norsemont undertaking the overall project management role for the execution phase of the project, for constructing the bulk earthworks using its construction fleet, and the owner-operator role for the operations phase of the project. Significant project and construction management capabilities will be required by the Owner for this and particularly to ensure the timely completion of the bulk earthworks. This approach risks causing delay to the project schedule and prolongation claims from other contractors with liability falling on Norsemont. Cost increases arising from the need to employ additional staff to provide project management are a further risk.




Additional geotechnical and related risks and opportunities identified by Knight Piésold are as follows: Construction approach Earthworks construction will be carried out using Owner operated equipment and personnel, rather than by a specialist contractor. The risks associated with this are: 

Insufficient or inappropriate planning and scheduling that could lead to delays in the construction and thus mine start up



Provision of insufficient types or numbers of construction equipment and/or personnel that could also lead to delays in the construction and thus mine start up



Construction cost overruns due to either of the above points.

Page 403


Mine pit slopes The most significant geotechnical risks for the pit are the potential for geological faults occurring within Sectors VI and VII. Present geologic data cannot rule out the projection into these sectors of certain faults that have been mapped elsewhere. The pit slopes in these sectors, as in all the sectors, have been analyzed assuming that free draining conditions will be maintained during pit operations. If these faults are confirmed and/or a non-free draining behaviour governs in the rock mass in these sectors, the pit slopes will have to be flatter than currently recommended. Additional field work in these areas is required to mitigate this risk. Catch bench geometries and performance estimates have been based on geotechnical interpretation and rock fracturing data developed from drill core and limited surface exposures, considering the expected effectiveness of controlled blasting techniques. These assumptions and observations have not been confirmed through full-scale exposures. There is some potential that the bench face angles and therefore the catch bench widths will have statistical distributions different from those calculated. Bog removal Existing information on the bog material characteristics and quantities within the TMF basin is limited. As such, there is a risk that the material may require different and more costly removal methods than those anticipated, and/or the quantity is significantly greater than currently assumed. The likelihood that the removal method may prove more costly is higher than the likelihood that the quantity may be larger. However, realisation of either would be expected to result in a potentially significant additional cost. TMF water quality Initial geochemical characterisation of the tailings shows that combined deposition of the rougher scavenger and cleaner scavenger tailings in a single stream into the TMF, as currently planned, will not result in the development of acidic conditions if the tailings deposition is managed properly. It is expected that the tailings surface at any time will not be exposed for periods of time long enough, before the overlying tailings layer is deposited, to enable acidic conditions to develop. Under this proviso, excess water from the TMF will be discharged to the Chilloroya River during annual wet periods. If the water quality in the TMF pond is degraded below release quality by improper tailings management or changes in the tailing characteristics this water and the collected underdrain flows may have to be treated prior to release. Similarly, in areas downstream of unlined portions of the TMF, seepage intercept and pump-back wells may have to be installed and operated. Water treatment capabilities Water from the PAG WRF containment pond and the in-pit sumps are expected to be acidic and will be neutralised at the plant site. It has been assumed that the treatment facilities at the plant will be sufficient to treat this water. If the facilities are insufficient for this, the resulting acidic water may have to be directed to the TMF. In this case, the water from the TMF would require treatment. Fresh water supply Fresh water for plant processes is expected to come from dewatering and water supply wells in the pits. Risks are associated with availability of groundwater. If groundwater extraction rates, including dewatering flows, are lower than currently predicted, additional dewatering wells and/or deeper wells would be required. Although the numerical modelling indicates that flows of 130 L/sec can be

Page 404


maintained throughout the life of the mine, uncertainties in the model exist due to geological complexities, and the greater length of the model predictions relative to the available data set for calibration. Further uncertainties exist with respect to water supply during an extended drought period. If the water does not meet required standards, a raw water pond, with a capacity of approximately 2 million m3 may be required. Poor groundwater quality during operations Quality of groundwater presents a risk; if the in-pit dewatering wells become impacted with poor quality recharge from the exposed pit walls, some of the water may require collection and treatment, or alternative sources of water supply may be required. Poor post-closure pit water quality There is a potential for poor water quality seepage from the open pit during pit flooding and post pit flooding. Ground water seepage surrounding the open pit will be monitored throughout the mine life and collected and treated if necessary. Reduction of baseflow to springs and lakes due to pit dewatering Two potential adverse impacts could occur to the limestone aquifers and related springs – (1) baseflow reduction to all springs located within the hydraulic influence of the pit dewatering system; and (2) hydraulic interconnectivity between the open pit and dewatering wells and the community water reservoirs situated north of the mine via the Yanak Fault. Baseflow to the lakes could be reduced if the phreatic levels drop and disconnects form the main springs within these basins. Maintaining baseflow to the springs and lakes during the dry season months from wells would be required to mitigate impacts. Tailings management facility Depending on the geochemical characteristics of the tailings and the hydrogeology at the TMF site as determined at the time of final design, it may be possible to reduce the extent of the geomembrane liner in the TMF. Alternatively, it may be possible to replace the geomembrane liner with an equivalent or near equivalent seepage mitigation measure such as a perimeter cut-off trench and grout curtain. The latter may or may not result in a cost reduction however, and would require a dedicated trade off study. It may also lead to construction complications since a trench and grout curtain may not be able to be constructed in parallel with the embankment.

Page 405


19.

INTERPRETATION AND CONCLUSIONS

A DFS has been completed for the Constancia Project, covering all disciplines, i.e. resource modelling, mining, metallurgical testwork and process design, plant and infrastructure design and engineering, project implementation, environmental studies, and capital and operating cost estimation. This work is reported in Sections 16 to 18. The accuracy of the cost estimates is considered to be ±15%, reflecting the level of detail of engineering design. Geotechnical site investigations have been completed to a feasibility level for the Constancia pit, TMF, plant site, and associated structures. Investigations at the site of the PAG WRF, including the containment pond and the potential borrow areas, were limited due to land access restrictions. The information available at this time is considered sufficient to support the feasibility, but additional field work should be performed at these locations to support the next stage of design. The stability analysis developed for the inter-ramp slopes in the pit provided factors of safety over 1.2 under static conditions, and over 1.0 for the case of seismic loads associated with a 100 year return period. These factors of safety are considered acceptable for this level of design and are consistent with Peruvian regulations. In consideration of potential bench scale failures, 77% of the bench face slopes in the pit are considered to have less than 27% probability of failure. The remaining 25% of the bench face slopes are considered to have a 55% failure probability. Overall, it is calculated that catch benches will meet the minimum width criterion with a reliability of 65% or better. Bench scale failures will be mitigated by scaling and maintenance of catch benches. Water for the mill will be supplied from the following sources: groundwater from dewatering wells located behind the pit walls, in pit surface water collected in the pit sumps, drainage from the PAG WRF that will be collected in the containment pond just below the PAG WRF, and reclaimed water from the surface water pond in the TMF. Water from the PAG WRF and in-pit sumps is expected to be acidic and will be neutralised with lime at the process plant before use in the process. Raw water for sustaining mill operations will be obtained, in part, from pit dewatering wells installed behind the pit perimeter. Based on the current hydrogeologic model, a total of 18 dewatering wells will be installed in and around the pit. Water from these wells will be pumped to a collection box immediately upstream of the plant and then conveyed by gravity to a dedicated tank in the plant. Some additional work is recommended to improve confidence in certain areas, but any changes are not expected to change the outcome of the DFS significantly. A post-tax cashflow model has been prepared by Norsemont, which indicates a NPV of $303.6 M and an IRR of 15.5%. However, these values do not take account of financing costs required to develop the Project, which will now be investigated by Norsemont. Project economics are most sensitive to the long-term copper price: a constant price of $2.00/lb has been assumed for the Base Case cashflow models. Other important economic variables are the total capital cost, treatment charges, the cost of diesel and electricity.

Page 406


Risk analysis identified the following issues: 

Budget over-run – the primary contributors to budget over-run are poor management resulting in scope creep and schedule prolongation. Typical of any project, budget over-run will ultimately affect the overall project viability.



Power supply – there is potential that all available capacity at the current supply point of Tintaya substation is secured by other parties. This situation is common with projects, as such Norsemont should commence negotiations to secure this supply, however, should this supply not be secured then Norsemont’s alternative is to fund the additional capital cost of obtaining a supply from the more distant Cotaruse Substation.



Implementation timing - in a regional context the current global financial crisis has resulted in the deferment of project expenditure. Should the activity in the construction market increase then escalation will increase commensurately with market activity. Opportunity exists to capitalise on the current market by the immediate execution of the project, the longer the delay the greater the potential impact of price escalation.



Power Price - the actual electrical charge rate paid by Norsemont during operations will be subject to market conditions, hence competition for power may result in the cost of power being higher that the current assumption within the DFS. This risk is considered as part of the sensitivity modelling.



Pit geotechnical - geotechnical parameters are key criteria in the mine design, in particular the assessment for overall pit slope stability. Additional operational costs would be incurred if the pit slope stability estimates prove to be too optimistic.



Norsemont as Developer/Operator - the implementation strategy relies upon Norsemont undertaking the overall project management role for the execution phase of the project and the owner operator role for the operations phase of the project. This strategy is the most cost effective and provides the effective management of construction risk, however, offsetting this is the risk in being able to recruit a highly capable delivery and operational team.




Opportunities exist to improve the Project in the following areas: 

The detailed planning for the bulk earthworks and purchase of the Owners civil construction fleet has potential to identify capital and operational costs savings and mitigate time and cost risks in the early construction phase of the project.



The investigation of the availability of second hand accommodation camps has the potential to save costs. Currently the capital cost is based on the provision of a new camp.



The investigation of the availability of new long lead equipment, i.e. cancelled orders etc. may provide schedule opportunities for the project. Currently these long lead item are critical path items, hence a reduction in time of the delivery of key equipment will provide schedule assurance and possibly an overall schedule reduction.

Page 407




Negotiations with the various organisations that generate and transmit power may allow more confidence on power pricing and availability. However, it is understood that firm commitments by government organisations can only be provided closer to the time of supply.



Detailed design of the improvement works for the access road upgrade and stakeholder consultation will increase the level of confidence around the scope of works. In addition, the acceleration of the delivery of the access road such that it is available for the construction of the mine will reduce transportation and travel risks and reduce travel time and delays.

Page 408


20.

RECOMMENDATIONS FOR FURTHER WORK

20.1

RESOURCES

No further database or resource modelling work is required until a decision is taken to develop the Project, at which time additional drilling and resource modelling will be required to support more detailed mine planning.. 20.2

MINING

The DFS pit designs include a quadrant where the probability of bench scale failure is 44%. The incremental costs associated with flattening the pit slope in this area should be assessed relative to the corresponding reduced probability of bench scale failures. The fuel (diesel) and explosive unit prices used in the study have been based on budget pricing for a recent Peruvian project, factored to address a potential freight differential. This approach was adopted when enquiries to Peruvian vendors were not answered in a timely way. These assumed rates should be confirmed by following up on the budget pricing submissions already requested from local suppliers. Heavy equipment budget pricing should be followed up as the starting point for future equipment tendering activities. Additionally, some vendors are prepared to estimate fleet requirements at their own cost but require more detailed mine design information to perform these analyses. The study has assumed that reticulated mains power will be available to allow electric shovels to operate from five months prior to plant start-up. Definition of an alternative diesel shovel based approach may be desirable in the event the provision of power is delayed. Analysis of available vendor and other information indicated that the operating cost savings associated with electric drills was not sufficient to justify the incremental capital investment relative to diesel units. This interim conclusion was based on incomplete data and should be supported by a more in depth analysis prior to, or as part of, the mine equipment tendering process The owner civil fleet to address tails dam construction and related mine work has been prepared somewhat independently of the mine production fleet assessment. These two fleets should be evaluated to determine if there are any synergies that might improve project value. 20.3

GEOTECHNICAL AND HYDROGEOLOGICAL STUDIES

Selected geotechnical investigations are recommended to provide support follow-up stages of engineering. These include: 

Additional field geotechnical investigations within the southern portion of the PAG WRF (limited investigations have been possible to date due to access restrictions).



Investigations to evaluate the need for removing the extremely weathered diorite to significant depths beneath the TMF embankment.



Additional geotechnical characterisation of the tailings and waste rock materials.

Page 409




Additional investigations in potential borrow areas to further characterise potential construction materials and refine the quantity estimates.



Within the mine pit, two additional drillholes in Sector VI and one in Sector VII are necessary to investigate the potential that locally observed geological faults may extend through this area.



Risk analyses should be carried out to investigate the probability and consequences of a rock mass slide occurring over the access ramp, and of removing the access to the pit in the Sector VI area.

With regards to hydrogeological studies, the following actions are recommended prior to project development: 

Further investigations within the TMF and PAG WRF to confirm the criteria used in the design of these facilities. Within the PAG WRF, the objective will be to confirm the containment capability of the existing regime.



A further, detailed investigation to assess the probability and potential influence of pit dewatering on the lakes and karstic area situated north of the pit is in progress. The investigation includes additional drilling, hydraulic testing and piezometer installations along the Yanak fault and within the lakes catchment area to assess the existing ground water system and the degree of hydraulic continuity and connection between the Yanak fault and lake basins. Results of this work will determine whether any further action is required.



It is recommended that a groundwater exploration program is undertaken to assess the availability of additional sources of process water as an alternative supply during drought conditions. Targets include: the Yanak fault south of the pit area and structures paralleling the Chilloroya valley below the glacial valley infill material.



Installation of additional ground water monitoring and sampling wells in the following locations: (1) within Quebrada Telaracaca to monitor potential seepages from the PAG WRF via fault zones; (2) upgradient and downgradient of the PAG WRF within Quebrada Huayllachane; and (3) downgradient of the proposed plant site.

20.4

METALLURGICAL TESTWORK

Testwork is warranted to evaluate the production of zinc concentrate, as a number of periods exist when this may be possible. These include: 

Year 0 to Year 4: during this period zinc is activated by virtue of the treatment of high zinc ore with supergene ore, leading to low requirements of zinc activation reagents. This period also represents the peak of copper production, with any zinc production requiring the additional capital for grinding, flotation, concentrate thickening and filtration equipment.



Year 9 to Year 11 and Year 18 to Year 20: during these periods, excess capacity is available in the flotation and filtration areas, and circuit reconfiguration can be conducted with small capital outlay to produce a zinc concentrate. Sphalerite will, however, require activation by CuSO4 with an increase in operating costs.

Page 410


Silver recovery calculations were hampered by ICP analysis for the locked cycle tailing streams at times being below the detection limit of 0.5 ppm. Further evaluation of silver recovery and deportments is required. Reduction in the talc/amphibole content of the molybdenum concentrate will produce increased revenue via reduced roasting charges. Further evaluation in an extended test work program is required into the extent of naturally floatable minerals in the Constancia deposit, their association with molybdenite and methods for their removal. Potential solutions include: 

Replacement or partial replacement of light fuel oil with other collectors such as A3302 in molybdenum flotation



Destruction of the fuel oil promoter after the early stages of molybdenum flotation and reverse flotation of naturally floating gangue minerals.

20.5

INFRASTRUCTURE

Negotiations with the various organisations that generate and transmit power may allow more confidence on power pricing and availability. However, it is understood that firm commitments by government organisations can only be provided closer to the time of supply. Detailed design of the improvement works for the access road upgrade and stakeholder consultation will increase the level of confidence around the scope of works. In addition, the acceleration of the delivery of the access road such that it is available for the construction of the mine will reduce transportation and travel risks and reduce travel time and delays. 20.6

ESIA AND PERMITTING

The ESIA is scheduled for completion by December 2009. Approval is expected by December 2010. The development of the permitting plan should begin in 2010. 20.7

PROJECT IMPLEMENTATION

The detailed planning for the bulk earthworks and purchase of the Owners civil construction fleet has potential to identify capital and operational costs savings and mitigate time and cost risks in the early construction phase of the project. The investigation of the availability of second hand accommodation camps has the potential to save costs. Currently the capital cost is based on the provision of a new camp. The investigation of the availability of new long lead equipment, i.e. cancelled orders etc., may provide schedule opportunities for the project. Currently these long lead item are critical path items, hence a reduction in time of the delivery of key equipment will provide schedule assurance and possibly an overall schedule reduction.

Page 411


21.

REFERENCES

Bieniawski, Z.T., 1989. Engineering Rock Mass Classifications, John Wiley and Sons, New York. Goodman, R.E., 1989. Introduction to Rock Mechanics, 2nd Edition, John Wiley and Sons, New York. Hoek, E. and Bray, J.W., 1981. Rock Slope Engineering, 3rd Edition, London. Hoek, E., Carranza-Torres, C. and Corkum, B., 2002. Hoek-Brown Failure Criterion – 2002 Edition, in NARMS-TAC 2002: Mining and Tunneling Innovation and Opportunity, Vol. 1 pp 267 – 273. R Hammah et al, Eds. University of Toronto Press, Toronto, 2002. Hoek, E., Kaiser, P.K. and Bawden, W.F., 1995. Support of Underground Excavations in Hard Rock, A.A. Balkema. Knight Piésold Ltd., October 2007. Proyecto Mina Justa – Geotechnical Investigation (Ref. No. LI20100209/08). Lytle, Murray, 2004. Technical Report of the Constancia Project, Department of Cusco, Peru. Technical Report prepared in compliance with National Instrument 43-101 and Form 43-101F. McRae, J.A., 2006. November 8, 2006.

Technical Report of the Constancia Copper, Molybdenum Porphyry Property,

Ryan, T.M., and Pryor, P.R., 2000. Designing Catch Benches and Inter-ramp Slopes: Chapter 3 in W. A. Hustrulid, M. K. McCarter, and D. J. A. Van Zyl, (eds.), Slope Stability in Surface Mining: Society for Mining, Metallurgy and Exploration, Littleton, CO, USA. Sjoberg, J., 1996. Large Scale Slope Stability in Open Pit - A Review. Technical Report, Division of Rock Mechanics, Lulea University of Technology - S-97187 Lulea Sweden. Snowden Mining Industry consultants Inc., May 2007. Constancia Project Technical Report, May 9, 2007. Wyllie Duncan.C., and Mah C.W., 2004. Rock Slope Engineering – Civil and Mining, 4th Edition, London.

Page 412


22.

DATE AND SIGNATURE PAGE

The undersigned Qualified Persons prepared this Technical Report, titled Constancia Copper Project Definitive Feasibility Study NI 43-101 Technical Report with an effective date of 28 September 2009. The format and content of the report are intended to conform to Form 43-101F1 of National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators. Title: Location:

Constancia Copper Project Definitive Feasibility Study, NI43-101 Technical Report Province of Cusco, Peru

Effective Dates: Effective Date of Technical Report:: Effective Date of Mineral Reserves: Effective Date of Mineral Resources:

28 September 2009 28 September 2009 28 September 2009

Qualified Persons: 

David D. (Dan) Greig, B.Sc, M.A.I.G. (#1684-1722), employed by GRD Minproc Limited as Principal Geologist, was responsible for overall preparation of the report.

Signed:

/s/ David D. Greig

Date: 28 September 2009

Dan Greig 

Lynn Widenbar, B.Sc. (Geology), M.Sc, MAusIMM, employed by GRD Minproc Limited as Consulting Geologist, was responsible for the preparation of Section 14 (Data Verification) and Section 17 (Mineral Resources), and for review of Section 12 and Section 13 of this report.

Signed:

/s/ Lynn Widenbar


Lynn Widenbar



Ross Oliver, B.Eng (Mining), AusIMM (#105137), employed by GRD Minproc Limited as Manager Mining & Geology, was responsible for the preparation of Section 17.14 (Mineral reserve) and Section 18.1 (Mining).

Signed:

/s/ Ross Oliver


Ross Oliver

Page 413




Greg Harbort, B.Eng (Met), University of Qld, 1985; Ph.D. University of Qld, 2005; MAusIMM Principal Process engineer for GRD Minproc Limited, was responsible for Mineral Processing and Metallurgical Testwork (Section 16), and for Plant Operating Costs and site infrastructure operating costs (Section18.14).

Signed:

/s/ Greg Harbort


Greg Harbort



Craig Cuttriss, B.Sc, M.Eng, R.P.E.Q., MAusIMM, Principal Process Engineer and Study Manager for GRD Minproc Limited,was responsible for GRD Minproc’s engineering aspects of the Project as contained in Section 18.4, Section 18.9, Section 18.11, Section 18.13 and Section 18.13.

Signed:

/s/ Craig Cuttriss


Craig Cuttriss



Thomas Kerr, B.Eng., (Civil), P.Eng., (British Columbia and Ontario), and P.E. (California and Alaska) in good standing, President of Knight Piésold and Co. (USA), was responsible for those parts of Section 18 relating to the Tailings Storage Facility, site geotechnical investigations and water and waste management.

Signed:

/s/ Thomas Kerr


Thomas Kerr



Robert Cummings M.Sc. Geol. Eng., Registered Professional Engineer (Arizona), Geotechnical Consultant and Principal of Saguaro Geoservices, was responsible for pit slope stability analysis and design parameters in Section 18.1.3 and Section 18.2.2.

Signed:

/s/ Robert Cummings

Date: : 28 September 2009

Robert Cummings

Page 414


23.

ADDITONAL REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION AND DEVELOPMENT PROPERTIES

As the Constancia Project has not yet reached the development or production stage, there are no additional requirements to report.

Page 415


24.

ILLUSTRATIONS

Illustrations are included as appropriate throughout the text, and are listed in the Index (List of Figures).

Page 416


25.

ANNEXURES

Two annexures are enclosed, namely 

A preliminary list of permits required for construction and operations



An English translation of the tax opinion provided by Picon & Asociados.

Page 417


Appendix 1 Permits

List of Required Permits and Licences – Construction Phase Regulatory Authority

Permit Name

Regulatory Agency

CONSTRUCTION ACTIVITIES MINING Acquisition of Mining Concessions

INGEMMET

DCM

Approval of Contract for taxation stability

MINEM

DPDM/DGM

Environmental Impact Assessment

MINEM

DGAAM

Emergency Mining Plan

MINEM

DGM

Registry and control of hazardous substances

MINEM

DGM

Contingency Plan

MINEM/SINADECI

DGM/ORDC

Establishment of a mining and security hygiene committee

MINEM

DGM

MINEM

DGAAM

MINEM

DGM

MINSA

DIGESA

Relocation

and/or

cooperation

agreements

with

local

communities within the area of the project. Approval of Feasibility Study Environmental Impact Assessment for Ministry of Health Permits Concession for Plant Operations

MINEM

DGM

Powder Box License (up to 5 years)

MININTER

DICSCAMEC

MINAG/ANA

ALA

Authorization of Studies and Execution of Work

MINAG/ANA

ALA

Permit for Surface Water Use for Potable Water Supply

MINAG/ANA

ALA

Permit for Groundwater Use for Potable Water Supply

MINAG/ANA

ALA

Permit for Surface Water Use for Industrial Water Supply

MINAG/ANA

ALA

Authorization to Conduct Studies for the Use of Groundwater for Consumption

Permit for Groundwater Use for Industrial Water Supply

MINAG/ANA

ALA

Authorization for the Construction of Hydrological Works

MINAG/ANA

ALA

MINEM

DGAAM

MINEM

DGAAM

MINEM

DGM

MTC

PVN/PVD

MTC

DGASA

Construction Permit

MINEM

DGM

Construction Site Inspection

MINEM

DGM

MINSA

DIGESA

MINSA

DIGESA

MINSA

DIGESA

MINSA

DIGESA

MINSA

DIRESA

Permit to Operate a Medical Centre

MINSA

DISA

Permit to Install an X-ray Machine for Medical Diagnoses

MINEM

IPEN

Permit to Operate an X-ray Machine for Medical Diagnoses

MINEM

IPEN

Declaration

and

identification

of

mining

environmental

liabilities. Environmental Liabilities Closure Plan ROADS Application for Approval of Investment in Public Infrastructure (This procedure includes the approval of the Feasibility Study and the SNIP) Contracts to supervision of the roads works Environmental Impact Assessment of the road, which will be constructed by Provias MINING CAMP

Sanitary Authorization for septic tanks and final disposal on the ground. Sanitary Authorization for a Sanitary Industrial Wastewater Treatment and Disposal System Sanitary Authorization for a Sanitary Domestic Wastewater Treatment and Disposal System Sanitary Authorization for a Potable Water Treatment Plant Sanitary Authorization for the Construction and Operation of a Kitchen and Warehouse

Regulatory Authority

Permit Name Registration of Installations and Sources of Radiation (for

Regulatory Agency

MINEM

IPEN

Authorization to Transport Radioactive/Nuclear Materials

MINEM

IPEN

Approval of Fuel Provision Plan

PCM

OSINERGMIN

Approval for Installation of Fuel Tanks

PCM

OSINERGMIN

Approval for Use of Fuel Tanks

PCM

OSINERGMIN

Registration of Use and Operation of a Fuel Station

MINEM

DREM

MTC

DDST

Permit to Import Telecommunications Equipment

MTC

DDST

Endorsement of Telecommunications Equipment and Apparati

MTC

DDST

Permit to Use Restricted Chemicals and Products

MITINCI

DPIQPF

MITINCI

DPIQPF

Annual Declaration of Waste Management

MINEM

DGM

Waste Management Plan

MINEM

DGAAM

Contingency Plan for Waste Management

MINEM

DGM

Solid Waste Manifestos

MINEM

DGAAM

Contracts for the transportation and disposal of solid wastes

MINEM

DGAAM

Deforestation Authorization.

MINAG

DFFF

nuclear densometers)

Authorization to Install and Operate Private Communications Systems

Monthly Report on the Movement of Restricted Chemicals and Products

Operation Permit for air transportation

MTC

DGAC

Direct Consumers Registry of Hydrocarbons

MINEM

DGH

A non-metallic mining concession

INGEMMET

INGEMMET

Approval of Contract for taxation stability

MINEM

DPDM/DGM

Environmental Impact Assessment

MINEM

DGAAM

Emergency Mining Plan

MINEM

DGM

QUARRY OPERATIONS


MINEM

DGM

Contingency Plan

MINEM/SINADECI

DGM/ORDC

MINEM

DGM

MINEM

DGAAM

Establishment of a mining and security hygiene committee Relocation

and/or

cooperation

agreements

with

local

communities within the area of the project. Approval of Feasibility Study

MINEM

DGM


MININTER

DICSCAMEC

MINEM

DGAAM

MINEM

DGAAM

MINEM

DGE

Declaration

and

identification

of

mining

environmental

liabilities. Environmental Liabilities Closure Plan ELECTRICITY Concession for Electrical Installations Establishment of Electrical Right-of-Way

MINEM

DGE

Approval of Electrical System Testing Plan

MINEM

DGE

MINEM

DGE

MINEM

DGE

Concession or authorization to perform electric generation activities for self consumption Approval of Completion Report

List of Required Permits and Licences – Operations Phase Permit Name

Authority/Autoridad

Agency/Agencia


MINEM

DGM

Annual Program of security and mining hygiene

MINEM

DGM

MINEM

DGAAM

OPERATIONS PHASE MINING

Relocation and/or cooperation agreements with local communities within the area of the project. Environmental Impact Assessment for Ministry of Health Permits

MINSA

DIGESA

Certificate of the Inexistence of Archaeological Remains (CIRA)

MINED

INC/DGPA

MINEM

DGM

MINEM

DGM

Authorization for the Start of Mineral Recovery, Metal

Mining

Concessions Certificate of Mining Operations (COM)/ Metal Operations Powder Box License (up to 5 years)

MININTER

DICSCAMEC

Explosives Transport License (1 year)

MININTER

DICSCAMEC

MINAG/ANA

ALA

Authorization of Studies and Execution of Work

MINAG/ANA

ALA

Permit for Surface Water Use for Potable Water Supply

MINAG/ANA

ALA

Permit for Groundwater Use for Potable Water Supply

MINAG/ANA

ALA

Permit for Surface Water Use for Industrial Water Supply

MINAG/ANA

ALA

Permit for Groundwater Use for Industrial Water Supply

MINAG/ANA

ALA

Authorization for the Construction of Hydrological Works

MINAG/ANA

ALA

Mining Closure Plan

MINEM

DGAAM

Construction Permit

MINEM

DGM

Construction Site Inspection

MINEM

DGM

MINSA

DIGESA

MINSA

DIGESA

MINSA

DIGESA

MINSA

DIGESA

MINSA

DIRESA

Permit to Operate a Medical Centre

MINSA

DISA

Permit to Install an X-ray Machine for Medical Diagnoses

MINEM

IPEN

Permit to Operate an X-ray Machine for Medical Diagnoses

MINEM

IPEN

MINEM

IPEN

Authorization to Transport Radioactive/Nuclear Materials

MINEM

IPEN

Approval of Fuel Provision Plan

PCM

OSINERGMIN

Authorization to Conduct Studies for the Use of Groundwater for Consumption

MINING CAMP

Sanitary Authorization for septic tanks and final disposal on the ground. Sanitary Authorization for a Sanitary Industrial Wastewater Treatment and Disposal System Sanitary Authorization for a Sanitary Domestic Wastewater Treatment and Disposal System Sanitary Authorization for a Potable Water Treatment Plant Sanitary Authorization for the Construction and Operation of a Kitchen and Warehouse

Registration of Installations and Sources of Radiation (for nuclear densometers)

Approval for Installation of Fuel Tanks

PCM

OSINERGMIN

Approval for Use of Fuel Tanks

PCM

OSINERGMIN

Registration of Use and Operation of a Fuel Station

MINEM

DREM

MTC

DDST

Permit to Import Telecommunications Equipment

MTC

DDST

Endorsement of Telecommunications Equipment and Apparati

MTC

DDST

Permit to Use Restricted Chemicals and Products

MITINCI

DPIQPF

Authorization to Install and Operate Private Communications Systems

Permit Name

Authority/Autoridad

Agency/Agencia

MITINCI

DPIQPF

Annual Declaration of Waste Management

MINEM

DGM

Waste Management Plan

MINEM

DGAAM

Contingency Plan for Waste Management

MINEM

DGM

Monthly Report on the Movement of Restricted Chemicals and Products

Solid Waste Manifestos

MINEM

DGAAM

Contracts for the transportation and disposal of solid wastes

MINEM

DGAAM

Deforestation Authorization.

MINAG

DFFF

Operation Permit for air transportation

MTC

DGAC

MINEM

DGH

Direct Consumers Registry of Hydrocarbons

OPERATION OF QUARRIES Authorization for the start of Exploitation of Quarries

MINEM

Quarries Closure Plan

MINEM

DGAAM


MINEM

DGM

Annual Program of security and mining hygiene

MINEM

DGM

Certificate of Mining Operations (COM)/Non Metal Operations

MINEM

DGM

MINEM

DGAAM

MINED

INC/DGPA

Global Explosives License

MININTER

DICSCAMEC


MININTER

DICSCAMEC

Explosives Transport License (1 year)

MININTER

DICSCAMEC

Relocation and/or cooperation agreements with local communities within the area of the project. Certificate of the Inexistence of Archaeological Remains (CIRA)

DGM

List of Definitions DEFINITIONS CIRA

Certificate of the Existence of Archaeological Remains

DCM

Directorate of Mining Concessions

DDST

Directorate of Telecommunications Service Department

DFFF

Directory of Forestry and Wildlife

DGAC

General Direction of Civil Aeronautics

DGAAA

General Direction of Agricultural Environmental Affairs

DGAAM

General Directorate of Mining Environmental Affairs

DGAAE

General Directorate of Environmental Affairs for Energy

DGASA

General Directorate of Social and Environmental Affairs

DGE

General Directorate of Energy

DGH

General Directorate of Hydrocarbons.

DGM

General Directorate of Mining

DGPA DICSCAMEC

General Directorate of Archaeological Patrimony General Directorate for Control of Security Services, Firearms, Munitions, and Explosives for Civil Use

DIGESA

Dirección General de Salud Ambiental

DIRESA

General Directorate of Environmental Health

DISA

Health District

DPDM

Directorate for the Promotion and Development of Mining

DPIQPF

Directorate of Industrial Procedures and Controlled Chemicals and Products

DREM

Regional Directorate of Energy and Mines

DTM

Technical Direction of Mining

INC

National Institute of Culture

INGEMMET

Geological Institute of Mining and Metallurgy

IPEN

Peruvian Nuclear Institute

MINAG

Ministry of Agriculture

MINED

Ministry of Education

MINEM

Ministry of Energy and Mines

MININTER

Ministry of the Interior

MINSA MITINCI

Ministry of Health Ministry

of

Industry,

Tourism,

Integration,

and

Commercial Negotiation

MTC

Ministry of Transportation and Communications

ORDC

Regional Office of Civil Protection

OSINERGMIN

Supervisory Agency for Energy and Mining Investment

PCM

Presidency of the Council of Ministers

PVD

Provías Descentralizado

PVN

Provías Nacional

RRPPMM

Public Registry

RRPPMM

Public Mining Registry

International


Appendix 2 Picon & Asociados Tax Opinion

REPORT S/N-2009-PICON/NORSEMONT

I.

TO

:

NORSEMONT MINING

FROM

:

PICÓN & ASOCIADOS S.A.C.

DATE

:

September 4, 2009

SUBJECT

:

Tax analysis of cash flow

TOPIC We have been requested to issue a tax opinion about cash flow prepared by the company regarding future investments.

II.

LEGISLATION        

III.

Supreme Decree Nº 014-92-EM,1 Revised Text of Mining General Law. Legislative Decree Nº 299,2 Leasing Law. Legislative Decree Nº 973,3 VAT Early Recovery Regime. Supreme Decree N° 084-2007-EF,4 VAT Early Recovery Regulations. Law Nº 27343,5 Juridical Stability Agreements Regime. Legislative Decrees N° 677 and 892, Workers participation in corporate profits. Supreme Decree Nº 009-98-TR – Regulations for workers participation in corporate profits. Law Nº 29230 and Supreme Decree Nº 147-2008-EF, Incentives for private sector to participate in public investments.

OPINION In general, cash flow reviewed complies with Peruvian tax laws, especially considering Income Tax provisions. In former meetings with members of your staff, we made some suggestions in order to improve the financial statement, which has been incorporated in cash flow subject of this report.

3.1

Preoperating costs (US $ 73 MM) 6 Regarding exploration investments, Article 74 of Mining General Law (hereinafter also MGL) indicates:

1 2

3 4 5 6

Published on June 2, 1982. Published on July 20, 1984. According to Transitory Provision 6 of Law Nº 27804, Leasing benefits might be included in Juridical Stability Agreements. Published on March 10, 2007. Published on June 29, 2007. Published on September 6, 2000 and clarified by Article 3 of Law Nº 27391. Line 533 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909.

Calle Manuel Gonzales Olaechea 271, San Isidro T. 7100224 F. 7100201 Pág. 1/10

Acquisition value of mining concessions will be amortized since year in which company must reach minimum production, for a time based on mine useful life, whose estimation relies on probed and probable reserves and mandatory minimum production. Amortization time must be informed to Tax Administration, through Annual Income Tax Return corresponding to tax period in which amortization starts. Acquisition value of mining concessions includes price paid, exploration investments incurred up to date in which mandatory minimum production must be complied, unless company chooses to deduct exploration expenses in year when they are executed (...) Consequently, Peruvian tax regulations provide two choices to treat exploration expenses in mining: 7 a. b.

100% deduction in one year. Inclusion as assets and amortization during concession useful life, beginning from tax period in which company starts minimum production.

Even further, company may adopt a tax treatment (for instance, 100% deduction) different to the accounting choice (for example, inclusion as asset and amortization). By other side, Article 75 of Mining General Law states that: Exploration expenses incurred after concession reaches mandatory minimum production might be deducted that year or amortized based on a mine useful life, whose estimation relies on probed and probable reserves and mandatory minimum production. Preparation and development expenses that allow mine exploitation for more than one year might be deducted in one year or amortized in three years. In each case, taxpayer´s election must be informed to Tax Administration through Annual Income Tax Return corresponding to tax period in which expenses were incurred. If reserves were consumed or concession was extinguished before exploration, development and preparation expenses were totally amortized, taxpayer may choose to immediately amortize any remaining amount or continue amortizing during the original period.

7

Peruvian Mining General Law, Art. 74.


As we can see in previous paragraph, “preparation and development expenses” that allow mine exploitation for more than one year might be completely deducted in one period or amortized in three years. 8 In summary, cash flow is correct in this point because tax losses might be carried forward according to Income Tax regulations, even when financial accounting may adopt a different treatment. 3.2

Mining (US $ 120 MM) 9 Fixed asset definition In order to better understand tax depreciation benefit granted by Mining General Law, we will proceed to define what are fixed assets. Fixed assets and intangibles are not defined in the Income Tax Law or in its Regulations, but we can find some useful concepts in International Accounting Standards (IAS), Generally Accepted Accounting Principles (U.S. GAAP) and accounting doctrine. Thus, Paragraph 6 of IAS 16 – Property, Plant and Equipment, characterizes these assets as tangible property that: (a) (b)

are used in production or supply of goods and services, are leased or used for administrative purposes; and, should be used during more than one year.

Gross value of fixed assets is reduced by effect of depreciation, which represents loss or reduction in that value because of time or obsolescence. Accounting terminology associates depreciation to tangible property and amortization to intangible property. Peruvian Income Tax Law and its Regulations have adopted same standards. Depreciation under Mining General Law Article 84 of Mining General Law indicates that stability agreements will guarantee, among other benefits, the right to use 20% annual depreciation for the majority of fixed assets, except buildings/edifications whose depreciation rate is up to 5%. According to Article 17 of MGL Regulations,10 when Mining General Authority approves project feasibility must also pronounce about maximum depreciation percentage allowed to taxpayer.

8 9 10

Peruvian Mining General Law, Art. 75. Line 537 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909. Supreme Decree N° 024-93-EM.


We have been informed that all assets registered under item “Mining” are machinery/equipment and consequently, 20% depreciation is allowed under MGL provisions. Leasing Contract Regulations According to Article 18 of Legislative Decree 299 – Leasing Law, this contract has following features in Peru: a.

Under IAS 17, assets included in the contract are registered and depreciated by the lessee.

b.

Useful life for those assets is duration of leasing contract, which can not be less than two (2) years for movable property and five (5) years for buildings/edifications. That implies possible depreciation rates of 50% and 20%, respectively.

c.

Call option (purchase decision) must be exercised at the end of the contract.

We have been informed that company would use leasing contracts to acquire machinery/equipment and consequently, cash flow should be modified as follows: ASSETS

PERCENTAGE

DEPRECIATION

Mining

100%

50%

Use of leasing contracts implies assumption of banking costs such as interests, insurance and other fees; notwithstanding, we understand that those amounts should not affect cash flow because they are not quite different to other financing source. 3.3

Process plant & associated infrastructure (US $ 367.2 MM) 11 We have been informed that this item only includes machinery/equipment and consequently, depreciation would operate as follows: a. Considering that 60% of those assets would be acquired using 24-month leasing contracts, depreciation per annum equals 50%. b. Other 40% of machinery/equipment would be directly purchased with applicable depreciation of 20%.

3.4

11 12

Tailing management facility (US $ 84.5 MM) 12

Line 538 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909. Line 539 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909.


According to mining tax provisions, movable assets might be annually depreciated to 20%. We have been informed that all assets registered under this item are machinery and consequently, cash flow should be modified as follows: ASSETS

PERCENTAGE

Tailing Management Facility Development Capital Tailing Management Facility Sustaining Capital

DEPRECIATION

100%

20%

100%

20%

Use of leasing contracts might be possible, in which case we refer to previous comments. 3.5

Project contingency (US $ 74 MM) 13 According to mining tax provisions, movable assets might be annually depreciated to 20%, while buildings and edifications to 5%. In that sense, we suggest to modify cash flow as follows:

3.6

ASSETS

PERCENTAGE

Project Contingency Project Contingency

90% 10%

DEPRECIATION 20% 5%

Owners cost (US $ 43 MM) 14 We have been informed that this item includes salaries and other payments for engineering, legal and audit services. Attending to this, cash flow should be modified showing 100% deduction for those concepts.

3.7

Accommodations (US $ 43 MM) 15 We have been informed that item “accommodations” includes both movable assets and buildings/edifications. As we indicated above, movable property such as machinery used in mining activities might be depreciated to 20% per annum, while buildings/edifications to 5% per year. However, because company has decided to use leasing contracts to acquire “accommodations”, depreciation rate may raise up to 50% per year

13 14 15

Line 546 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909. Line 547 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909. Line 542 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909.


for movable property and up to 20% for buildings/edifications. Because of this, cash flow should be modified as follows:

3.8

ASSETS

PERCENTAGE

DEPRECIATION

Accommodations Accommodations

60% 40%

20% 50%

Power line (US $ 24.5 MM) 16 According to mining tax provisions, movable assets might be annually depreciated to 20%. We have been informed that all assets registered under this item are machinery and consequently, cash flow should be modified as follows: ASSETS

PERCENTAGE

DEPRECIATION

Power Line

100%

20%

Use of leasing contracts is possible, in which case we refer to previous comments. 3.9

Access road (US $ 20.2 MM) 17 Regarding this item, company might have following choices: i. ii. iii.

Consider it as a non deductible expense, because it does not integrate company assets. Consider it as a deductible expense, inasmuch as might qualify as road infrastructure previous approval from Peruvian Public Sector. 18 Consider it as credit against Corporate Income Tax (maximum credit up to 50% of prior year tax), when special public investment provisions are accomplished. 19

Being third choice the most advantageous, it has some legal complications; notwithstanding, we have been informed that Peruvian Government is working in an easier legal framework for this benefit. In this sense, it has been announced that two projects have been approved to apply this regimen. In our calculations (Excel file, lines 569 to 572) we have applied this benefit, but it must be clear that it is new and several procedures before Peruvian Government have to be complied with.

16 17

18 19

Line 543 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909. Line 541 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909.

Mining General Law, Art. 72, Inc. d). Law 29230.


3.10 Owners cost closure (US $ 55.4 MM) 20 Deduction of this item in some specific years is explained as we understand, in company´s foresight about periods in which expenses will be executed. 3.11 Sale of Remaining Assets (US $ 36.9 MM) 21 This item considers revenues from sale of transport fleet used in company activities. If sale was in 2014, revenues would be subject to Corporate Income Tax that year. This item also considers revenues from sale of remaining equipment used in company activities. Sale will be taxed by Corporate Income Tax and Value Added Tax. Considering that equipment should be totally depreciated, no deductible cost would be recognized for Income Tax purposes. 3.12 Residual Depreciation22 Under this cash flow item we find remaining depreciation not deducted up to 2027. According to Income Tax provisions, two treatments may be granted to that amount: a.

If assets were sold, remaining depreciation would be deducted as sale cost.

b.

If assets were not useful for taxpayer, he would draw them from his accounting and deduct as expense remaining depreciation.

3.13 Value Added Tax (VAT) VAT early recovery Peruvian VAT operates under the following scheme: tax paid (19%) when taxpayer purchases goods and services, is used as credit when he performs taxed activities; however, this mechanism may be useless from financial perspective when taxpayers have long preoperative periods, inasmuch as they could not use immediately as fiscal credit the VAT paid in their acquisitions because they do not have sales or other taxed transactions. In certain cases, Peruvian tax legislation allows taxpayers to obtain the reimbursement of VAT paid on purchases made during preoperative stage. This tax benefit requires accomplish a set of procedures before Peruvian Government. 20 21 22

Line 556 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909. Line 560 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909. Line 558 of Excel page “Dep_Amort Cal” in file called “Economic Model Version 3 04b 030909.


According to provisions of Legislative Decree 973, VAT early recovery requires: a. b. c.

Sign an Investment Agreement with Peruvian Government. Invest, at least, US $5 million dollars. Preoperative stage must be no lesser than two (2) years.

After the accomplishment of these requirements, taxpayer must file a request to Tax Administration in order to get the VAT reimbursement. Attending to usual extension of recovery process, it would take around one year; however, Tax Administration is currently shortening that time. In that sense, cash flow may consider that VAT paid will be reimbursed in the same tax period. VAT payment in advance During two first preoperative years, company importations might be subject to 3.5% tax payments in advance, calculated on: CIF value + customs duties + VAT. It is possible to be exempted from this regime, but that is subject to a process before Tax Administration. Assuming that: (i) company begins exemption process and (ii) files a request to get reimbursement of tax payments in advance, usual extension of recovery process would take around one year; however, Tax Administration is currently shortening that time. In that sense, cash flow may consider that VAT paid will be reimbursed in the same tax period. 3.14 Customs duties Regarding machinery and equipment that will be imported, we have been noticed that cash flow includes customs duties and other taxes on imports. As reference, the majority of mining machinery and equipment is exempted in Peru of customs duties, but there is a limited number of items subject to 12%. In addition, we want to highlight that Peru is actively signing Free Trade Agreements with several countries such as U.S., Canada, China, etc. In that sense, it is advisable to make specific tax analysis when machinery and equipment will be imported in order to determine whether they are taxed or not. 3.15 Stability Tax Agreement


According to Chapter 9 of MGL, mining companies might celebrate contracts with Peruvian Government by which, they stabilize for fifteen (15) years tax provisions in force when investment program is approved by Mining Authority. Benefits covered by this regimen, include: a)

Income Tax, in which case, stabilized rate would be corporate rate in force plus 2%. Current corporate Income Tax rate is 30% and applying this provision, stabilized rate would be 32%. Notwithstanding stabilized corporate tax rate would be 32%, in analyzed cash flow 30% tax rate has been used because there is no certainty that a Stability Tax Agreement will be subscribed.

b) c) d)

Value Added Tax. Special provisions for tax reimbursements, as well as importation and exportation regimens. In case of exemptions and other tax benefits, stabilization is subject to provisions in force when agreement is signed.

Mining concession must accomplish following requirements in order to access to stability agreements:   3.16

Projects must have 5,000 metric tons per day as initial capacity. Investment programs must not be lesser than US $20 millions dollars.

LABOR PARTICIPATION IN CORPORATE PROFITS According to Peruvian Law, mining companies with 20 or more workers must distribute among them, 8% of net income before taxes. This percentage is applied on the same basis used to calculate Income Tax. Distribution has to occur within 30 calendar days next to file of Annual Income Tax Return. Maximum participation per worker is 18 salaries; however, this limit does not affect cash flow because any amount in excess is paid to Government.

Sincerely,

JORGE L. PICÓN PARTNER



Constancia Technical Report Ni 43 101

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