Report to:
Technical Report and Preliminary Assessment of the Zafranal Project, Perú Document No. 1295840100-REP-R0001-05
1295840100-REP-R0001-05
Report to:
TECHNICAL REPORT AND PRELIMINARY ASSESSMENT OF THE ZAFRANAL PROJECT, PERÚ
EFFECTIVE DATE: JANUARY 16, 2013
Prepared by Marinus Andre De Ruijter, P.Eng. Gregory Z. Mosher, P.Geo. Hassan Ghaffari, P.Eng. Monica Danon-Schaffer, Ph.D., P.Eng. Sabry Abdel Hafez, Ph.D., P.Eng. Carlos Guzmán Wilson Muir, P.Eng.
AS/jc
Suite 800, 555 West Hastings Street, Vancouver, British Columbia V6B 1M1 Phone: 604-408-3788 Fax: 604-408-3722
1295840100-REP-R0001-05
TABLE OF CONTENTS
1.0 SUMMARY ...... 1-1 1.1 INTRODUCTION ...... 1-1 1.2 PROPERTY DESCRIPTION ...... 1-3 1.3 HISTORY ...... 1-3 1.4 GEOLOGICAL SETTING ...... 1-4 1.5 MINERAL RESOURCES ...... 1-4 1.6 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 1-5 1.6.1 MINERAL PROCESSING ...... 1-5 1.6.2 METALLURGICAL TESTING ...... 1-5 1.7 MINING ...... 1-5 1.8 PROJECT INFRASTRUCTURE ...... 1-6 1.9 ENVIRONMENTAL ...... 1-7 1.10 CAPITAL AND OPERATING COSTS ...... 1-8 1.10.1 CAPITAL COST ESTIMATE ...... 1-8 1.10.2 OPERATING COST ESTIMATE ...... 1-8 1.11 ECONOMIC ANALYSIS ...... 1-9 1.11.1 SENSITIVITY ANALYSIS ...... 1-10 1.12 PROJECT DEVELOPMENT PLAN ...... 1-12 1.13 OPPORTUNITIES AND RECOMMENDATIONS ...... 1-12 2.0 INTRODUCTION ...... 2-1 3.0 RELIANCE ON OTHER EXPERTS ...... 3-1 4.0 PROPERTY DESCRIPTION AND LOCATION ...... 4-1 5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ...... 5-1 6.0 HISTORY ...... 6-1 6.1 ZAFRANAL MAIN ZONE ...... 6-1 6.2 VICTORIA ZONE ...... 6-2 6.3 SICERA NORTE ZONE ...... 6-2 6.4 SICERA SUR ZONE ...... 6-2 7.0 GEOLOGICAL SETTING AND MINERALIZATION ...... 7-1 7.1 REGIONAL GEOLOGY...... 7-1 7.2 PROPERTY GEOLOGY ...... 7-2 7.2.1 ZAFRANAL MAIN ZONE ...... 7-2 7.2.2 VICTORIA ZONE ...... 7-4 7.2.3 SICERA NORTE ZONE ...... 7-4
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7.2.4 SICERA SUR ZONE ...... 7-5 7.3 MINERALIZATION ...... 7-6 7.3.1 ZAFRANAL MAIN ZONE ...... 7-7 7.3.2 VICTORIA ZONE ...... 7-7 7.3.3 SICERA NORTE ZONE ...... 7-8 7.3.4 SICERA SUR ZONE ...... 7-8 8.0 DEPOSIT TYPES ...... 8-1 9.0 EXPLORATION ...... 9-1 9.1 ZAFRANAL MAIN ZONE ...... 9-1 9.2 VICTORIA ZONE ...... 9-1 9.3 SICERA NORTE ZONE ...... 9-1 9.4 SICERA SUR ZONE ...... 9-1 10.0 DRILLING ...... 10-1 10.1 ZAFRANAL MAIN ZONE ...... 10-1 10.2 VICTORIA ZONE ...... 10-2 10.3 SICERA NORTE ZONE ...... 10-3 10.4 SICERA SUR ZONE ...... 10-4 11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY ...... 11-1 11.1 REVERSE CIRCULATION SAMPLING ...... 11-1 11.2 DRILL CORE SAMPLING ...... 11-1 11.3 QUALITY ASSURANCE/QUALITY CONTROL PROCEDURES ...... 11-2 11.4 STANDARDS ...... 11-2 11.5 BLANKS ...... 11-6 11.6 DUPLICATES ...... 11-8 12.0 DATA VERIFICATION ...... 12-1 13.0 MINERAL PROCESSING AND METALLURGICAL TESTING ...... 13-1 13.1 INTRODUCTION ...... 13-1 13.2 CHRONOLOGY OF METALLURGICAL TEST WORK ...... 13-1 13.3 MINERALOGY ...... 13-2 13.4 FEED GRADE ...... 13-5 13.5 MINERALIZED MATERIAL CHARACTERISTICS ...... 13-6 13.6 GRINDABILITY ...... 13-7 13.6.1 CRUSHER WORK INDEX...... 13-7 13.6.2 ABRASION INDEX ...... 13-8 13.6.3 BOND BALL MILL WORK INDEX ...... 13-9 13.6.4 JK SAG MILL COMMINUTION TEST RESULTS ...... 13-9 13.7 FLOTATION TESTS ...... 13-10 13.7.1 BULK SULPHIDE FLOTATION ...... 13-11 13.7.2 FLOTATION PROGRAMS ...... 13-12 13.7.3 FLOTATION TESTING – GRIND SIZE EVALUATION ...... 13-12 13.7.4 FLOTATION – COLLECTOR REAGENT EVALUATION ...... 13-13
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13.7.5 FLOTATION – PH EVALUATION ...... 13-14 13.7.6 FLOTATION – ROUGHER FLOTATION TIME ...... 13-15 13.7.7 FLOTATION – REGRIND...... 13-15 13.7.8 FLOTATION – CLEANER PH EVALUATION ...... 13-16 13.7.9 FLOTATION – LOCKED CYCLE TESTING ...... 13-16 13.7.10 FLOTATION – FINAL CONCENTRATE ELEMENTAL ANALYSIS ...... 13-19 13.7.11 FLOTATION – VARIABILITY TESTS (ROUGHER) ...... 13-19 13.8 SETTLING TEST WORK (TAILINGS) ...... 13-24 13.9 ENVIRONMENTAL TESTING...... 13-24 13.10 SOLVENT EXTRACTION AND ELECTROWINNING PROCESSING FACILITY ...... 13-24 13.11 CONCLUSIONS ...... 13-24 13.12 RECOMMENDED ADDITIONAL TEST WORK ...... 13-24 14.0 MINERAL RESOURCE ESTIMATES ...... 14-1 14.1 EXPLORATORY DATA ANALYSIS ...... 14-1 14.1.1 ASSAYS ...... 14-1 14.1.2 CAPPING ...... 14-2 14.1.3 COMPOSITES ...... 14-2 14.2 BULK DENSITY ...... 14-2 14.3 GEOLOGICAL INTERPRETATION ...... 14-3 14.4 SPATIAL ANALYSIS ...... 14-6 14.5 RESOURCE BLOCK MODEL ...... 14-10 14.6 INTERPOLATION PLAN ...... 14-10 14.7 MINERAL RESOURCE CLASSIFICATION ...... 14-11 14.8 MINERAL RESOURCE TABULATION ...... 14-11 14.9 BLOCK MODEL VALIDATION ...... 14-14 15.0 MINERAL RESERVE ESTIMATES ...... 15-1 16.0 MINING METHODS ...... 16-1 16.1 INTRODUCTION ...... 16-1 16.2 INITIAL INFORMATION ...... 16-1 16.3 PIT OPTIMIZATION ...... 16-4 16.3.1 TECHNICAL AND ECONOMIC PARAMETERS ...... 16-5 16.3.2 PIT OPTIMIZATION OUTPUT ...... 16-6 16.3.3 MINING INVENTORY ...... 16-9 16.4 MINING PHASES ...... 16-12 16.5 MINING SCHEDULE ...... 16-13 16.5.1 PLANNING CRITERIA ...... 16-13 16.5.2 PRODUCTION PLAN ...... 16-14 16.6 MINE EQUIPMENT FLEET ...... 16-20 16.7 SCHEDULE OF MINE EQUIPMENT ...... 16-20 17.0 RECOVERY METHODS ...... 17-1 17.1 INTRODUCTION ...... 17-1
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17.2 SUMMARY ...... 17-1 17.3 PLANT DESIGN ...... 17-4 17.3.1 MAJOR DESIGN CRITERIA ...... 17-4 17.3.2 OPERATING SCHEDULE AND AVAILABILITY ...... 17-5 17.4 PROCESS PLANT DESCRIPTION ...... 17-5 17.4.1 PRIMARY CRUSHING ...... 17-5 17.4.2 COARSE MATERIAL STOCKPILE AND RECLAIM ...... 17-6 17.4.3 GRINDING AND CLASSIFICATION ...... 17-6 17.4.4 FLOTATION AND REGRIND CIRCUITS ...... 17-7 17.4.5 CONCENTRATE HANDLING ...... 17-10 17.4.6 TAILINGS THICKENER ...... 17-10 17.4.7 REAGENT HANDLING AND STORAGE ...... 17-11 17.4.8 ASSAY AND METALLURGICAL LABORATORY ...... 17-13 17.4.9 WATER SUPPLY ...... 17-13 17.4.10 AIR SUPPLY ...... 17-14 17.4.11 ON-LINE SAMPLE ANALYSIS ...... 17-14 17.5 SOLVENT EXTRACTION-ELECTROWINNING PROCESSING FACILITY...... 17-14 17.5.1 INTRODUCTION ...... 17-14 17.5.2 SUMMARY ...... 17-14 17.5.3 PLANT DESIGN ...... 17-15 17.6 PROCESS PLANT DESCRIPTION ...... 17-16 17.6.1 DUMP LEACHING ...... 17-16 17.6.2 SOLVENT EXTRACTION ...... 17-17 17.6.3 ELECTROWINNING CIRCUIT...... 17-18 17.7 WATER SUPPLY – REVERSE OSMOSIS PLANT ...... 17-18 18.0 PROJECT INFRASTRUCTURE ...... 18-1 18.1 INTRODUCTION ...... 18-1 18.2 PROCESS PLANT AND ANCILLARY FACILITIES ...... 18-4 18.2.1 FACILITIES DESCRIPTIONS ...... 18-5 18.3 WAREHOUSE, MAINTENANCE, AND CONCENTRATE OPERATIONS FACILITIES ...... 18-7 18.3.1 MAINTENANCE SHOP AND WAREHOUSE ...... 18-7 18.3.2 CONCENTRATOR OPERATING FACILITY ...... 18-7 18.3.3 FUEL STORAGE AND DISTRIBUTION ...... 18-7 18.4 TAILINGS MANAGEMENT FACILITY ...... 18-7 18.5 LEACH FACILITY ...... 18-8 18.6 ROADS ...... 18-9 18.6.1 OVERVIEW ...... 18-9 18.6.2 EXISTING ACCESS ROADS ...... 18-10 18.6.3 NEW ACCESS ROADS ...... 18-10 18.6.4 SITE ROADS ...... 18-10 18.7 FRESH WATER SUPPLY AND DISTRIBUTION ...... 18-11 18.7.1 SEAWATER DESALINATION PLAN ...... 18-11 18.7.2 PUMPING AND CONVEYING THE WATER TO THE MINE ...... 18-12 18.7.3 DRINKING WATER PLANT ...... 18-12 18.8 POWER SUPPLY AND DISTRIBUTION ...... 18-12
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18.9 OTHER INFRASTRUCTURE...... 18-13 19.0 MARKET STUDIES AND CONTRACTS ...... 19-1 20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT ...... 20-1 20.1 ENVIRONMENTAL STUDIES...... 20-1 20.2 ENVIRONMENTAL PERMITTING ...... 20-2 20.2.1 ENVIRONMENTAL IMPACT ASSESSMENTS ...... 20-2 20.2.2 OTHER PERMITS ...... 20-2 20.2.3 MINE CLOSURE PLAN ...... 20-3 20.2.4 ENVIRONMENTAL LEGISLATION ...... 20-3 20.2.5 PERUVIAN ENVIRONMENTAL GUIDELINES ...... 20-5 20.2.6 APPLICABLE INTERNATIONAL STANDARDS ...... 20-5 20.3 ENVIRONMENTAL MANAGEMENT PLAN ...... 20-6 20.3.1 POTENTIAL ENVIRONMENTAL IMPACTS ...... 20-6 20.4 TAILINGS MANAGEMENT ...... 20-7 20.4.1 TAILINGS DRAINAGE ...... 20-7 20.4.2 TAILINGS MONITORING ...... 20-8 20.4.3 TAILINGS FACILITY CLOSURE ...... 20-9 20.5 SULPHIDE TAILINGS MANAGEMENT ...... 20-9 20.6 MINE CLOSURE AND RECLAMATION ...... 20-10 20.6.1 PROGRESSIVE RECLAMATION ...... 20-10 20.6.2 FINAL CLOSURE ...... 20-10 20.7 SOCIAL AND COMMUNITY CONSIDERATIONS ...... 20-13 20.7.1 PROJECT LOCATION AND AREA OF DIRECT INFLUENCE ...... 20-13 20.7.2 CONSULTATION WITH COMMUNITIES ...... 20-14 20.7.3 COMMUNITY RELATIONS PLAN ...... 20-15 21.0 CAPITAL AND OPERATING COSTS ...... 21-1 21.1 CAPITAL COST ESTIMATES ...... 21-1 21.1.1 INTRODUCTION ...... 21-1 21.1.2 ESTIMATE BASE DATE AND VALIDITY PERIOD EXCHANGE RATE ...... 21-2 21.1.3 ESTIMATE APPROACH ...... 21-2 21.1.5 OWNER’S COSTS ...... 21-8 21.1.6 SUSTAINING CAPITAL ...... 21-10 21.1.8 MINING CAPITAL COST ESTIMATE ...... 21-10 21.1.9 PROCESS PLANT CAPITAL COST ESTIMATE ...... 21-14 21.1.10 TAILINGS MANAGEMENT FACILITY CAPITAL COST ESTIMATE ...... 21-14 21.1.11 LEACH FACILITY ...... 21-15 21.1.12 POWER SUPPLY ...... 21-18 21.1.13 WATER SUPPLY ...... 21-18 21.2 OPERATING COST ESTIMATES ...... 21-19 21.2.1 SUMMARY ...... 21-19 21.2.2 MINING OPERATING COST ESTIMATE ...... 21-20 21.2.3 PROCESS PLANT OPERATING COST ESTIMATE ...... 21-23 21.2.4 TAILINGS MANAGEMENT FACILITY OPERATING COST ESTIMATE ...... 21-24 21.2.5 GENERAL AND ADMINISTRATIVE COSTS ...... 21-25
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21.2.6 LEACH FACILITY OPERATING COST ESTIMATE ...... 21-27 21.2.7 WATER SUPPLY OPERATING COST ESTIMATE ...... 21-29 22.0 ECONOMIC ANALYSIS ...... 22-1 22.1 INTRODUCTION ...... 22-1 22.2 PRE-TAX MODEL ...... 22-2 22.2.1 MINE/METAL PRODUCTION IN FINANCIAL MODEL ...... 22-2 22.2.2 BASIS OF FINANCIAL EVALUATIONS ...... 22-4 22.3 SUMMARY OF FINANCIAL RESULTS ...... 22-5 22.4 SENSITIVITY ANALYSIS ...... 22-6 22.5 POST-TAX FINANCIAL ANALYSIS ...... 22-8 22.6 ROYALTIES ...... 22-10 22.7 SMELTER TERMS ...... 22-10 22.8 TRANSPORTATION LOGISTICS ...... 22-10 22.8.1 INSURANCE...... 22-11 23.0 ADJACENT PROPERTIES ...... 23-1 24.0 OTHER RELEVANT DATA AND INFORMATION ...... 24-1 24.1 PROJECT EXECUTION PLAN ...... 24-1 24.1.1 INTRODUCTION ...... 24-1 24.1.2 PROJECT APPROACH ...... 24-1 24.1.3 PROJECT EXECUTION SUMMARY ...... 24-3 24.1.4 ENGINEERING ...... 24-4 24.1.5 PROCUREMENT ...... 24-4 24.1.6 CONSTRUCTION MANAGEMENT ...... 24-5 24.1.7 CONSTRUCTION SCHEDULE ...... 24-7 24.1.8 PRE-OPERATIONAL TESTING AND START-UP ...... 24-10 25.0 INTERPRETATION AND CONCLUSIONS ...... 25-1 25.1 RISK ANALYSIS ...... 25-1 25.2 GEOLOGY ...... 25-3 25.3 MINING ...... 25-3 25.4 CONCENTRATOR ...... 25-5 25.5 LEACH FACILITY ...... 25-5 25.6 TAILINGS AND WATER MANAGEMENT ...... 25-6 25.7 OWNER’S COSTS ...... 25-6 25.8 ENVIRONMENTAL ...... 25-6 26.0 RECOMMENDATIONS ...... 26-1 26.1 SUMMARY ...... 26-1 26.2 RISK ANALYSIS ...... 26-2 26.2.1 SOCIAL/POLITICAL ...... 26-2 26.2.2 ENVIRONMENTAL ...... 26-2 26.3 MINING ...... 26-3 26.4 GEOLOGY ...... 26-3
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26.5 CONCENTRATOR ...... 26-4 26.5.1 WATER MANAGEMENT ...... 26-4 26.5.2 CLEANER FLOTATION CIRCUIT CONFIGURATION ...... 26-4 26.5.3 CYCLONE OVERFLOW/ROUGHER FLOTATION FEED ...... 26-4 26.5.4 WATER TREATMENT PLANT ...... 26-5 26.5.5 PRIMARY GRINDING CIRCUIT ...... 26-5 26.5.6 COPPER RECOVERY ...... 26-5 26.6 FUTURE METALLURGICAL TEST WORK ...... 26-5 26.7 CONCENTRATE HANDLING TRADE-OFF ...... 26-6 26.8 PRODUCT TRANSPORTATION ...... 26-6 26.9 LEACH FACILITY ...... 26-6 26.10 SITE LAYOUT ...... 26-7 26.11 TAILINGS MANAGEMENT FACILITY ...... 26-7 26.12 WATER SUPPLY ...... 26-8 26.13 POWER SUPPLY AND DISTRIBUTION ...... 26-8 26.14 ENVIRONMENTAL ...... 26-8 27.0 REFERENCES ...... 27-1 28.0 CERTIFICATES OF QUALIFIED PERSON ...... 28-1
LIST OF TABLES
Table 1.1 General Project Information ...... 1-1 Table 1.2 Zafranal Property Resource Estimate Synopsis ...... 1-4 Table 1.3 Summary of Mining Details ...... 1-6 Table 1.4 Capital Cost Summary ...... 1-8 Table 1.5 Operating Cost Summary ...... 1-9 Table 1.6 Metal Production from the Project ...... 1-10 Table 2.1 Summary of QPs ...... 2-1 Table 4.1 Zafranal Property Mineral Concessions ...... 4-3 Table 10.1 Zafranal Main Zone Assay Descriptive Statistics ...... 10-1 Table 10.2 Victoria Zone Assay Descriptive Statistics ...... 10-2 Table 10.3 Sicera Norte Zone Assay Descriptive Statistics ...... 10-3 Table 10.4 Sicera Sur Zone Assay Descriptive Statistics ...... 10-4 Table 11.1 Zafranal Property Control Samples ...... 11-2 Table 11.2 Zafranal Property Standards ...... 11-3 Table 11.3 Zafranal Property Over-Limit Standards ...... 11-4 Table 11.4 Zafranal Property Blank Descriptive Statistics ...... 11-6 Table 11.5 Zafranal Property Over-limit Blanks ...... 11-6 Table 11.6 Zafranal Property Duplicate Sample Descriptive Statistics ...... 11-8 Table 13.1 Test Work Programs and Reports ...... 13-2 Table 13.2 Mineral Composition of Phase III Composite Samples ...... 13-3 Table 13.3 Mineral Liberation of Phase III Composite Samples ...... 13-4
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Table 13.4 Head Values – Composite Samples ...... 13-5 Table 13.5 Head Values – Variability Samples ...... 13-5 Table 13.6 Specific Gravity Results by Rock Type Classification ...... 13-6 Table 13.7 Crusher Bond Work Index Results – Phase I ...... 13-7 Table 13.8 Crusher Bond Work Index Results by Rock Type Classification ...... 13-8 Table 13.9 Abrasion Index Results by Rock Type Classification ...... 13-8 Table 13.10 Bond Ball Mill Index Results by Rock Type Classification ...... 13-9 Table 13.11 Drop Weight Index Results by Rock Type Classification ...... 13-10 Table 13.12 Locked Cycle Concentrate Summary ...... 13-19 Table 13.13 Variability Tests – Comparison of Results ...... 13-22 Table 14.1 AMEC-Minproc 2011 Zafranal Main Zone Resource Estimate at 0.2% Copper Cut-off ...... 14-1 Table 14.2 Zafranal Property Drillhole and Sample Quantities...... 14-1 Table 14.3 Zafranal Mineral Zone Bulk Density Measurements ...... 14-3 Table 14.4 Zafranal Property Variograms and Search Ellipse Parameters ...... 14-7 Table 14.5 Zafranal Property Block Model Parameters ...... 14-10 Table 14.6 Zafranal Property Kriged Resource Synopsis at 0.2% Copper Cut-off ...... 14-11 Table 14.7 Zafranal Main Zone Kriged Resource Estimate at 0.2% Copper Cut-off ...... 14-12 Table 14.8 Victoria Zone Kriged Resource Estimate at 0.2% Copper Cut-off ...... 14-13 Table 14.9 Sicera Norte and Sicera Sur Zones Kriged Resource Estimates at 0.2% Copper Cut-off ...... 14-13 Table 16.1 Grade-tonnage Table for Zafranal Main...... 16-2 Table 16.2 Grade-tonnage Table for Victoria ...... 16-2 Table 16.3 Grade-tonnage Table for Sicera Sur ...... 16-3 Table 16.4 Grade-tonnage Table for Sicera Norte ...... 16-3 Table 16.5 Inter-ramp and Overall Slopes Recommended for Zafranal Main Pit...... 16-4 Table 16.6 Mining Costs ...... 16-5 Table 16.7 Whittle™ Run Results – Zafranal Main Zone ...... 16-7 Table 16.8 Whittle™ Run Results - Victoria Zone ...... 16-7 Table 16.9 Whittle™ Run Results – Sicera Norte Zone ...... 16-8 Table 16.10 Whittle™ Run Results – Sicera Sur Zone ...... 16-8 Table 16.11 Zafranal Main Zone Whittle™ Pit Mineral Resource ...... 16-9 Table 16.12 Victoria Zone Whittle™ Pit Mineral Resource ...... 16-10 Table 16.13 Sicera Norte Zone Whittle™ Pit Mineral Resource ...... 16-11 Table 16.14 Sicera Sur Zone Whittle™ Pit Mineral Resource ...... 16-11 Table 16.15 Total Mineral Inventory from Whittle™ Pits ...... 16-12 Table 16.16 80,000 t/d Mining Plan Summary By Planning Period ...... 16-15 Table 16.17 80,000 t/d Mining Plan Summary By Planning Period ...... 16-16 Table 16.18 Mining Equipment Requirements ...... 16-21 Table 17.1 Major Design Criteria...... 17-4 Table 17.2 Major Design Criteria for the Leach SX-EW Plant ...... 17-15 Table 19.1 Commercial Terms for Sale of Products ...... 19-1 Table 21.1 Capital Cost Summary ...... 21-1 Table 21.2 Details of Owner’s Costs ...... 21-9 Table 21.3 Summary of Mining Costs ...... 21-11 Table 21.4 Mining Pre-production Unit Costs ...... 21-11 Table 21.5 Initial Mining Equipment Fleet ...... 21-12 Table 21.6 Other Mining Investments ...... 21-13 Table 21.7 Process Plant Capital Cost ...... 21-14 Table 21.8 Tailings Management Facility Capital Costs ...... 21-15 Table 21.9 Leach Facility Capital Cost ...... 21-15
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Table 21.10 Power Supply and Distribution Capital Cost ...... 21-18 Table 21.11 Water Supply Capital Costs ...... 21-19 Table 21.12 Operating Cost Summary ...... 21-19 Table 21.13 Mine Unit Operating Cost Summary ...... 21-20 Table 21.14 Relevant Consumables Prices ...... 21-20 Table 21.15 Mining Operating Cost Summary ...... 21-22 Table 21.16 Concentrator Annual Operating Cost Summary ...... 21-23 Table 21.17 Tailings Management Facility Sustaining and Operating Costs ...... 21-24 Table 21.18 G&A Cost Estimate ...... 21-25 Table 21.19 Assay Laboratory Costs ...... 21-26 Table 21.20 Leach Facility Supply Costs ...... 21-27 Table 21.21 Leach Facility Operating Cost – 10,000 t/a Cathode ...... 21-28 Table 21.22 Water Supply Operating Cost Estimate ...... 21-30 Table 22.1 Metal Production from the Zafranal Project ...... 22-2 Table 22.2 Summary of Pre-tax Financial Results ...... 22-5 Table 22.3 Components of the Various Taxes ...... 22-9 Table 22.4 Summary of Post-tax Financial Results ...... 22-10 Table 25.1 Mining Plan ...... 25-4 Table 25.2 Mining Pre-production and Production Costs ...... 25-4 Table 26.1 Recommended Future Work ...... 26-1
LIST OF FIGURES
Figure 1.1 Property Location Map ...... 1-2 Figure 1.2 Pre-tax NPV Sensitivity Analysis...... 1-11 Figure 1.3 Pre-tax IRR Sensitivity Analysis ...... 1-11 Figure 1.4 Pre-tax Payback Sensitivity Analysis ...... 1-12 Figure 4.1 Property Location ...... 4-2 Figure 4.2 Zafranal Property Mineral Concessions and Mineral Zone Locations ...... 4-5 Figure 7.1 Tectonic Belts of Southeastern Perú ...... 7-1 Figure 7.2 Zafranal and Victoria Geology...... 7-3 Figure 7.3 Sicera Norte Zone Geology ...... 7-5 Figure 7.4 Sicera Sur Zone Geology ...... 7-6 Figure 13.1 Copper Head Grade versus Recovery for Rock Type ...... 13-11 Figure 13.2 LCT Procedure – Plenge ...... 13-17 Figure 13.3 LCT Procedure – G&T ...... 13-18 Figure 13.4 Rougher Recovery versus Copper Feed Grade – G&T Variability Tests ...... 13-20 Figure 13.5 Rougher Recovery versus Copper Feed Grade – Plenge Variability Tests .. 13-21 Figure 13.6 Assay Comparison, Copper Head Grade ...... 13-23 Figure 14.1 Perspective View of the Zafranal and Victoria Mineral Zones ...... 14-4 Figure 14.2 Perspective View of the Sicera Norte Mineral Zones ...... 14-5 Figure 14.3 Perspective View of the Sicera Sur Mineral Zones...... 14-5 Figure 16.1 Inter-ramp and Overall Slope Angle by Domain ...... 16-4 Figure 16.2 General Mine Configuration ...... 16-13 Figure 16.3 Total Tonnage Mining Plan ...... 16-17 Figure 16.4 Concentrator Feed ...... 16-18 Figure 16.5 Leach Facility Feed ...... 16-19
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Figure 17.1 Simplified Process Flowsheet – Zafranal Project ...... 17-3 Figure 18.1 Overall Site Layout ...... 18-2 Figure 18.2 Site Layout...... 18-3 Figure 18.3 Site View from Concentrator Location ...... 18-4 Figure 18.4 Plant Site Layout ...... 18-5 Figure 22.1 Total Metal Production ...... 22-3 Figure 22.2 Pre-tax Undiscounted Annual and Cumulative Net Cash Flows ...... 22-5 Figure 22.3 Pre-tax NPV Sensitivity Analysis...... 22-6 Figure 22.4 Pre-tax IRR Sensitivity Analysis ...... 22-7 Figure 22.5 Pre-tax Payback Sensitivity Analysis ...... 22-7 Figure 24.1 Project Management Organization Chart ...... 24-2 Figure 24.2 CM Organization Chart ...... 24-6 Figure 24.3 Preliminary Project Development Schedule Summary ...... 24-8
GLOSSARY
UNITS OF MEASURE above mean sea level ...... amsl acre ...... ac ampere ...... A annum (year) ...... a billion ...... B billion tonnes ...... Bt billion years ago ...... Ga British Thermal Unit ...... BTU centimetre ...... cm cubic centimetre ...... cm3 cubic feet per minute ...... cfm cubic feet per second ...... ft3/s cubic foot...... ft3 cubic inch ...... in3 cubic metre ...... m3 cubic yard...... yd3 Coefficients of Variation ...... CVs day ...... d days per week ...... d/wk days per year (annum) ...... d/a dead weight tonnes ...... DWT decibel adjusted ...... dBa decibel ...... dB degree ...... ° degrees Celsius ...... °C diameter ...... ø
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dollar (American) ...... US$ dollar (Canadian) ...... Cdn$ dry metric ton ...... dmt foot ...... ft gallon ...... gal gallons per minute (US) ...... gpm gigajoule...... GJ gigapascal ...... GPa gigawatt...... GW gram ...... g grams per litre ...... g/L grams per tonne ...... g/t greater than ...... > hectare (10,000 m2) ...... ha hertz ...... Hz horsepower ...... hp hour ...... h hours per day ...... h/d hours per week ...... h/wk hours per year ...... h/a inch ...... " kilo (thousand) ...... k kilogram ...... kg kilograms per cubic metre ...... kg/m3 kilograms per hour ...... kg/h kilograms per square metre ...... kg/m2 kilometre ...... km kilometres per hour ...... km/h kilopascal ...... kPa kilotonne...... kt kilovolt ...... kV kilovolt-ampere ...... kVA kilovolts ...... kV kilowatt ...... kW kilowatt hour ...... kWh kilowatt hours per tonne (metric ton) ...... kWh/t kilowatt hours per year ...... kWh/a less than...... < litre ...... L litres per minute ...... L/m megabytes per second ...... Mb/s megapascal ...... MPa megavolt-ampere ...... MVA megawatt ...... MW metre ...... m metres above sea level ...... masl
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metres Baltic sea level ...... mbsl metres per minute ...... m/min metres per second ...... m/s metric ton (tonne) ...... t microns ...... µm milligram...... mg milligrams per litre ...... mg/L millilitre ...... mL millimetre...... mm million ...... M million bank cubic metres ...... Mbm3 million bank cubic metres per annum ...... Mbm3/a million tonnes ...... Mt million years ago ...... Ma minute (plane angle) ...... ' minute (time) ...... min month ...... mo Nuevo sol (Peruvian currency) ...... S/. ounce ...... oz pascal ...... Pa centipoise ...... mPa∙s parts per million ...... ppm parts per billion ...... ppb percent ...... % pound(s) ...... lb pounds per square inch ...... psi revolutions per minute ...... rpm second (plane angle) ...... " second (time) ...... s specific gravity ...... SG square centimetre ...... cm2 square foot ...... ft2 square inch ...... in2 square kilometre ...... km2 square metre ...... m2 thousand tonnes ...... kt three dimensional ...... 3D Three Dimensional Model ...... 3DM tonne (1,000 kg) ...... t tonnes per day ...... t/d tonnes per hour ...... t/h tonnes per year ...... t/a two dimension ...... 2D volt ...... V week ...... wk weight/weight ...... w/w
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wet metric ton ...... wmt year (annum) ...... a
ABBREVIATIONS AND ACRONYMS Abeinsa Infrastructuras Medio Ambiente S.A...... ABEIMA abrasion index ...... Ai acid rock drainage ...... ARD acid-base accounting ...... ABA Amdel Mineral Laboratories ...... Amdel ammonium nitrate/fuel oil ...... ANFO AQM Copper Inc...... AQM areas of direct influence ...... ADI artisanal and small-scale miners ...... ASM Asociación Nacional del Agua ...... ANA atomic absorption spectrophotometer ...... AAS atomic absorption ...... AA audio-frequency magneto-telluric ...... ATM Australia and New Zealand Banking Group Limited ...... ANZ Autoridad Autónoma de Majes ...... Autodema bill of materials ...... BOMs Bond ball mill work index ...... BWi Bond rod mill work Index ...... RWi Certificate of Non Existence of Archaeological Remains ...... CIRA CESEL Ingenieros ...... CESEL chalcocite ...... Ch chalocopyrite ...... Cp community relations program ...... CRP Compañia Minera Zafranal S.A.C...... CMZ construction management ...... CM copper ...... Cu crusher work index ...... CWi cumulative net cash flow ...... CNCF cut-off grade ...... COG diamond drillhole ...... DDH Dirección General de Asuntos Ambientales Mineros ...... DGAAM Dirección General de Capitanias y Guardacostas del Perú ...... DICAPI Directorate the Dirección General de Asuntos Ambientales Energéticos ...... DGAAE distributed control system ...... DCS drop weight index ...... DWi Engineering, Procurement and Construction Management ...... EPCM environmental impact assessment ...... EIA Environmental Management Plan ...... EMP Free Carrier ...... FCA front end loader ...... FEL G&T Metallurgical Services Ltd...... G&T
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general and administrative ...... G&A glass reinforced plastic ...... GRP gold ...... Au health and safety management plan ...... HSMP Health, Safety, and Environmental ...... HSE heating, ventilation, and air-conditioning ...... HVAC high-density polyethylene ...... HDPE induced polarization ...... IP inductively coupled plasma ...... ICP inductively coupled plasma-spectrometer ...... ICP-MS Instituto Geológico Minero Metalúrgico ...... INGEMMET intermediate leach solution ...... ILS internal rate of return...... IRR International Finance Corporation ...... IFC International Organization for Standardization ...... ISO iron ...... Fe joint venture ...... JV Knight Piésold Ltd...... Knight Piésold Ley del Sistema Nacional de Evaluación de Impacto Ambiental ...... Ley del SEIA life-of-mine ...... LOM locked cycle tests ...... LCTs London Metal Exchange ...... LME Maintenance and Repair Contract ...... MARC material take-offs ...... MTOs Maximum Permissible Levels ...... MPLs mechanically stabilized earth ...... MSE metal leaching ...... ML methyl isobutyl carbinol...... MIBC Minera AQM Copper Perú SAC ...... AQM Perú Ministerio de Energía y Minas ...... MEM National Instrument 43-101 ...... NI 43-101 NCL Ingeniería y Construcción S.A...... NCL net cash flow ...... NCF net present value ...... NPV net smelter royalty ...... NSR non-acid generating ...... NAG operation, maintenance and surveillance manual ...... OMS ordinary kriging ...... OK organic/aqueous ...... O/A Personal Protective Equipment ...... PPE piping and instrumentation diagrams ...... P&IDs Plenge Laboratories ...... Plenge polyethersulfone ...... PES polyvinyl-pyrollidone...... PVP potassium amyl xanthate ...... PAX potentially acid generating ...... PAG
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pregnant leach solution ...... PLS preliminary economic assessment ...... PEA process design criteria ...... PDC Project Execution Plan ...... PEP project management system ...... PMS project management team ...... PMT project procedures manual ...... PPM pyrite ...... Py Qualified Person ...... QP quality assurance/quality control ...... QA/QC request for proposal ...... RFP reverse circulation ...... RC reverse osmosis ...... RO rock quality designation ...... RQD Royal Bank of Canada ...... RBC run-of-mine ...... ROM SAG Mill Comminution ...... SMC salt water reverse osmosis ...... SWRO selective mining unit...... SMU semi-autogenous grinding ...... SAG semi-detailed Environmental Impact Assessment ...... EIAsd sewage treatment plant...... STP SGS Lakefield Research Ltd...... SGS Lakefield SociedadTerral S.A...... Terral solvent extraction-electrowinning ...... SX-EW sulphur ...... S tailings management facility ...... TMF Tetra Tech Wardrop ...... Tetra Tech time-domain electromagnetic ...... TEM Traffic and Logistics ...... T&L Universal Transverse Mercator ...... UTM volume to volume ...... v/v World Geodetic System ...... WGS Zafranal Project ...... the Project Zafranal Property ...... the Property
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1.0 SUMMARY
1.1 I NTRODUCTION
Tetra Tech Wardrop (Tetra Tech) was retained by AQM Copper Inc. (AQM) to prepare a National Instrument 43-101 (NI 43-101) compliant preliminary economic assessment (PEA) for the Zafranal Project (the Project) in Perú. The Zafranal Property (the Property) is located about 90 km northwest of the city of Arequipa in southern Perú (Figure 1.1). The Property contains four zones of porphyry-style copper mineralization: Zafranal Main, Victoria, Sicera Norte and Sicera Sur. The Zafranal concentrator has been designed to process a nominal 29,200,000 t/a (80,000 t/d) of copper-gold porphyry material from an open pit operation and produce a marketable copper concentrate of 28% copper containing about 3 g/t gold. The Zafranal leach facility has been designed to treat approximately 20,000 t/d of oxide and secondary copper material to produce 10,000 t/a of copper cathode.
General information for the Project is summarized in Table 1.1.
Table 1.1 General Project Information
Description Unit Amount Estimated Mineral Resources (Measured + Indicated)1 t 557,191,000 Estimated Mineral Resources (Inferred)2 t 53,738,000 Concentrator Feed t 425,310,000 Leach Facility Feed t 87,226,000 Life-of-mine (LOM) years 15 Milling Rate t/d 80,000 Strip Ratio waste/mineralized material 1.06 Total Project Initial Capital Cost3 US$ million 1,520 Total Overall Operating Cost4 US$/t processed 8.29 Net Present Value (NPV) at 8% Discount Rate before US$ million 1,332 Tax and Government Royalities Payback After Start of Mill Production years 1.9 Internal Rate of Return (IRR) % 26.7 Note: 1Measured + Indicated Resource constrained by a Whittle economic shell is 510.7 Mt. 2Inferred Resource constrained by a Whittle economic shell is 4.9 Mt. 3This is a rounded figure. 4Includes feed for flotation and leach; operating cost for flotation only is US$8.86/t.
All dollar figures presented in this report are stated in US dollars, unless otherwise specified.
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Figure 1.1 Property Location Map
This PEA report has been prepared by Tetra Tech for AQM based on work by the following independent consultants:
• NCL Ingeniería y Construcción S.A. (NCL): commissioned by AQM to conduct the mining engineering. • Knight Piésold Ltd. (Knight Piésold): power, tailings management and overall site water balance.
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• CESEL Ingenieros (CESEL): completed a conceptual level study to determine the point at which electrical infrastructure to the Project site could reasonably interconnect with the Peruvian electric power system. • Transmin Metallurgical Consultants (Transmin): supervised metallurgical test work for the Project. • Abeinsa Infrastructuras Medio Ambiente S.A. (ABEIMA): performed engineering for the desalinated water supply and pipeline.
Each section of this report has been prepared by a Qualified Person (QP). Details of the QP’s responsibility for each section of this report are provided in Section 28.0.
1.2 P ROPERTY D ESCRIPTION
The Property is located in southern Perú about 150 km by road (90 km straight-line distance) northwest of the city of Arequipa on the border between the District of Huancarqui in the Province of Castilla and the District of Lluta in the Province of Caylloma. The approximate centre of the Property is located at 16°02’ 28” south latitude and 72°14’ 19” west longitude (Universal Transverse Mercator (UTM) 18S 794000E/8224000N using the World Geodetic System (WGS) (WGS84).
The nearest major centre is Arequipa which is serviced by scheduled flights and is the major supply center for regional mining activity. Approximately 85 km of the estimated 150 km of road between Arequipa and the Property is paved, which includes a section of the Pan-American Highway. The last 65 km is gravel, of which 15 km is via private access road constructed to access the Property. Within the Property, AQM has constructed and maintains a network of access and drill roads.
The regional climate is dry, with temperatures ranging between about 12°C in winter and 28°C in summer. Precipitation is scarce and agriculture generally possible only in river valleys with accesible irrigation. The Property is located outside zones of agricultural activity and there are no communities located on the Property. Characteristic vegetation is comprised generally of scarce grasses, cacti and shrubs. Valley bottoms within the Property are typically filled with active alluvium and are bare of vegetation.
1.3 H ISTORY
The mineral potential of the Main Zone deposit was initially discovered by artisanal small-scale miners (ASM) working the narrow gold veins in the nearby batholith formation. Through their exploration of the area, they became aware of surface exposures of copper mineralization and in 2003 brought this mineralization to the attention of Teck geologists who were conducting regional exploration for porphyry deposits. The Zafranal Main Zone was immediately staked by Teck. Currently, ASM
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activity continues on the periphery of the Zafranal porphyry systems where artisanal miners are allowed to actively mine auriferous quartz veins.
1.4 G EOLOGICAL S ETTING
Bedrock within the Property area includes Jurassic volcaniclastic and sedimentary rocks, Cretaceous granodiorite and Paleocene to late Cenozoic diorite intrusives and dikes. Most of these units have been deformed by regional-scale and local shear zones, often resulting in schistose and gneissose textures. The four mineral zones of interest, Zafranal Main, Victoria, Sicera Norte and Sicera Sur, have similar bedrock geology but differ sufficiently to warrant the separate descriptions that follow.
1.5 M INERAL R ESOURCES
Mineral resources for the Project were classified in accordance with NI 43-101 requirements. Table 1.2 summarizes the estimated Measured, Indicated, and Inferred Mineral Resources.
Table 1.2 Zafranal Property Resource Estimate Synopsis
Tonnes Cu Au (t) (%) (g/t) Kriged: All Zones Measured Zafranal Main Zone Measured Total 107,907,000 0.43 0.08 Victoria Zone Measured Total 13,184,000 0.31 0.03 Kriged: Total Measured 121,091,000 0.42 0.08 Kriged: All Zones Indicated Zafranal Main Zone Indicated Total 319,042,000 0.38 0.08 Victoria Zone Indicated Total 54,731,000 0.25 0.04 Sicera Norte Zone Indicated Total 34,005,000 0.26 0.02 Sicera Sur Zone Indicated Total 28,322,000 0.30 0.03 Kriged: Total Indicated 436,100,000 0.35 0.06 Kriged: Total Measured + Indicated 557,191,000 0.36 0.07 Kriged: All Zones Inferred Zafranal Main Zone Inferred Total 12,665,000 0.29 0.04 Victoria Zone Inferred Total 31,470,000 0.25 0.03 Sicera Norte Zone Inferred Total 6,170,000 0.25 0.02 Sicera Sur Zone Inferred Total 3,432,000 0.31 0.03 Kriged: Total Inferred 53,738,000 0.27 0.03 Note: Mineral resources which are not mineral reserves do not have demonstrated economic viability. Inferred mineral resources have a high degree of uncertainty as to their existence, and a great uncertainty as to their economic and legal feasibility. It cannot be assumed that all or any part of an Inferred resource will ever be upgraded to a higher category.
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1.6 M INERAL P ROCESSING AND M ETALLURGICAL T ESTING
1.6.1 MINERAL PROCESSING
The Project will be an open pit operation with mill feed rate of 80,000 t/d with the processing of copper-bearing plant feed based on crushing, grinding and flotation methodology. The final concentrate will be transported from site by truck to a regional port site facility. The process method includes a single semi-autogenous grinding (SAG) mill followed by two ball mills.
The tailings will be thickened to an approximate pulp density of 65% solids by weight, then transported by pipeline using gravity to the tailings management facility (TMF) roughly 10 km downhill from the process plant to a natural basin.
The recovery of copper from oxidized copper-bearing material and secondary sulphide will also be conducted. A solvent extraction-electrowinning (SX-EW) facility will produce 10,000 t of cathode copper per year from the run-of-mine (ROM) dump leaching of the mineralized material. Approximately 20,000 t/d of material with a feed grade of about 0.23% copper will be dumped on suitably prepared leach pads and treated with acid solution. The overall copper recovery is estimated to be 60%. After upgrading the copper-bearing solution by solvent extraction, the copper will be recovered by electrowinning.
1.6.2 METALLURGICAL TESTING
Metallurgical test work conducted at various laboratories has defined the process parameters required for the overall design of the concentrator. These parameters include, amongst others, a mineralogical investigation to describe the copper-bearing phases and gold present in the plant feed material, the determination of the crushability, grinding and abrasion index values required for the design of the comminution circuit, and various flotation details such as duration of flotation, the amount of reagents required, and establishing the recovery and concentrate grade of the final product. The concentrate produced will have a copper grade of 28% copper and contain about 3 g/t of gold with the respective recoveries of 87.7% and 49.0%. Limited test work was carried out to recover copper from oxide and secondary sulphide material using acid and bacterial leaching processes followed by SX-EW.
1.7 M INING
The Project will be an open pit mine utilizing a truck-and-shovel mining approach. A mine production schedule was developed for the 17-year project starting with approximately 2 years of pre-stripping followed by 15 years of production.
A variable cut-off grade of between 0.18% and 0.25% total copper was used for planning mine production to supply the concentrator. A fixed cut-off grade of 0.15%
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total copper was used for sending material to the leach pads. The annual mine production, comprised of plant feed material, leach feed material and waste, peaked at 78,208,000 t/a, with a 15-year LOM stripping ratio of 1.06. The ultimate pit design was optimized based on a Gemcom Whittle™ pit generated by using a price of $2.70/lb copper and $1,100/oz gold, then modified to incorporate haul roads and safety berms. Table 1.3 summarizes the mining details.
Table 1.3 Summary of Mining Details
Description Unit Amount Mine Life (Production) years 15 Mining Inventory t 1,056,497,000 Strip Ratio waste/mineralized material 1.06 Total Feed (Mill and Leach) t 512,566,000 Mining Operating Cost US$/t milled and leached 2.56
1.8 P ROJECT I NFRASTRUCTURE
As a comprehensive greenfield development, the Project will require the development of supporting infrastructure.
These items include:
• a process plant or concentrator that will include crushing, grinding, flotation, regrinding, concentrate filtration, concentrate thickener, tailings thickener facilities and assay laboratory • separate structures that will be erected for facilities such as a warehouse, maintenance shop, administration offices, and other supporting infrastructure • a partially-lined TMF, using geomembrane, and two structural earth dams, two waste rock dumps and a low grade sulphide stockpile • a leach facility including an SX-EW plant and heap leach pads • a network of access and on-site roads • a fresh water supply and distribution system, including a seawater desalination plant on the coast • power supply and distribution infrastructure, including a 220 kV substation • other infrastructure including truck shop, temporary construction camp, permanent camp, and laydown areas.
Details of the Project infrastructure are provided in Section 18.0.
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1.9 E NVIRONMENTAL
Based on the Ley del Sistema Nacional de Evaluación de Impacto Ambiental (Ley del SEIA) or National Environmental Impact Assessment System Law, an informal baseline study for the Project was performed in 2011 and 2012 and covered physical, biological, socio-economic, and cultural areas. Environmental settings, permits and registrations, and environmental management strategies that may be required for the Project are summarized in Section 20.0. The Project will be subject to an environmental assessment in the future. A conceptual level Plan de Manejo Ambiental or Environmental Management Plan (EMP), was prepared and potential impacts associated with the Zafranal TMF were identified as part of this PEA. Further details are discussed in Section 20.0.
A more detailed environmental management and closure plan for the Project will be developed in the prefeasability study. However, a cost estimate has been included in this PEA as part of the capital and operating cost estimates.
A preliminary environmental baseline study was performed for the Project. This study will be used to define the scope of work for the formal baseline study, which will commence during the next stage of project evaluation and is mandatory in order to complete an obligatory Environmental Impact Assessment (EIA) and secure the mine operating permits. AQM has had a limited EMP in place for the past 34 months as a result of a semi-detailed Environmental Impact Assessment (EIAsd) that was submitted for exploration activities only, and subsequently approved for the Project by the Peruvian government.
There are Peruvian mining environmental guidelines provided by the Mining and Energy Ministry that cover monitoring protocols for air and water quality, water management, acid mine drainage management, evaluation of impacts resulting from mining-metallurgical activities and others.
With regards to tailings water management, free water from discharged tailings and seasonal runoff will collect in a small pond in the south end of the tailings impoundment, from which the water will evaporate. No operational spillway will be required for the impoundment during operations; however, a spillway will be constructed at closure. Sufficient freeboard to contain extreme precipitation events will be maintained during mine operations. The TMF will be a zero-discharge facility and the tailings embankments will incorporate a geomembrane liner and grout curtains to minimize seepage to the environment (as required).
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1.10 C APITAL AND O PERATING C OSTS
1.10.1 CAPITAL COST ESTIMATE
As summarized in Table 1.4, an initial capital cost of US$1,519.7 million is estimated for the Project. All currencies in this section are expressed in US dollars. The accuracy range of the capital cost estimate is ±35%.
This estimate is prepared with a base date of Q3 2012 and does not include any escalation past this date. The quotations used in this estimate were obtained in Q3 2012, and are budgetary and non-binding.
CESEL provided some equipment prices and the remainder were obtained from recent historical data. Freight costs and taxes were provided by AQM. For non- major equipment (i.e. equipment less than $100,000), costing was based on in-house data or quotes from recent similar projects.
All equipment and material costs include Free Carrier (FCA) manufacturer plant Incoterms® 2010. Other costs such as spares, taxes, duties, freight, and packaging were covered separately in the Indirects section of the estimate.
Table 1.4 Capital Cost Summary
Capital Cost Description ($ '000) Mining and Pre-production Development 367,689 Process Plant 240,995 Leach Facility 37,099 Tailing and Water Management 55,108 Plant Infrastructure 115,340 Fresh Water Supply 204,535 Power Supply and Distribution 50,255 Indirects Owner’s Costs 67,778 Indirects 177,961 Contingency (18%) 202,912 Total Capital Cost 1,519,672
1.10.2 OPERATING COST ESTIMATE
On-site operating costs are estimated to be US$8.29/t of mineralized material processed and include mining, processing, leaching, general and administrative (G&A), and plant services cost. The unit costs summarized in Table 1.5 are based on treatment of 425,310,000 t by flotation and 87,256,000 t by leaching and 365 days of operation.
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Table 1.5 Operating Cost Summary
Unit Cost (US/$/t milled Area and leached) Mining 2.56 Processing 4.23 Leaching 0.94 G&A 0.49 Tailings and Water Management 0.07 Total Operating Cost 8.29
Details of the Project’s capital and operating costs are provided in Section 21.0.
1.11 E CONOMIC A NALYSIS
An economic evaluation of the Project was prepared by Tetra Tech based on a pre- tax financial model. All currencies in this section are expressed in US dollars. For the 15-year mine life, the following pre-tax financial parameters were calculated using the base case metal prices of $3.00/lb copper and $1,274/oz gold:
• 26.7% IRR • 1.9-year payback on an initial capital cost of $1,520 million • $1,332 million NPV using an 8% discount rate.
AQM commissioned Ernst & Young, Perú to prepare a tax model for the a post-tax economic evaluation of the Project with the inclusion of applicable taxes and the Peruvian government’s mining royalty (Section 22.0). The post-tax financial results are as follow:
• 17.4% IRR • 2.6-year payback on an initial capital cost of $1,520 million • $588 million NPV using an 8% discount rate.
Metal revenues projected in the cash flow models were based on the average metal values indicated in Table 1.6.
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Table 1.6 Metal Production from the Project
Description Value Mine Life (Years) 151 Material Milled/Leached Total Tonnes to Mill ('000) 425,310 Average Annual Tonnes to Mill ('000) 28,354 Total Tonnes to Leaching (‘000) 87,256 Average Annual Tonnes to Leaching (‘000) 6,712 Average Grade Copper (%) – Mill 0.378 Gold (g/t) – Mill 0.071 Copper (%) – Leaching 0.230 Gold (g/t) – Leaching 0.085 Total Production Copper (‘000 lb) – Mill 3,105,452 Gold (‘000 oz) – Mill 479 Copper (‘000 lb) – Leaching 265,863 Average Annual Production Copper (‘000 lb) – Mill 207,030 Gold (‘000 oz) – Mill 32 Copper (‘000 lb) – Leaching 20,451 Note: 1Leaching is active for 13 years.
1.11.1 SENSITIVITY ANALYSIS
Sensitivity of the Project’s pre-tax NPV, IRR to the Project key variables was investigated. Using the base case as a reference, each of the key variables was changed between ±30% at 10% intervals while holding the other variables constant. The following key variables were investigated:
• copper price • gold price • capital costs • operating costs.
As shown in Figure 1.2, the Project NPV, calculated at an 8% discount, is most sensitive to the copper price and, in decreasing order operating costs, capital costs, and gold price.
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Figure 1.2 Pre-tax NPV Sensitivity Analysis
3,500 3,000 2,500 2,000 Copper price 1,500 Gold price Capital costs 1,000 Operating costs 500 0 NPV@8% Discount Rate (US$M) Rate Discount NPV@8% -30% -20% -10% 0% 10% 20% 30% -500 % Change from Base Case
As shown in Figure 1.3, the Project IRR is most sensitive to the copper price followed by the capital costs, operating costs and gold price.
Figure 1.3 Pre-tax IRR Sensitivity Analysis
45% 40% 35% 30% Copper price 25% Gold price 20% Capital costs 15% Operating costs 10%
Internal Rate Rate of Internal Return (%) 5% 0% -30% -20% -10% 0% 10% 20% 30% % Change from Base Case
As shown in Figure 1.4, the payback period is most sensitive to the copper price and less sensitive to the rest of parameters.
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Figure 1.4 Pre-tax Payback Sensitivity Analysis
6.0
5.0
4.0 Copper price 3.0 Gold price Capital costs 2.0 Operating costs
Payback Period (years) Period Payback 1.0
0.0 -30% -20% -10% 0% 10% 20% 30% % Change from Base Case
1.12 P ROJECT D EVELOPMENT P LAN
The Project will take approximately three years of construction activities to complete from the time all engineering is completed and permits granted. A high-level project development plan and schedule is provided in Section 24.0.
1.13 O PPORTUNITIES AND R ECOMMENDATIONS
Based on the work carried out in this PEA and the resultant economic evaluation, this study should be followed by an additional study in order to further assess the economic viability of the Project.
Detailed opportunities and recommendations for the Project are provided in Section 26.0.
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2.0 INTRODUCTION
Tetra Tech, in collaboration with NCL and Knight Piésold, prepared this technical report for the Project in general accordance with the guidelines provided in NI 43-101 Standards of Disclosure for Mineral Projects.
A summary of QPs responsible for each section of this report is provided in Table 2.1.
Table 2.1 Summary of QPs
Report Section Company QP 1.0 Summary All Sign off by Section 2.0 Introduction Tetra Tech Hassan Ghaffari, P.Eng. 3.0 Reliance on Other Experts Tetra Tech Hassan Ghaffari, P.Eng. 4.0 Property Description and Location Tetra Tech Gregory Z. Mosher, P.Geo. 5.0 Accessibility, Climate, Local Resources, Tetra Tech Gregory Z. Mosher, P.Geo. Infrastructure and Physiography 6.0 History Tetra Tech Gregory Z. Mosher, P.Geo. 7.0 Geological Setting and Mineralization Tetra Tech Gregory Z. Mosher, P.Geo. 8.0 Deposit Types Tetra Tech Gregory Z. Mosher, P.Geo. 9.0 Exploration Tetra Tech Gregory Z. Mosher, P.Geo. 10.0 Drilling Tetra Tech Gregory Z. Mosher, P.Geo. 11.0 Sample Preparation, Analyses, and Security Tetra Tech Gregory Z. Mosher, P.Geo. 12.0 Data Verification Tetra Tech Gregory Z. Mosher, P.Geo. 13.0 Mineral Processing and Metallurgical Testing Tetra Tech Marinus Andre De Ruijter, P.Eng. 14.0 Mineral Resource Estimates Tetra Tech Gregory Z. Mosher, P.Geo. 15.0 Mineral Reserve Estimates NCL Carlos Guzmán 16.0 Mining Methods NCL Carlos Guzmán 17.0 Recovery Methods Tetra Tech Marinus Andre De Ruijter, P.Eng. 18.0 Project Infrastructure - - 18.1 Introduction Tetra Tech Hassan Ghaffari, P.Eng. 18.2 Process Plant and Ancillary Tetra Tech Hassan Ghaffari, P.Eng. Facilities 18.3 Warehouse, Maintenance, and Tetra Tech Hassan Ghaffari, P.Eng. Concentrate Operations Facilities 18.4 Tailings Management Facility Knight Piésold Wilson Muir, P.Eng. 18.5 Leach Facility Tetra Tech Hassan Ghaffari, P.Eng. 18.6 Roads Tetra Tech Hassan Ghaffari, P.Eng. 18.7 Fresh Water Supply and Tetra Tech Hassan Ghaffari, P.Eng. Distribution table continues…
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Report Section Company QP 18.8 Power Supply and Distribution Knight Piésold Wilson Muir, P.Eng. 18.9 Other Infrastructure Tetra Tech Hassan Ghaffari, P.Eng. 19.0 Market Studies and Contracts Tetra Tech Sabry Abdel Hafez, Ph.D., P.Eng. 20.0 Environmental Studies, Permitting and Tetra Tech Monica Danon-Schaffer, Ph.D., Social or Community Impact P.Eng. 21.0 Capital and Operating Costs - - 21.1 Capital Cost Estimates - - 21.1.1 Introduction Tetra Tech Hassan Ghaffari, P.Eng. 21.1.2 Estimate Base Date and Validity Tetra Tech Hassan Ghaffari, P.Eng. Period Exchange Rate 21.1.3 Estimate Approach Tetra Tech Hassan Ghaffari, P.Eng. 21.1.4 Cost Basis by Project Tetra Tech Hassan Ghaffari, P.Eng. 21.1.5 Owner’s Costs Tetra Tech Hassan Ghaffari, P.Eng. 21.1.6 Sustaining Capital Tetra Tech Hassan Ghaffari, P.Eng. 21.1.7 Elements of Costs Tetra Tech Hassan Ghaffari, P.Eng. 21.1.8 Mining Capital Cost Estimate NCL Carlos Guzmán 21.1.9 Process Plant Capital Cost Tetra Tech Hassan Ghaffari, P.Eng. Estimate 21.1.10 Tailings Management Facility Knight Piésold Wilson Muir, P.Eng. Capital Cost Estimate 21.1.11 Leach Facility Tetra Tech Hassan Ghaffari, P.Eng. 21.1.12 Power Supply Knight Piésold Wilson Muir, P.Eng. 21.1.13 Water Supply Tetra Tech Hassan Ghaffari, P.Eng. 21.2 Operating Cost Estimates - - 21.2.1 Summary Tetra Tech Marinus Andre De Ruijter, P.Eng. 21.2.2 Mining Operating Cost Estimate NCL Carlos Guzmán 21.2.3 Process Plant Operating Cost Tetra Tech Marinus Andre De Ruijter, P.Eng. Estimate 21.2.4 Tailings Management Facility Knight Piésold Wilson Muir, P.Eng. Operating Cost Estimate 21.2.5 General and Administrative Costs Tetra Tech Marinus Andre De Ruijter, P.Eng. 21.2.6 Leach Facility Operating Cost Tetra Tech Marinus Andre De Ruijter, P.Eng. Estimate 21.2.7 Water Supply Operating Cost Tetra Tech Marinus Andre De Ruijter, P.Eng. Estimate 22.0 Economic Analysis Tetra Tech Sabry Abdel Hafez, Ph.D., P.Eng. 23.0 Adjacent Properties Tetra Tech Gregory Z. Mosher, P.Geo. 24.0 Other Relevant Data and Information Tetra Tech Hassan Ghaffari, P.Eng. 25.0 Interpretations and Conclusions All Sign off by Section 26.0 Recommendations All Sign off by Section 27.0 References All Sign off by Section 28.0 Certificates of Qualified Persons All Sign off by Section
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3.0 RELIANCE ON OTHER EXPERTS
In preparation of this report, Tetra Tech has relied upon others for information, and disclaims responsibility for information derived from reports pertaining to mineral tenure, property ownership, surface rights, environment, royalties, and social issues. Neither Tetra Tech nor the authors are qualified to provide extensive comment on legal issues, including mineral tenure status associated with the Project, and therefore, ownership information is provided for information purposes only.
Tetra Tech has reviewed and analyzed data and reports provided by AQM, together with publicly available information augmented by direct field examination, drawing its own conclusions. Information from third-party sources is quoted or referenced throughout this technical report and is disclosed in Section 27.0. Tetra Tech used (or relied upon) information from these sources under the assumption that the contents are accurate.
The QPs who prepared this report relied upon information provided by the following experts who are not QPs for the purposes of this report:
• Mr. Alvaro Fernandez-Baca, Consulting Geologist of AQM has been relied on for advice on matters relating to geological exploration, drilling and quality assurance/quality control (QA/QC). • Mr. Lorenzo de la Puente, General Manager of AQM Perú has been relied on for advice on matters relating to the mineral concessions and ownership of the Project and permitting requirements pertaining to the Project. • Mr. Fernando Tori Vargas and Ms. Jessica Espinola, of Ernst & Young, Perú in Lima Perú, have been relied on for advice on matters relating to taxes in the economic modelling. Mr. Tori Vargas is a Partner in the Tax Services department. Ms. Jessica Espinola is a Senior Tax Manager and conducted the detailed review of the post-tax model. • Mr. Carlos Derpsch of Ayrmin Consultorias y Representacions Ltda (Ayrmin) located in Santiago Chile, has been relied on for advice on matters relating to smelting, refining and transportation terms for copper concentrate as well as copper cathode.
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4.0 PROPERTY DESCRIPTION AND LOCATION
The Property is located in southern Perú about 150 km by road (90 km straight-line distance) northwest of the city of Arequipa on the border between the District of Huancarqui in the Province of Castilla and the District of Lluta in the Province of Caylloma. The approximate centre of the Property is located at 16°02’ 28” south latitude and 72°14’ 19” west longitude UTM 18S 794000E/8224000N using WGS84 (Figure 4.1). The datum selected for the Project reporting is WGS84 coordinate system using the UTM 18S horizontal projection. Earlier work was performed using the datum of PSAD 56 and where appropriate this work has been re-stated in WGS84.
The Property is 33,552.19 ha in area and is comprised of 33 concessions and ten mining claims, the details of which are presented in Table 4.1 and shown in Figure 4.2.
Ownership of a concession carries the right to both explore and exploit a property, although additional permits are necessary prior to obtaining the actual right to mine. Concessions are held for as long as annual “Mining Good Standing Fees” are paid. Failure to pay the annual “Mining Good Standing Fees” for two consecutive years will result in the mining title being terminated.
Table 4.1 shows the next annual “Mining Good Standing Fees” payment, which is June 30, 2013. By that time, the annual “Mining Good Standing Fees” will be paid for the 2013 to 2014 tax year to keep the concessions in good standing for the following two years.
For the purposes of this report, there are four mineral zones of interest: Zafranal Main, Victoria, Sicera Norte and Sicera Sur. The approximate locations of these zones are also shown on Figure 4.2.
In 2009, AQM, through its wholly-owned Peruvian subsidiary Minera AQM Copper Perú SAC (AQM Perú), acquired from Teck Perú the right to earn a 51% interest in the Property. Teck Perú retained the right to earn back to a 60% interest in the Property. In the event Teck Perú did not elect to exercise the back-in, AQM had the right to increase its interest up to 100%.
In 2010, the option agreement was amended whereby AQM Perú and Teck Perú formed a 50/50 joint venture (JV) to explore and develop the Property. AQM was immediately vested to a 50% interest in the Property and all cash payment obligations held by AQM, and all royalties and back-in rights held by Teck Perú were cancelled in exchange for the granting of five million AQM shares to Teck Perú and the commitment by AQM to fund $10.7 million in exploration expenditures in addition to the $7.5 million already spent by AQM on the Property. Under the terms of the JV
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agreement, Teck Perú has the right to become project operator when a production decision has been made. In the meantime, AQM is the Project operator.
The Property includes the Chicharron option (4,700 ha) that Teck Perú purchased from BHBP Minerals but which remains subject to a 1.5% net smelter royalty (NSR) capped at $2.0 million. None of the four mineral zones discussed in this report falls within the Chicharron option area.
Figure 4.1 Property Location
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Table 4.1 Zafranal Property Mineral Concessions
Next Annual “Mining Good Area Unique Standing Fee” Name (ha) Code Holder Payment Zafranal Concessions AMALIA GUILLERMINA 200.00 010172503 AQM 30/06/2013 CAMPANERO 1 1,000.00 010278208 CMZ 30/06/2013 CAMPANERO 2 400.00 010278108 CMZ 30/06/2013 CHARO 1 1,000.00 010554007 CMZ 30/06/2013 CHICHARRON_11 1,000.00 010210403 CMZ 30/06/2013 CHICHARRON_N_5 1,000.00 010209003 CMZ 30/06/2013 CHICHARRON_N_6 700.00 010209103 CMZ 30/06/2013 CHICHARRON_N_7 1,000.00 010209203 CMZ 30/06/2013 CHICHARRON_N_8 1,000.00 010209303 CMZ 30/06/2013 SICERA 1 1,000.00 010248903 CMZ 30/06/2013 SICERA 2 500.00 010295003 CMZ 30/06/2013 SICERA 3 900.00 010313703 CMZ 30/06/2013 SICERA 4 1,000.00 010330303 CMZ 30/06/2013 ZAFRANAL 1 1,000.00 010135403 CMZ 30/06/2013 ZAFRANAL 10 600.00 010360803 CMZ 30/06/2013 ZAFRANAL 11 600.00 010360903 CMZ 30/06/2013 ZAFRANAL 12 1,000.00 010260704 CMZ 30/06/2013 ZAFRANAL 13 1,000.00 010260804 CMZ 30/06/2013 ZAFRANAL 14 1,000.00 010260904 CMZ 30/06/2013 ZAFRANAL 15 500.00 010261004 CMZ 30/06/2013 ZAFRANAL 16 1,000.00 010261104 CMZ 30/06/2013 ZAFRANAL 17 1,000.00 010261204 CMZ 30/06/2013 ZAFRANAL 18 1,000.00 010261304 CMZ 30/06/2013 ZAFRANAL 2 27.20 010175103 CMZ 30/06/2013 ZAFRANAL 21 1,000.00 010261604 CMZ 30/06/2013 ZAFRANAL 3 525.00 010175303 CMZ 30/06/2013 ZAFRANAL 34 1,000.00 010262904 CMZ 30/06/2013 ZAFRANAL 35 700.00 010263004 CMZ 30/06/2013 ZAFRANAL 36 500.00 010263104 CMZ 30/06/2013 ZAFRANAL 4 799.99 010269403 CMZ 30/06/2013 ZAFRANAL 7 1,000.00 010313803 CMZ 30/06/2013 ZAFRANAL 8 500.00 010340003 CMZ 30/06/2013 ZAFRANAL 9 500.00 010357503 CMZ 30/06/2013 Total Area 25,952.20 - - - table continues…
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Next Annual “Mining Good Area Unique Standing Fee” Name (ha) Code Holder Payment Zafranal Mining Claims AQM IV 500.00 010033810 AQM 30/06/2013 AQP I 800.00 010209809 AQM 30/06/2013 AQP II 500.00 010209909 AQM 30/06/2013 AQP III 900.00 010210009 AQM 30/06/2013 AQP V 800.00 010145910 AQM 30/06/2013 AQP VII 1,000.00 010294310 AQM 30/06/2013 AQP VIII 900.00 010479310 AQM 30/06/2013 AQP X 500.00 010015111 AQM 30/06/2013 AQP XI 700.00 010166411 AQM 30/06/2013 AQP XII 1,000.00 010166311 AQM 30/06/2013 Total Area 7,600.00 - - - Note: CMZ = Compañia Minera Zafranal S.A.C.
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Figure 4.2 Zafranal Property Mineral Concessions and Mineral Zone Locations
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The main permits, required to carry out all current and planned exploration work including drilling activities are the EIAsd, the authorization for water use, and the surface access rights to the Property. The EIAsd permit currently in place is valid until August 1, 2013 and allows for drilling up to 226 platforms on the Zafranal Main, Victoria, Sicera Norte and Sicera Sur zones. A modification to the EIAsd permit to allow drilling additional areas on the Property is in progress at the time this report was prepared by Tetra Tech.
Surface rights to the Property are held by the Autoridad Autónoma de Majes (Autodema), the agency of the Arequipa Regional Government that manages the Majes Siguas Special Project. AQM acquires the right of access to the Property from Autodema on an annual basis through a renewable contract which was recently renewed and became effective on April 1, 2012. A fee equivalent to $30.00/ha is payable on an annual basis to maintain access rights to the Property. In April 2012, AQM paid Autodema $167,000 for the right of access to 5,531 ha of mining concessions currently being explored on the Property.
A “Mining Good Standing Fee” equivalent to $3.00/ha is also payable on an annual basis to maintain the concessions in good standing. In addition, a penalty of $6.00/ha is charged for those concessions older than seven years that have not recorded a minimum investment by AQM. In June 2012, AQM paid a total of $199,000 in concession fees.
To record an investment, AQM is required to file an annual information form detailing the investments in mining activities in the Property and the amount of money spent on sustainability development programs, including community relations and environmental management programs.
In compliance with the recently approved regulation to formalize artisanal and small mining activity, AQM is engaged in a process to sign an exploitation agreement with the company formed by approximately 200 artisanal miners who are operating within the Property (in areas of no interest to the Project). Additionally, a closure plan is required to be submitted to the authority for the reclamation of the works areas disturbed by the artisanal miners. The informal miners have up to two years to comply with all the requirements established by the current legislation to be formalized as artisanal or small miners.
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5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY
The Property is located in the Cordillera Occidental about 90 km (150 km by road) northwest of Arequipa. The Property area is characterized by steep, dissected topography with elevations between about 1,000 and 3,300 masl. The four mineral zones of interest to this report are located between elevations of 2,000 and 2,850 masl. The Property is well-removed from agricultural areas and does not contain any communities or settled areas.
The climate in this part of Perú is dry, with temperatures ranging between about 12°C in winter and 28°C in summer. The climate is such that exploration and mining can readily be carried out on a year-round basis. Rain is scarce and agriculture is possible only in river valleys where irrigation is available. However, occasionally the area is subject to heavy downpours and flash flooding in the river valleys. Characteristic vegetation on the Property is comprised primarily of grasses, cacti and shrubs. Valley bottoms within the Property are typically filled with active alluvium and are bare of vegetation.
The nearest major centre is the city of Arequipa which is serviced by scheduled flights and is the major supply center for mining activity in this part of Perú. About 85 km of the 150 km of roadway from Arequipa and the Property is paved, including the Pan-American Highway. The last 65 km is by gravel road of which 15 km are via a private access road originally constructed by Teck Resources. Within the Property, AQM has constructed and maintains a network of access and exploration roads.
Through an agreement with Autodema, AQM has rights to explore and exploit the Property subject to securing the necessary environmental and sectorial permits. As the Property and surrounding area is virtually uninhabited and the land used for no other purpose, it is improbable that any difficulty would be encountered in obtaining the necessary surface rights for any contemplated mining activity.
There are sufficient areas within the Property for processing plants as well as storage of mining waste and tailings disposal as well as leach pads for material determined to be viable for this means of metal extraction.
Water does not appear readily available within the Property. For the purposes of this report, it was assumed that the water for the Project will be supplied from a saltwater desalination plant on the shores of the Pacific Ocean, approximately 115 km to the southwest of the proposed plant site.
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Electricity would be obtained from a national grid substation which is accessible within 115 km of the location currently considered most suitable for a processing plant. Whether sufficient excess electricity would be available for a new mining operation is not known as a power deficit is forecast for the area; however, additional power lines and upgrades of existing lines are under construction and several hydroelectric projects and gas fired power stations are also planned which would mitigate these forecasted shortfalls in power supply.
Skilled mining personnel are currently available in the general area. Arequipa is a mining district with a number of large operating mines, however a deficit of trades people is forecasted as the mining industry grows within the region. Expected shortages of skilled labour in the future will need to be mitigated by extensive training and recruiting programs.
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6.0 HISTORY
Artisanal miners actively mine for gold in the auriferous quartz veins on the periphery of the Zafranal porphyry systems. It is not known when the artisanal miners began working in this area but they were well aware of surface exposures of copper mineralization in what is now the Zafranal Main Zone. In 2003, these miners brought this mineralization to the attention of Teck geologists who were conducting regional exploration for porphyry deposits. The Zafranal Main Zone was immediately staked by Teck. With the exception of gold production by artisanal miners, there has been no commercial production from the Property.
6.1 Z AFRANAL M AIN Z ONE
Teck explored the Zafranal Main Zone during the period 2003 through 2005, starting with surface rock sampling that confirmed the extensive presence of porphyry-style geology and alteration and copper and gold mineralization.
In 2004 Teck drilled 22 reverse circulation (RC) holes (7,000 aggregate metres) and four core holes (1,556 aggregate metres). The RC drilling outlined a 1 km- long enriched copper blanket in porphyritic dioritic intrusive rocks. Mineralized intercepts ranged from a minimum thickness of 4 m with an average grade of 0.62% copper (ZFRC04-016) to a maximum intercept of 110 m with an average grade of 1.02% copper (ZFRC04-008).
The core holes were drilled to confirm the RC drill results. Diamond drillhole (DDH) DDH04-001 which was a twin of ZFRC04-008 intersected 166 m with an average grade of 1% copper. Two other holes intersected significant intervals of copper mineralization (ZFDDH04-002 intersected 92 m with an average grade of 0.94% copper and ZFDDH04-004 intersected 77 m with an average grade of 1.80% copper).
A further 10 RC holes were drilled in 2005 to test for possible eastern, northern and western extensions of mineralization intersected by previous drill campaigns. Extensions to the then-known limits of mineralization were found in all three directions.
In 2004, following the first phase of drilling, orientation-scale time-domain electromagnetic (TEM) and audio-frequency magneto-telluric (ATM) surveys were carried out. The ATM survey was successful in differentiating between altered and unaltered and mineralized and unmineralized rocks.
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Between 2006 and 2007, Teck conducted scout drilling in peripheral areas of the Property, looking for possible porphyry targets concealed beneath the Tertiary cover. No further work was done on the Zafranal Main Zone between 2005 and May 2009 when AQM acquired an option on the Property.
6.2 V ICTORIA Z ONE
The Victoria Zone is contiguous with the eastern end of the Zafranal Main Zone but was not discovered until 2011 as AQM extended their drill coverage of the Zafranal Main Zone in that direction.
6.3 S ICERA N ORTE Z ONE
The Sicera Norte Zone is located about 10 km northwest of the Zafranal Main Zone and was identified and explored by Teck in 2004 when they carried out surface sampling and geological mapping. This work identified the geological setting as being similar to other porphyry occurrences in the area and established the general distribution of surface copper and gold mineralization.
No further exploration took place here until AQM began their current exploration program in 2010. Teck did not own the concession hosting Sicera Norte, as it was held by an individual until purchased by AQM in 2010.
6.4 S ICERA S UR Z ONE
Sicera Sur is located about 5 km south of Sicera Norte and in the 1990s was owned by Compañia Minera Milpo S.A.A. who carried out surface sampling. During the mid- 1990s, Phelps Dodge optioned the Property, carried out additional surface sampling and drilled two core holes, the results of which did not encourage them to continue exploring the prospect.
Teck optioned the Property from Phelps Dodge in 2004, confirmed the surface sampling results obtained by Phelps Dodge, and drilled four RC holes to test for the possible presence of mineralization beneath gravels that border the area of exposed mineralization. Three of these holes were drilled in an area where faulting has placed bedrock over recent gravels and so were unsuccessful in testing for the presence of mineralization beneath the gravel cover. The fourth hole (CH07RC-011) was drilled within 120 m of outcropping mineralization and intersected 120 m of copper mineralization with an average grade of 0.23% copper.
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7.0 GEOLOGICAL SETTING AND MINERALIZATION
7.1 R EGIONAL G EOLOGY
Southern Perú is divisible into three major, northwest-trending tectonic belts, from east to west, the Sierra Oriental, Altiplano and Sierra Occidental. All three developed as a result of the subduction of the Nazca oceanic plate beneath the continental South American Plate. Subduction is estimated to have begun during the Late Triassic, about 200 Ma and the ongoing process is termed the Andean Orogeny (Figure 7.1).
Figure 7.1 Tectonic Belts of Southeastern Perú
Source: Rivera et al 2010.
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The Property is located within the Sierra Occidental and is underlain by thin Paleozoic-age sedimentary strata and by Mesozoic-age intrusive rocks.
Subduction-related igneous intrusive and related volcanic rocks were emplaced into the Sierra Occidental throughout Mesozoic and Cenozoic time. The Coastal Batholith, of high-potassium granitoid composition and of Lower Jurassic to Upper Eocene age and comprised of multiple intrusions, is the dominant feature of the Cordillera Occide. Major intrusive periods have been dated to around 100, 80, and 75 to 60 Ma before present day.
Cenozoic-age volcanic and volcaniclastic rocks cap the Cordillera Oriental.
The distribution of mineralization within the Property is controlled by the Iquipi- Clavelinas Fault System, an east-to-northwest-trending regional-scale system of transcurrent faults that are related to displacements caused by subduction. These faults are deep-seated and movement on them within the Property area appears to have developed a dilational shear system that focused the emplacement of both the host diorites as well as the porphyry-style copper mineralization that is contained within them.
7.2 P ROPERTY G EOLOGY
Bedrock within the Property area includes deformed and metamorphosed intrusive and mafic volcanic rocks of undetermined age, Jurassic volcaniclastic and sedimentary rocks, Cretaceous granodiorite and Paleocene to late Cenozoic diorite intrusives and dikes. Most of these units have been deformed by regional-scale and local shear zones, often resulting in schistose and gneissose textures. The four mineral zones of interest, Zafranal Main, Victoria, Sicera Norte and Sicera Sur, have similar bedrock geology but differ sufficiently to warrant the separate descriptions that follow.
7.2.1 ZAFRANAL MAIN ZONE
The Zafranal Main Zone is located within an east-trending block that is bounded by faults that belong to the Iquipi-Clavelinas Fault System. This block is bounded to the north by gneissose rocks with local mylonitic bands that formed as a result of at least two phases of deformation, including movement on the Iquipi-Clavelinas Fault, and to the south by sedimentary rocks, primarily quartzite, siltstone and wacke, tentatively assigned to the Upper Jurassic to Lower Cretaceous-age Yura Group that are also tectonically deformed (Figure 7.2).
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Figure 7.2 Zafranal and Victoria Geology
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The fault block contains volcano-sedimentary rocks of the Lower Jurassic-age Guaneros Formation that have been intruded by several phases of diorite to Quartz- Diorite plugs and dikes of Upper Cretaceous age. The Guaneros Formation is comprised of siltstone, sandstone, debris flows, tuff, volcanic breccia and andesitic flows together with sub-volcanic units of andesitic composition.
Four phases of dioritic intrusive rocks have been identified: Zafranal Diorite, Microdiorite, Quartz-diorite and Post-mineral Diorites.
The Zafranal Diorite is the oldest and most abundant of the intrusive rocks, is greenish-grey and porphyritic, and occurs as stocks and dykes. The emplacement of the Zafranal Diorite is interpreted to have been accompanied by an early hypogene copper-mineralizing phase of mineralization.
The Microdiorite is fine-crystalline and greenish and occurs at the eastern and western ends of the Zafranal Main Zone. The Microdiorite cuts the Zafranal Diorite and its emplacement appears to coincide with the main copper-mineralizing event.
The Quartz-diorite is dark grey, commonly contains disseminated pyrite and magnetite, and cuts both the Zafranal Diorite and Microdiorite. The Quartz-diorite characteristically contains low-grade hypogene mineralization.
Post-mineral dikes and minor apophyses are the last-documented intrusive event on the Property. This phase cuts the preceding three intrusive phases and is unmineralized although in some areas it is propylitically altered.
7.2.2 VICTORIA ZONE
Mineralization within the Victoria Zone (Figure 7.2) is hosted by a variety of strongly- deformed and locally metamorphosed extrusive and intrusive rocks of undetermined age, principally hornblende diorite which appears to have been intruded by mafic to ultramafic andesitic dykes and later porphyritic diorite bodies similar in texture to the Zafranal diorite in the Zafranal Main Zone. Primary volcanic textures in mafic units suggest that the intrusive rocks were emplaced into a sequence of basaltic andesite volcanic tuffs and/or flows. All of these rock units have been strongly deformed by a regional, dextral shear zone marking the contact with the Coastal Batholith. The strong deformation obscures cross-cutting relationships and primary textures and locally the deformation is so intense that the rocks are mylonitized.
7.2.3 SICERA NORTE ZONE
Sicera Norte is underlain by highly-deformed diorite and quartz-feldspar porphyritic rocks. These intrusive rocks have been thrust over unaltered and unmineralized andesitic volcanic and volcaniclastic rocks along a north-west trending, low-angle reverse fault which marks the southwestern limit of the hydrothermal system (Figure 7.3).
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Figure 7.3 Sicera Norte Zone Geology
7.2.4 SICERA SUR ZONE
Sicera Sur is underlain by a sequence of limestone beds and calcareous sedimentary units belonging to the Upper Jurassic/Lower Cretaceous-age Yura Group. An east-verging reverse thrust has placed older, unmineralized and unaltered, Jurassic Guaneros Formation andesite flows over the Yura Group. This sequence is intruded by at least three generations of dioritic intrusive stocks, plugs and dykes, all with very similar compositions but slightly different textures. The emplacement of these bodies appears to have been controlled by three sets of faults: an older, high-angle northwest-trending system, a later north-trending reverse system and a late, low-angle northwest-trending set of reverse faults. Most of these faults have been reactivated as normal faults (Figure 7.4).
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Figure 7.4 Sicera Sur Zone Geology
Alteration is widespread within all four mineral zones and was used as a primary exploration tool during the initial stage of exploration of the Property. Propylitic (quartz, chlorite, epidote, and calcite) alteration is the most extensive type and phyllic (quartz, sericite, and pyrite) is most-closely associated with mineralization. Potassic alteration is also present in inner portions of alteration zones and is commonly replaced by retrograde phyllic alteration, as evidenced by relict, secondary biotite.
7.3 M INERALIZATION
Copper is present in all four zones in quantities of potential economic significance and occur within veins and stockworks and as disseminations. Primary (hypogene) mineralization, is nearly everywhere overlain by zones of secondary enrichment mineralization that have been categorized on the basis of copper solubility into leached, mixed, oxide, supergene and transitional.
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The emplacement of copper and gold are interpreted by AQM geologists as belonging to separate events; there appears to have been more than one copper- mineralizing event but gold appears to have arrived in a late event that was copper- poor and is associated with quartz veins.
7.3.1 ZAFRANAL MAIN ZONE
The Zafranal Main Zone contains a leach cap as well as oxide and supergene zones that overlie an extensive hypogene zone. The Victoria Zone, which is contiguous with the eastern end of the Zafranal Main Zone, contains mixed, leach, supergene and transitional zones. The Sicera Norte Zone contains minor mixed, supergene and transitional zones above a wedge-shape, steeply-dipping hypogene zone. The Sicera Sur Zone contains the complete set of hypogene, mixed, leach, oxide, supergene and transition zones although, with the exception of the leach zone, all are irregular and discontinuous.
Hypogene mineralization is most commonly associated with potassic alteration (quartz, biotite, chlorite and potassium feldspar) and Microdiorite is the most common host.
Oxide mineralization is associated with intense phylitic (sericite, quartz, clay) alteration and characteristically is located immediately beneath the zone of leaching. Copper oxide minerals include brochantite, chrysocolla, chalcocite and tenorite as well as psilomelane. In the Zafranal Main Zone, copper oxide mineralization commonly contains remnants of secondary copper sulphides.
Supergene mineralization is represented by chalcocite replacing chalcopyrite and as overgrowths on pyrite. Supergene mineralization is best developed in volcanic strata and particularly in the Zafranal Diorite in association with phylitic alteration (sericite, quartz, chlorite/biotite, clay minerals and pyrite) and is interpreted to have formed during Upper Eocene to Lower Miocene time, as evidenced by the lack of structural deformation except that caused by the youngest normal faults.
Continuity of mineralization within the Zafranal Main Zone has been demonstrated by drilling to extend for about 2,500 m along strike with a width between 500 to 800 m, and has been traced over a vertical range of nearly 1,000 m.
7.3.2 VICTORIA ZONE
Mineralization occurs as a tabular body parallel to the main foliation, suggesting that it predates or is contemporaneous with the principal deformation event. Chalcopyrite occurs both as fine-grained disseminations and as patches associated with veins and stockworks. Mineralization within the Victoria Zone has dimensions of about 900 m along strike by up to 500 m width by about 400 m vertical.
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7.3.3 SICERA NORTE ZONE
Mineralization occurs as fine-grained disseminated chalcopyrite that is mainly hosted by phyllically-altered diorite rocks and its distribution appears to be controlled by transverse, east-northeast trending structures that crosscut the main northwest- trending foliation. Mineralization is cut by late, unmineralized diorite dikes. The Sicera Norte Zone has a strike length of about 400 m, width of about 300 m and a vertical extent of about 300 m.
7.3.4 SICERA SUR ZONE
Mineralization is principally hosted by two of the three main diorite intrusive units. Two areas with strong phyllic alteration and hypogene chalcopyrite mineralization have been found, one on the northern and one on the far southern side of Sicera Sur suggesting two distinct foci for the emplacement of the mineralized porphyry bodies. Both mineralized zones remain open. Mineralization at the Sicera Sur Zone has been found over a strike length of about 800 m and both a width and vertical extent of up to 400 m.
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8.0 DEPOSIT TYPES
Mineralization within the Property belongs to the calc-alkalic porphyry copper deposit model. Salient characteristics of this type of mineral deposit are presented below. This description is excerpted from Panteleyev et al. 1995.
DESCRIPTION
This generalized model includes various subtypes all of which contain chalcopyrite in stockwork veinlets in hydrothermally altered porphry and adjacent country rock.
GEOLOGICAL ENVIRONMENT
Rock Types: Tonalite to monzogranite or syenitic porphyry intruding granitic, volcanic, calcareous sedimentary, and other rocks.
Textures: Porphyry has closely spaced phenocrysts and microaplitic quartz- feldspar groundmass.
Age Range: Mainly Mesozoic and Cenozoic, but may be any age.
Depositional Environment: High-level intrusive rocks contemporaneous with abundant dikes, breccia pipes, faults as well as cupolas of batholiths.
Tectonic Setting(s): Rift zones contemporaneous with Andean or island-arc volcanism along convergent plate boundaries. Uplift and erosion to expose subvolcanic rocks.
Associated Deposit Types: Base-metal skarn, epithermal veins, polymetallic replacement, volcanic hosted massive replacement.
DEPOSIT DESCRIPTION
Mineralogy: Chalcopyrite + pyrite ± molybdenite; chalcopyrite + magnetite ± bornite ± Au; assemblages may be superposed. Quartz + K-feldspar + biotite ± anhydrite; quartz + sericite + clay minerals. Late veins of enargite, tetrahedrite, galena, sphalerite and barite in some deposits.
Texture/Structure: Stockwork veinlets and disseminated sulfide grains.
Alteration: From bottom, innermost zones outward: sodic-calcic, potassic, phyllic, and argillic to propylitic. High-alumina alteration in upper part of some
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deposits. Propylitic or phyllic alteration may overprint early potassic assemblage.
Ore Controls: Stockwork veins in porphyry, along porphyry contact, and in favorable country rocks such as carbonate rocks, mafic igneous rocks, and older granitic plutons.
Weathering: Green and blue Cu carbonates and silicates in weathered outcrops, or where leaching is intense, barren outcrops remain after Cu is leached, transported downward, and deposited as secondary sulfides at water table or paleowater table. Fractures in leached outcrops are coated with hematitic limonite having bright red streak. Deposits of secondary sulfides contain chalcocite and other CU2S minerals replacing pyrite and chalcopyrite. Residual soils overlying deposits may contain anomalous amounts of rutile.
Geochemical Signature: Cu + Mo ± Au + Ag + W + B + Sr center, Pb, Zn, Au, As, Sb, Se, Te, Mn, Co, Ba, and Rb outer. Locally Hi and Sn form most distal anomalies. High S in all zones. Some deposits have weak U anomalies.
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9.0 EXPLORATION
9.1 Z AFRANAL M AIN Z ONE
AQM began fieldwork on the Property in June 2009 with a program of rock sampling and geological re-mapping of the Zafranal Main Zone. This work confirmed the earlier Teck Cominco results with respect to the extent of copper-gold mineralization, bedrock geology and alteration.
A magneto-telluric survey (5.7 line-kilometres) was run over the Zafranal Main Zone in November 2009 and the results were combined with those from a similar survey of 5 line-kilometres carried out by Teck Cominco in 2004. The combined surveys defined a zone of low resistivity measuring about 1,000 m by 500 m that is coincident with the known surface indications of porphyry-style alteration and mineralization.
9.2 V ICTORIA Z ONE
No work other than drilling was done specifically within the Victoria Zone as it is immediately on strike with the Zafranal Main Zone and was covered by surface sampling and geophysical programs as part of the assessment of that zone.
9.3 S ICERA N ORTE Z ONE
Surface rock geochemical sampling was carried out prior to the 2010 drill program described in the following section, and geophysical surveys (19.7 km of induced polarization (IP) and 36.4 km of magnetics) were carried out in 2011. These surveys identified structural domains and areas in which sulphides are present and therefore were useful in designing the subsequent drill program.
9.4 S ICERA S UR Z ONE
A surface rock-geochemical sampling program was completed at Sicera Sur in late 2009 and indicated the presence of a zone of surface copper enrichment: of the 223 samples that were collected, primarily along existing exploration access roads, 50% contained more than 300 ppm copper and 14% more than 0.1% copper. These results, together with surface indications of a phyllic alteration zone measuring 3 km by 1.8 km as well as a leached cap, suggested the potential for the presence of significant mineralization at depth.
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The survey work carried out on all four zones was conventional in nature and is not considered by Tetra Tech to have introduced any biases into the results obtained.
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10.0 DRILLING
10.1 Z AFRANAL M AIN Z ONE
As of November 2011, AQM had tested the Zafranal Main Zone with 188 core holes with an aggregate length of 66,126 m and 41 RC holes with an aggregate length of 13,328 m. Of these, 10 core holes and 10 RC holes are located outside the volume for which the following resource estimate was made. AQM incorporated into their database results from four core holes (1,556 aggregate metres) and 32 RC holes (9,625 aggregate metres) that had been drilled by Teck in 2004 and 2005.
Descriptive statistics for assays from both core and RC holes are presented in Table 10.1.
Table 10.1 Zafranal Main Zone Assay Descriptive Statistics
Length Cu Au (m) (%) (g/t) DDH Statistics Mean 2.00 0.06 0.24 Standard Error 0.00 0.00 0.00 Median 2.00 0.04 0.14 Mode 2.00 0.00 0.11 Standard Deviation 0.22 0.12 0.38 Sample Variance 0.05 0.01 0.14 Kurtosis 96.71 1,052.75 1,719.98 Skewness 2.07 23.21 23.39 Range 10.80 8.34 32.80 Minimum 0.20 0.00 0.00 Maximum 11.00 8.34 32.80 Sum 67,677.70 2,120.68 8,007.36 Count 33,860.00 33,860.00 33,860.00 RC Statistics Mean 1.46 0.02 0.19 Standard Error 0.00 0.00 0.00 Median 1.00 0.02 0.08 Mode 1.00 0.00 0.00 Standard Deviation 0.50 0.60 0.31 Sample Variance 0.25 0.36 0.10 Kurtosis -1.97 217.09 39.81 table continues…
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Length Cu Au (m) (%) (g/t) Skewness 0.18 -14.05 4.28 Range 1.00 17.59 7.85 Minimum 1.00 -9.00 0.00 Maximum 2.00 8.59 7.85 Sum 22,760.00 267.50 3008.69 Count 15,636.00 15,636.00 15,636.00
10.2 V ICTORIA Z ONE
During 2011, AQM drilled 31 core holes (10,765 aggregate metres) and seven RC holes (2,103 aggregate metres) to test the Victoria Zone. The assay dataset used for the Victoria resource estimate includes two holes that were drilled by Teck but excludes four holes (three RC and one DDH) that fall outside the geological model of the zone. Descriptive statistics of the assay dataset are presented in Table 10.2.
Table 10.2 Victoria Zone Assay Descriptive Statistics
Length Cu Au (m) (%) (g/t) DDH Statistics Mean 2.00 0.16 0.03 Standard Error 0.00 0.00 0.00 Median 2.00 0.11 0.02 Mode 2.00 0.12 0.00 Standard Deviation 0.24 0.19 0.07 Sample Variance 0.06 0.03 0.01 Kurtosis 10.45 27.68 2,551.34 Skewness 0.56 3.69 43.50 Range 3.55 2.91 4.37 Minimum 0.45 0.00 0.00 Maximum 4.00 2.91 4.37 Sum 10,765.40 861.40 141.35 Count 5,379.00 5,379.00 5,379.00 RC Statistics Mean 1.64 0.06 0.02 Standard Error 0.01 0.00 0.00 Median 2.00 0.03 0.01 Mode 2.00 0.00 0.00 Standard Deviation 0.48 0.10 0.10 Sample Variance 0.23 0.01 0.01 table continues…
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Length Cu Au (m) (%) (g/t) Kurtosis -1.66 68.25 1,490.04 Skewness -0.59 6.71 37.66 Range 1.00 1.68 3.98 Minimum 1.00 0.00 0.00 Maximum 2.00 1.68 3.98 Sum 2,727.00 99.21 27.65 Count 1,662.00 1,662.00 1,662.00
10.3 S ICERA N ORTE Z ONE
During 2010 and 2011, AQM drilled 47 core holes (14,503 aggregate metres) and 11 RC holes (3,232 aggregate metres) to assess the Sicera Norte Zone. Descriptive statistics of the assay dataset are presented in Table 10.3.
Table 10.3 Sicera Norte Zone Assay Descriptive Statistics
Length Cu Au (m) (%) (g/t) DDH Assays Mean 2.02 0.10 0.01 Standard Error 0.00 0.00 0.00 Median 2.00 0.06 0.01 Mode 2.00 0.01 0.00 Standard Deviation 0.26 0.12 0.03 Sample Variance 0.07 0.01 0.00 Kurtosis 8.08 12.70 2,503.66 Skewness 0.92 2.54 44.05 Range 3.60 1.62 1.83 Minimum 0.40 0.00 0.00 Maximum 4.00 1.62 1.83 Sum 14,917.10 742.19 81.64 Count 7,381.00 7,381.00 7,381.00 RC Assays Mean 1.13 0.08 0.03 Standard Error 0.01 0.00 0.00 Median 1.00 0.04 0.01 Mode 1.00 0.01 0.01 Standard Deviation 0.35 0.12 0.07 Sample Variance 0.12 0.01 0.01 Kurtosis 29.79 19.97 483.51 table continues…
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Length Cu Au (m) (%) (g/t) Skewness 3.59 3.33 17.73 Range 6.00 1.62 2.51 Minimum 1.00 0.00 0.00 Maximum 7.00 1.62 2.51 Sum 3,219.00 237.92 74.05 Count 2,861.00 2,861.00 2,861.00
10.4 S ICERA S UR Z ONE
In 2011, AQM tested the Sicera Sur Zone with three core holes (1,018 aggregate metres) and 34 RC holes (11,456 aggregate metres). Descriptive statistics of the assay dataset are presented in Table 10.4.
Table 10.4 Sicera Sur Zone Assay Descriptive Statistics
Length Cu Au (m) (%) (g/t) DDH Assays Mean 2.01 0.15 0.01 Standard Error 0.26 0.01 0.00 Median 2.00 0.07 0.01 Mode 2.00 0.07 0.00 Standard Deviation 5.50 0.16 0.02 Sample Variance 30.28 0.02 0.00 Kurtosis 447.27 0.76 20.79 Skewness 21.13 1.22 3.61 Range 117.50 0.70 0.17 Minimum 1.00 0.00 0.00 Maximum 3.50 0.70 0.17 Sum 899.30 67.03 5.81 Count 448.00 448.00 448.00 RC Assays Mean 1.08 0.11 0.01 Standard Error 0.00 0.00 0.00 Median 1.00 0.07 0.01 Mode 1.00 0.02 0.00 Standard Deviation 0.28 0.16 0.04 Sample Variance 0.08 0.03 0.00 Kurtosis 7.11 83.84 3,718.09 Skewness 3.02 6.64 50.00 table continues…
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Length Cu Au (m) (%) (g/t) Range 1.00 3.38 2.86 Minimum 1.00 0.00 0.00 Maximum 2.00 3.38 2.86 Sum 11,456.00 1,194.44 124.20 Count 10,575.00 10,575.00 10,575.00
The drillholes intersected mineralization at a range of angles relative to true thickness; therefore it is not possible to generalize with respect to the relationship between true and intersected thickness.
Drill plans and representative sections through the four mineral zones of interest are presented in Section 14.0.
The collar location and elevation of all drillholes were surveyed by professional surveyors and locations are referenced to the PSAD56 datum and later converted to WGS84 datum. Downhole surveys were run with a variety of instruments, both optical and gyroscopic. Deviflex and Gyroscope tools are non-magnetic and collected readings from the bottom of the drill hole to surface at two-second intervals. The FlexIT instrument is magnetic and was used to take readings at 50 m intervals down the hole.
Core was transported by AQM personnel from the drill site to camp several times each day. Core was photographed then logged for recovery and rock quality designation (RQD) prior to being logged for lithology, alteration, mineralization and structure. The core was then marked for sampling.
Tetra Tech does not believe that there are any drilling, sampling or recovery factors that could materially affect the accuracy and reliability of the results. Because both core and RC drilling were used, the possibility exists that the drilling methodologies resulted in real differences between the assays for the two populations, particularly as the average gold and copper grades of the RC sample population are, in all four zones, lower than those of the drill core samples. The existence of meaningful differences between the two datasets was assessed using a T-test and the assumption that the variances of the populations were the same. In all four cases the differences between means were indicated to be not significant. On the basis of this result, the RC and drill core sample populations were aggregated.
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11.0 SAMPLE PREPARATION, ANALYSES AND SECURITY
11.1 R EVERSE C IRCULATION S AMPLING
RC chips were collected during drilling in 1 m intervals. Samples were weighed then put though a sample splitter and a one-third split (approximately 7 kg on average) was placed into a laboratory mesh sack and the remainder of the sample was placed into a rice bag.
Samples were then transported to camp where blanks of granodiorite (five per 100 samples), duplicates (five per 100 samples) and six standards per 100 samples (two each of three different standards), were added to the sample stream at random intervals.
Samples were shipped to the ALS Chemex laboratory in Arequipa for crushing and splitting. Shipping took place in entire-hole lots that were transported from the Property to Arequipa by an AQM representative.
Following preparation in Arequipa, pulps were sent to the ALS Chemex laboratory in Lima for analyses.
11.2 D RILL C ORE S AMPLING
Core samples were transported to the camp where, as detailed in the previous section, blanks, duplicates and standards were added to the sample stream at random intervals. Drill core was then sampled at the core processing facility located in the camp. Core was sawn in half and one half was placed in a numbered plastic sample bag and the other half was returned to the core box for archival purposes. Most samples were collected in two-metre lengths. Individual sample bags were collected into rice bags that were then sealed with numbered “zap straps”. Samples were transported to the ALS Chemex laboratory in Arequipa for processing and subsequently to Lima for analysis.
The ALS Chemex laboratory in Lima is International Organization for Standardization (ISO) 17025 certified. All samples were analyzed for 32 elements by inductively coupled plasma (ICP) (MEICP41). Gold was analyzed by atomic absorption (AA) (AA23) using a 30 g sample.
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Samples with ICP copper assays greater than 0.2% were subjected to a series of analyses to establish sequential solubility. The values obtained from these analyses were used to model the mineral zones that form the basis of the following resource estimate.
Total copper content was determined by aqua regia digestion followed by an atomic absorption spectrophotometer (AAS) (AA46 for mineralized material-grade copper between 0.01% and 40% copper). Samples were then dissolved by weak sulphuric acid (AA06) to establish the proportion of total copper belonging to the secondary “oxide” phase. Sulphuric acid digestion was followed by cyanide digestion (AA16s) to determine the remaining portion of the total copper belonging to other secondary phases. The cyanide digestion was followed by hydrochloric and nitric acid digestion (AA62S) to determine the residual copper content which is equated to primary or hypogene mineralization. The rationale for the assignment of copper to various secondary mineral zones is explained in Section 14.0.
11.3 Q UALITY A SSURANCE/QUALITY C ONTROL P ROCEDURES
AQM employed standards, blanks and duplicate samples in their quality assurance/ quality control (QA/QC) program. Table 11.1 enumerates the number of control samples used for each zone and the percentage of the sample population represented by each type of control sample. The responses obtained from the control program are discussed below.
Table 11.1 Zafranal Property Control Samples
Standards Blanks Duplicates Total Zone Samples Standards Blanks Duplicates (%) (%) (%) (%) Zafranal Main 33,860 2,889 2,673 2,658 8.53 7.89 7.85 24.28 Victoria 7,041 439 378 379 6.23 5.37 5.38 16.99 Sicera Norte 10,242 699 593 588 6.82 5.79 5.74 18.36 Sicera Sur 11,023 519 445 457 4.71 4.04 4.15 12.89 Total/Average 62,166 4,546 4,089 4,082 7.31 6.58 6.57 20.46
11.4 S TANDARDS
AQM had standards prepared by SGS in Lima from coarse drill core sample rejects. The samples were prepared in a conventional manner and were subjected to independent round-robin analyses, the results of which were used to determine the expected mean and standard deviation of each standard. Table 11.2 is a list of the standards used to date, mean, plus/minus two standard deviation limits as well as the number of each that have been used.
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Table 11.2 Zafranal Property Standards
No. Standard Mean X-2SD X+2SD Used 101 1.047 0.983 1.111 251 111 1.500 1.390 1.610 45 201 0.620 0.589 0.650 291 222 2.260 2.140 2.380 48 301 0.306 0.287 0.324 279 333 6.870 6.710 7.030 44 444 1.290 1.210 1.370 345 555 0.639 0.586 0.692 344 666 0.313 0.291 0.335 308 700 0.347 0.366 0.328 707 777 0.344 0.294 0.394 45 800 0.666 0.678 0.655 622 888 0.742 0.672 0.812 41 900 1.166 1.189 1.143 540 999 1.550 1.350 1.750 35 1000 0.257 0.235 0.279 211 2000 0.529 0.494 0.565 204 3000 0.865 0.821 0.907 186 Total 4,546
Analyses of standards were reviewed to determine the number that exceeded the expected mean plus/minus two standard deviations. Those that exceeded those limits are listed in Table 11.3. It is probable that some of these outliers are due to the errors in documentation of standards at the time they were entered into the sample stream i.e. they were assigned the wrong standard identification. Also, the high failure rate of Standard 333 (6 out of 44 samples) suggests the possibility that the standard is inhomogeneous. However, it is notable that 29 of the 43 analyses that exceeded the two standard deviation limits occur as pairs or higher multiples from the same analytical batch. This may indicate that the fault lies with the analytical procedure. AQM did not request that any of the samples be re-analyzed so that it is not possible to determine the cause.
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Table 11.3 Zafranal Property Over-Limit Standards
Cu Standard Two Standard Absolute Batch No. Standard Location Hole ID Sample (%) Mean Deviations Difference AR10025706 101 Victoria ZFRC10-058 162434 0.319 1.047 0.128 0.728 AR10033503 333 Zafranal ZFDDH10-025 158159 7.240 6.870 0.320 0.370 AR10034751 333 Zafranal ZFDDH10-034 165478 7.210 6.870 0.320 0.340 AR10035337 333 Zafranal ZFDDH10-027 166614 7.430 6.870 0.320 0.560 AR10035337 333 Zafranal ZFDDH10-027 166752 7.350 6.870 0.320 0.480 AR10040822 333 Zafranal ZFRC10-066 170250 7.200 6.870 0.320 0.330 AR10040822 333 Zafranal ZFRC10-066 170314 7.220 6.870 0.320 0.350 AR10073212 555 Sicera Norte SNRC10-003 203175 0.101 0.639 0.106 0.538 AR10073213 555 Sicera Norte SNRC10-004 203435 0.024 0.639 0.106 0.615 AR10086588 666 Zafranal ZFDDH10-077 176418 0.624 0.313 0.106 0.311 AR10086588 555 Zafranal ZFDDH10-077 176433 0.312 0.639 0.106 0.327 AR10110580 666 Zafranal ZFDDH10-097 185883 1.510 0.313 0.106 1.197 AR10110580 555 Zafranal ZFDDH10-097 185965 1.500 0.639 0.106 0.861 AR10110580 555 Zafranal ZFDDH10-097 187023 0.753 0.639 0.106 0.114 AR10125538 900 Zafranal ZFDDH10-102 182407 1.120 1.166 0.046 0.046 AR10127969 900 Zafranal ZFDDH10-114 187980 1.215 1.166 0.046 0.049 AR10133802 900 Zafranal ZFDDH10-122 190239 1.120 1.166 0.046 0.046 AR10142905 900 Zafranal ZFDDH10-125 192568 1.075 1.166 0.046 0.091 AR10142905 900 Zafranal ZFDDH10-125 192810 1.000 1.166 0.046 0.166 AR10147935 900 Zafranal ZFDDH10-136 205803 1.120 1.166 0.046 0.046 AR11096696 900 Sicera Norte SNDDH11-025 216303 1.115 1.166 0.046 0.051 AR11096697 900 Zafranal ZFDDH11-157 218294 1.120 1.166 0.046 0.046 AR11107062 900 Sicera Norte SNDDH11-029 221694 1.115 1.166 0.046 0.051 AR11107062 900 Sicera Norte SNDDH11-029 221734 1.115 1.166 0.046 0.051 table continues…
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Cu Standard Two Standard Absolute Batch No. Standard Location Hole ID Sample (%) Mean Deviations Difference AR11107063 900 Zafranal ZFDDH11-167 225494 1.120 1.166 0.046 0.046 AR11117323 900 Sicera Norte SNRC11-011 231421 1.220 1.166 0.046 0.054 AR11118573 900 Sicera Norte SNRC11-014 232234 1.115 1.166 0.046 0.051 AR11122810 900 Sicera Norte SNRC11-013 231973 1.100 1.166 0.046 0.066 AR11122810 900 Sicera Norte SNRC11-013 231993 1.105 1.166 0.046 0.061 AR11125030 900 Sicera Norte SNRC11-015 232713 1.120 1.166 0.046 0.046 AR11129154 700 Zafranal ZFDDH11-177 226334 1.170 0.347 0.106 0.823 AR11129154 700 Zafranal ZFDDH11-177 226361 1.180 0.347 0.106 0.833 AR11160762 900 Zafranal ZFDDH11-191 227203 1.120 1.166 0.046 0.046 AR11160762 900 Zafranal ZFDDH11-191 227334 1.090 1.166 0.046 0.076 OCT0014.R11 3000 Victoria ZFDDH11-212 246303 0.777 0.865 0.086 0.088 OCT0014.R11 3000 Victoria ZFDDH11-212 246343 0.752 0.865 0.086 0.113 OCT0014.R11 3000 Victoria ZFDDH11-212 246543 0.765 0.865 0.086 0.100 SEP0348.R11 3000 Zafranal ZFDDH11-208 227994 0.777 0.865 0.086 0.088 SEP0348.R11 3000 Zafranal ZFDDH11-208 252034 0.768 0.865 0.086 0.097 SEP0348.R11 3000 Zafranal ZFDDH11-208 252061 0.767 0.865 0.086 0.098 SEP0348.R11 3000 Zafranal ZFDDH11-208 252194 0.768 0.865 0.086 0.097 SEP0402.R11 3000 Victoria ZFDDH11-210 244561 0.767 0.865 0.086 0.098 SEP0402.R11 3000 Victoria ZFDDH11-210 244603 0.773 0.865 0.086 0.092 SEP0402.R11 3000 Victoria ZFDDH11-210 244694 0.766 0.865 0.086 0.099
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11.5 B LANKS
AQM used locally-obtained, unaltered granodiorite as blank material on the assumption that it is unmineralized. The average copper content for all blanks (4,088) is 0.003% and the standard deviation is also 0.003%. Therefore the anomalous threshold was taken as ±0.009% (Table 11.4). Table 11.5 is a list of those samples that exceeded this limit. As with the standards, a large percentage of the samples (55%) occur in pairs or higher multiples from the same analytical batch suggesting the possibility of cross-sample contamination or analytical error. However, given the fact that the source material may contain trace quantities of copper it cannot be determined with certainty whether the over-limit values arose because of contamination or analytical error. AQM did not take any remedial action on the basis of over-limit blanks.
Table 11.4 Zafranal Property Blank Descriptive Statistics
Blanks: Sicera Sicera Copper % Zafranal Victoria Norte Sur Average Mean 0.0024 0.002 0.002 0.004 0.003 Standard Error 0.0000 0.000 0.000 0.001 - Median 0.0019 0.002 0.002 0.003 - Mode 0.0014 0.002 0.002 0.002 - Standard Deviation 0.0021 0.003 0.001 0.013 0.003 Sample Variance 0.0000 0.000 0.000 0.000 - Kurtosis 272.7601 117.158 53.267 416.082 - Skewness 11.7419 9.945 5.175 20.105 - Range 0.0623 0.038 0.017 0.263 - Minimum 0.0005 0.000 0.001 0.001 - Maximum 0.0628 0.039 0.018 0.264 - Sum 6.4972 0.766 1.355 1.659 - Count 2673 377 593 445 4,088
Table 11.5 Zafranal Property Over-limit Blanks
Cu Batch Hole ID Location Sample (%) AR09146061 ZFRC09-033 Zafranal 150184 0.009 AR09146061 ZFRC09-033 Zafranal 150253 0.011 AR10000401 ZFRC09-034 Zafranal 151004 0.010 AR10000401 ZFRC09-034 Zafranal 151008 0.010 AR10003799 ZFRC10-039 Zafranal 152476 0.011 AR10003799 ZFRC10-039 Zafranal 152507 0.010 AR10003799 ZFRC10-039 Zafranal 152532 0.011 table continues…
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Cu Batch Hole ID Location Sample (%) AR10004119 ZFRC10-042 Zafranal 153576 0.013 AR10004992 ZFRC10-041 Zafranal 153176 0.012 AR10005620 ZFRC10-043 Zafranal 154454 0.009 AR10006266 ZFDDH10-008 Zafranal 154253 0.012 AR10006267 ZFRC10-045 Zafranal 155137 0.012 AR10006267 ZFRC10-045 Zafranal 155141 0.019 AR10006268 ZFRC10-044 Zafranal 154866 0.011 AR10008915 ZFRC10-050 Zafranal 159878 0.011 AR10008915 ZFRC10-050 Zafranal 159882 0.010 AR10008915 ZFRC10-050 Zafranal 159907 0.010 AR10008915 ZFRC10-050 Zafranal 159978 0.010 AR10010814 ZFRC10-047 Zafranal 155782 0.012 AR10011319 ZFRC10-048 Zafranal 159047 0.009 AR10011319 ZFRC10-048 Zafranal 159116 0.011 AR10017022 ZFRC10-053 Zafranal 161023 0.024 AR10017022 ZFRC10-053 Zafranal 161107 0.012 AR10019220 ZFDDH10-013 Zafranal 156925 0.036 AR10019224 ZFRC10-054 Zafranal 161380 0.013 AR10023984 ZFDDH10-017 Zafranal 168409 0.009 AR10027330 ZFRC10-060 Zafranal 162725 0.008 AR10027335 ZFDDH10-018 Zafranal 166109 0.009 AR10027335 ZFDDH10-018 Zafranal 166125 0.011 AR10027335 ZFDDH10-018 Zafranal 166184 0.010 AR10028952 ZFDDH10-019 Zafranal 168784 0.015 AR10028952 ZFDDH10-019 Zafranal 168825 0.015 AR10028956 ZFRC10-062 Zafranal 163253 0.010 AR10031730 ZFRC10-064 Zafranal 163851 0.010 AR10031730 ZFRC10-064 Zafranal 163882 0.012 AR10042849 ZFDDH10-030 Zafranal 158497 0.010 AR10047312 ZFDDH10-042 Zafranal 169921 0.010 AR10050837 ZFDDH10-043 Zafranal 158668 0.013 AR10064892 SNRC10-002 Sicera Norte 179784 0.018 AR10066018 SSRC10-009 Sicera Sur 202408 0.264 AR10098444 ZFDDH10-088 Zafranal 188418 0.010 AR10151585 ZFDDH10-134 Zafranal 190745 0.063 AR11078859 SNDDH11-013 Sicera Norte 215609 0.012 AR11096698 SSRC11-013 Sicera Sur 228016 0.036 AR11117381 ZFDDH11-171 Victoria 217209 0.039 AR11156831 ZFDDH11-189 Zafranal 213909 0.009 SEP0402.R11 ZFDDH11-210 Victoria 244553 0.019 SEP0402.R11 ZFDDH11-210 Victoria 244625 0.030
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11.6 D UPLICATES
Copper assays from duplicate samples show a very high degree of correlation for all four mineral zones. Descriptive statistics for duplicate samples are shown below in Table 11.6.
Table 11.6 Zafranal Property Duplicate Sample Descriptive Statistics
Copper Duplicates (%) Zafranal Victoria Sicera Norte Sicera Sur Statistic Original Duplicate Original Duplicate Original Duplicate Original Duplicate Mean 0.221 0.224 0.14 0.14 0.100 0.101 0.111 0.111 Standard 0.006 0.006 0.01 0.01 0.005 0.005 0.006 0.006 Error Median 0.126 0.126 0.09 0.08 0.060 0.059 0.073 0.074 Mode 0.131 0.111 0.17 0.14 0.064 0.104 0.112 0.119 Standard 0.325 0.335 0.19 0.20 0.111 0.118 0.131 0.129 Deviation Sample 0.106 0.112 0.04 0.04 0.012 0.014 0.017 0.017 Variance Kurtosis 102.205 78.471 51.11 84.11 5.927 9.894 32.658 28.217 Skewness 6.571 6.070 5.16 6.91 2.056 2.537 4.210 3.882 Range 7.471 6.980 2.37 2.79 0.791 0.973 1.505 1.445 Minimum 0.000 0.000 0.00 0.00 0.000 0.000 0.000 0.000 Maximum 7.470 6.980 2.37 2.79 0.791 0.973 1.505 1.445 Sum 587.931 594.935 54.86 53.81 58.671 59.353 50.744 50.927 Count 2658 2658 379 379 588 588 457 457 Correlation - 0.97 - 0.98 - 0.94 - 0.99
The difference between means of only 40 pairs of duplicates exceeded the mean value by two standard deviations and it is probable that the difference is attributable to natural variance in mineral distribution between samples. No samples were re- assayed on the basis of differences between duplicates.
Tetra Tech considers that the sample preparation, security and analytical procedures meet industry standards.
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12.0 DATA VERIFICATION
Tetra Tech inspected the Zafranal Main and Sicera Norte and Sicera Sur Zones. The Victoria Zone was not examined. Drillhole and surface sample locations were observed as well as surface manifestations of copper mineralization, lithology and alteration. The core-logging and processing procedures were reviewed at the core processing facility located in the main camp. Drill core is also stored at this location and core from several holes that demonstrated the principal rock types and styles of alteration and mineralization were examined.
Data capture and processing procedures were also observed in the field geology office where an explanation was also provided, by the geologists involved, of the geological interpretation of the four mineral zones.
Copies of laboratory certificates (in .pdf format) together with the corresponding Microsoft Excel® spreadsheets were obtained for five separate shipments of (1,003) samples submitted by AQM to ALS Chemex. The assay values for copper and gold on the certificates were compared to the corresponding values on the spreadsheets as well as against the values contained in the AQM database. No discrepancies were found.
Tetra Tech did not collect any verification samples for three reasons: 1) to date over 62,000 samples from the Property have been assayed for copper and gold, all by a highly reputable commercial laboratory; 2) chalcopyrite and secondary copper minerals are clearly visible in core and outcrop so that the presence of copper is not in question; 3) informal miners are or have been mining gold within the Property for an extended period and, in Sicera Norte are actively mining within the area that has been tested by drilling. This was considered sufficient evidence of the presence of gold.
Tetra Tech considers the data to be of adequate quality for the purposes of the resource estimate that follows.
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13.0 M INERAL P ROCESSING AND M ETALLURGICAL T ESTING
13.1 I NTRODUCTION
In June 2012, Transmin prepared a comprehensive report encompassing the most recent metallurgical test programs which had been developed and monitored by Transmin. The report was provided to Tetra Tech and forms the basis of this section. The testing occurred concurrently at Plenge Laboratories (Plenge) in Lima, Peru and G&T Metallurgical Services Ltd. (G&T) in Kamloops, BC, Canada. Both laboratories are accredited metallurgical facilities. Transmin’s report also discusses test work from 2010 with special reference to the grinding programs that were completed at that time and which will be used as a basis for the plant design in this study.
13.2 C HRONOLOGY OF M ETALLURGICAL T EST W ORK
Earlier metallurgical test work programs are listed in Table 13.1. The June 2012 report reviewed some of the earlier test work; however not all of the earlier test work was included in the Transmin report. There have been three phases of test work and reporting; the latest is Phase III, which forms the basis for the majority of this analysis.
Phases I and II were carried out under supervision by AMEC Minproc, Amdel Mineral Laboratories (Amdel), JKTech, SGS (in Santiago, Chile) and Philips Enterprises. Phase I included comminution testing, leaching, mineralogical identification, flotation optimization of three composites and some variability flotation testing. A final report incorporating all the results was written by Amdel at the end of the Phase I test work.
The Phase II test program included locked cycle flotation testing, further comminution testing and leach tests. A report of the flotation test work was issued in October 2011 with inconsistencies and errors, and was re-issued in June 2012. The test work program was poorly monitored and reported with unclear objectives. The write-up was unclear and data was incorrectly reported rendering the test results unusable. Of all the Phase I and II test work which was completed, only the comminution testing was used for the final process design.
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Table 13.1 Test Work Programs and Reports
Author or Ref. Document or Test Program Laboratory Date No. Mineralized material characterization study Teck 2005 1 Phase I – flotation optimization, flotation variability testing, Amdel May to August 2 leaching, mineralogical identification, sizing analysis 2010 SMC test on 15 samples JK Tech April 2010 3 Abrasion tests SGS 2010 4 Impact tests Philips May 2010 5 Enterprises Phase II – comminution tests, locked cycle and variability Amdel March to 6 flotation tests, leach amenability tests December 2011 Bond, abrasion and SAG Mill Comminution tests SGS January 2011 7 Comminution test update AMEC Minproc August 2010 8 SAG Mill Comminution test report on 42 samples from the JK Tech January 2011 9 Project Zafranal Copper Project – Technical Report December AMEC Minproc February 2011 10 2010 Resource Estimate Document No. 60246-00000-23-002-001 Zafranal Copper Project – Stage 2 Flotation Testwork AMEC Minproc May 2012 11 Document No. 60246-00000-21-002-003 Zafranal Copper Project – Dewatered Tailings Testwork AMEC Minproc May 2012 12 Document No. 60246-00000-21-002-004 Zafranal Copper Project – Leaching Test Work Document AMEC Minproc, May 2012 13 No. 60246-00000-21-002-005 Amdel Phase III test G&T November 2011 14 Plenge to June 2012 TM 621 Zafranal, Metallurgical Test Work Report for AQM Transmin June 2012 15 Report on Phase III Test Work
Section 13.0 will build on the results and premises of the results obtained and explain the parameters chosen for the process design as required.
13.3 M INERALOGY
A mineralized material characterization study was performed by Teck in 2005 and reported in a resource estimate. The following mineralogical details were documented within that report:
• The chip samples were dominated by quartz, plagioclase feldspar and muscovite/sericite. Biotite, chlorite, kaolinite, iron-titanium oxides and apatite occurred in minor to trace amounts. Sulphide minerals consisted mainly of pyrite, with lesser amounts of chalcopyrite, chalcocite and covellite.
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• Secondary copper minerals, such as chalcocite and covellite, hosted 78% of the total copper in the samples studied. A calculated 94.5% of these minerals exhibited surface exposures of 10% or greater, suggesting good heap leaching potential at a relatively coarse particle size. • Copper distribution by mineral species showed a predominance of secondary copper minerals near the surface of the deposit and an increasing amount of chalcopyrite downhole.
• Chalcocite and chalcopyrite were similar in grain size with a mean P80 size value of 38 μm and 39 μm, respectively. Covellite was found to be finer- grained with a mean P80 of 17 μm. • No deleterious minerals were identified at the detection limit of the study. • No carbonate minerals were identified, indicating that tailings and waste rock may be acid generating in the presence of water. • No gold-related mineralogy was reported.
G&T performed mineralogical analysis on the three composite samples which were received for testing in Phase III. G&T conducted a mineralogical evaluation to identify the percent liberation by size and distribution by association class, metal content and mineral content. Table 13.2 provides a summary of the findings as presented in July 2012 by G&T.
Table 13.2 Mineral Composition of Phase III Composite Samples
Mineral Content Percent Mineral ZFC-MC01 ZFC-MC02 ZFC-MC03
Chalcopyrite 1.3 0.6 0.7 Bornite <0.1 <0.1 <0.1 Chalcocite <0.1 0.3 0.1 Covellite <0.1 0.1 0.1 Pyrite 4.5 4.5 4.8 Non-sulphide Gangue 94.2 94.4 94.3
The copper oxide mineral identified was turquoise, a hydrous aluminium and copper phosphate.
Mineral distribution by class of association at a particular grind size was also presented for the three composite samples. The mineral liberation in two dimensions (2D) is shown in Table 13.3.
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Table 13.3 Mineral Liberation of Phase III Composite Samples
Mineral Distribution by Class of Association (%) Sample ID Mineral Status Cp Ch Py Gn ZFC-MC01 Liberated 59.2 25.3 78.4 98.5 P80 = 146 µm Binary - Cp - 46.2 5.8 0.8 Binary - Ch 1.3 - 0.0 0.0 Binary - Py 6.4 2.2 - 0.7 Binary - Gn 31.7 20.8 15.5 - Multiphase 1.4 5.6 0.3 0.0 Total 100.0 100.0 100.0 100.0 ZFC-MC02 Liberated 44.7 51.1 77.2 98.6 P80 = 138 µm Binary - Cp - 14.5 4.7 0.5 Binary - Ch 6.2 - 0.5 0.2 Binary - Py 12.9 0.5 - 0.7 Binary - Gn 34.1 29.7 16.1 - Multiphase 2.2 4.2 1.5 0.1 Total 100.0 100.0 100.0 100.0 ZFC-MC03 Liberated 43.7 33.0 67.9 98.7 P80 = 149 µm Binary - Cp - 24.3 10.9 0.4 Binary - Ch 5.6 - 0.2 0.0 Binary - Py 17.6 6.5 - 0.7 Binary - Gn 28.8 23.1 17.2 - Multiphase 4.2 13.2 3.8 0.1 Total 100.0 100.0 100.0 100.0 Notes: Cp = chalocopyrite, Ch = chalcocite, Py = pyrite, Gn = gangue Numbers are rounded however the totals equal 100.0.
Chalcopyrite appears to be reasonably well-liberated for flotation purposes at the indicated grind sizes for composite sample ZFC-MC01 (about 60%), although the extent of liberation is borderline for samples ZFC-MC02 and ZFC-MC03 (about 45%).
The data also indicates that the pyrite is fairly well-liberated from the copper minerals in composite ZFC-MC01, slightly less in composite sample ZFC-MC02, while composite ZFC-MC03 indicates that there is a higher extent of association of copper with the pyrite of about 11% at this grind size. There is no indication as to where or how the gold is mineralogically associated.
The primary grind P80 of 150 µm was selected for purposes of the design of the comminution section.
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13.4 F EED G RADE
For metallurgical testing, a number of the variability samples were used as a feed source by Plenge. Three composite samples were prepared with a portion of each sent to G&T for their test program. The composite samples which were prepared for testing were as follows:
• ZFC-MC01 – hypogene material • ZFC-MC02 – supergene material • ZFC-MC03 – mixed material (sulphide with oxide material)
The assays of the samples tested, together with the head values used in the design, are provided in Table 13.4
Table 13.4 Head Values – Composite Samples
Assays Cu CuSS CuCN CuRes Ag Au Fe Laboratory Composite (%) (%) (%) (%) (g/t) (g/t) (%) Plenge ZFC-MC01 0.45 0.020 0.070 0.35 1.20 0.15 3.70 ZFC-MC02 0.49 0.070 0.26 0.15 0.60 0.13 3.60 ZFC-MC03 0.41 0.090 0.12 0.19 0.60 0.13 4.10 G&T ZFC-MC01 0.46 0.010 0.020 0.42 <1 0.084 3.50 ZFC-MC02 0.52 0.030 0.25 0.24 <1 0.069 3.40 ZFC-MC03 0.44 0.050 0.12 0.27 <1 - 3.80 Design Mill Feed 0.54 to 0.84 - - - - 0.10 to 0.13 -
Samples were prepared at Plenge for testing locally and sub-samples were shipped to G&T for the parallel variability test program. Chemical characterization was performed at both facilities with slight differences in the results obtained for the actual head grade values for each sample. For example the copper feed ranges for the samples were 0.20 to 0.99% copper for Plenge and 0.22 to 0.88% copper for G&T as shown in Table 13.5. Note that the average values from both Plenge and G&T are lower than the nominal design feed grade value of 0.54% copper.
Table 13.5 Head Values – Variability Samples
Assays Composite Cu Au Fe S Laboratory Identification (%) (g/t) (%) (%) Plenge Variability Samples Minimum 0.20 0.02 2.25 0.09 (35 samples tested) Maximum 0.99 0.37 7.10 7.64 Average 0.44 0.13 4.04 2.52 table continues…
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Assays Composite Cu Au Fe S Laboratory Identification (%) (g/t) (%) (%) G&T Variability Minimum 0.22 0.03 2.00 0.08 Samples (30 samples tested) Maximum 0.88 0.26 8.50 8.73 Average 0.49 0.11 2.73 2.73
The gold feed assay value range for Plenge and G&T are similar, and the average values are similar to the design feed grade value in the process design criteria (PDC).
Also of note is that the sequential copper classification assay values also differ for the two laboratories. All the assay results are available in the master documents.
13.5 M INERALIZED M ATERIAL C HARACTERISTICS
Specific gravity (SG) determinations were conducted for samples in both Phase I and II test programs. The SG values obtained ranged from 2.29 to 2.81 with an average SG value of 2.58, and a 75th percentile value of 2.65. SG values of the samples grouped by copper rock type classification are shown in Table 13.6.
Table 13.6 Specific Gravity Results by Rock Type Classification
Rock Type Classification Statistical Outcome Hypogene Supergene Oxide Minimum 2.36 2.36 2.29 Maximum 2.73 2.81 2.57 Average 2.60 2.59 2.47 75th Percentile 2.67 2.63 2.54 No. of Samples Tested 36 25 4
The average value of the non-oxide rock types is 2.59 with a corresponding average 75th percentile SG of 2.65.
An SG value of 2.70 was given in the Transmin report and used in the PDC.
The Phase III test program was made up of samples representing two zones, namely Zafranal Main and Victoria Zones. These samples had copper head grades of higher than 0.2% copper with similar geological parameters and were samples taken from the applicable diamond drill cores.
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13.6 G RINDABILITY
Variability samples for comminution tests were sourced from diamond drill core material and tested at the SGS laboratories in Santiago, Chile. Phase I testing included 21 samples and the Phase II program tested 51 samples. Samples were classified by copper rock type namely, supergene, hypogene and oxide. Comminution tests included crusher work index (CWi), abrasion index (Ai), Bond ball mill work index (BWi) and SAG Mill Comminution (SMC) tests.
The detailed results have been documented in the original reports and were summarized in the Transmin report which also records the sample location. The results from these comminution tests have been used as a basis for the design of the crushing and grinding circuit.
13.6.1 CRUSHER WORK INDEX
The CWi was only determined on the Phase I samples. Table 13.7 lists the results obtained. Relevant graphs of the data are documented in the Transmin report.
Table 13.7 Crusher Bond Work Index Results – Phase I
Sample Rock Type CWi (Average) ID Classification (kWh/t) SGS-01 Supergene 3.15 SGS-02 Supergene 3.73 SGS-03 Hypogene 4.27 SGS-04 Supergene 5.31 SGS-05 Supergene 7.43 SGS-06 Supergene 5.71 SGS-07 Supergene 8.07 SGS-08 Supergene 6.54 SGS-09 Supergene 8.05 SGS-10 Hypogene 9.51 SGS-11 Hypogene 8.31 SGS-12 Supergene 7.65 SGS-13A Oxide 5.60 SGS-13B Oxide 4.47 SGS-14 Hypogene 6.44 Average - 6.28 75th Percentile - 7.85
The 75th percentile CWi value for the samples tested is 7.85 kWh/t which is the value used in the design and shown in the PDC.
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A further breakdown of results into copper rock type classification yields the results as shown in Table 13.8.
Table 13.8 Crusher Bond Work Index Results by Rock Type Classification
Rock Type Classification Statistical Outcome (kWh/t) Hypogene Supergene Oxide Minimum 4.27 3.15 4.47 Maximum 9.51 8.07 5.60 Average 7.13 6.18 5.04 75th Percentile 8.61 7.65 5.32 No. of Samples Tested 4 9 2
From a crushing perspective, the hypogene material is classified as hard, the supergene as moderately hard, and the oxide as moderately soft. The average value of the two hardest mineral types is 6.67 kWh/t with a corresponding 75th percentile hardness of 8.13 kWh/t, which is slightly higher than the value used in the PDC.
13.6.2 ABRASION INDEX
The Ai values were determined for samples in both Phase I and II test programs. The Ai values obtained ranged from 0.014 to 0.630 g with an average of 0.213 g. The 75th percentile value was 0.279 g which is used in the design and shown in the PDC. AMEC Minproc hypothesized that the wide distribution obtained for the abrasion results appear to indicate that factors other than mineral type contribute to the abrasion properties of the materials. The factors considered included sample depth below ground level, lithology, and mineralogical alteration characteristics.
The Ai values of the samples grouped by copper rock type classification are shown in Table 13.9.
Table 13.9 Abrasion Index Results by Rock Type Classification
Rock Type Classification Statistical Outcome (g) Hypogene Supergene Oxide Minimum 0.034 0.014 0.015 Maximum 0.598 0.630 0.172 Average 0.243 0.198 0.098 75th Percentile 0.330 0.257 0.137 No. of Samples Tested 38 28 6
The hypogene material is the most abrasive from a crushing and grinding perspective, although the most abrasive sample tested was that from the supergene mineralization, namely 0.630 g.
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The average value of the two most abrasive rock types average values is 0.221 g with a corresponding 75th percentile Ai of 0.294 g.
The detailed results are presented in the Transmin report.
13.6.3 BOND BALL MILL WORK INDEX
The BWi was conducted on both Phase I and II samples and generated a total of 72 results. The BWi values obtained ranged from 5.3 to 15.8 kWh/t, with an arithmetic average of 11.2 kWh/t and a 75th percentile value of 12.2 kWh/t. The 75th percentile value of 12.2 kWh/t is used in the design.
The BWi values obtained for the samples as grouped by copper rock type classification are shown in Table 13.10.
Table 13.10 Bond Ball Mill Index Results by Rock Type Classification
Rock Type Classification Statistical Outcome (kWh/t) Hypogene Supergene Oxide Minimum 7.04 5.30 8.39 Maximum 14.24 15.82 10.49 Average 11.38 11.28 9.31 75th Percentile 12.56 12.05 9.46 No. of Samples Tested 38 28 6
The hypogene and supergene materials were determined to be quite similar from a crushing and grinding perspective. The average value of the hypogene and supergene rock types is 11.3 kWh/t with a corresponding 75th percentile BWi value of 12.3 kWh/t, although 12.1 kWh/t was selected for purposes of the design criteria. However, for purposes of the design of the grinding circuit, the equivalent Bond Rod Mill Work Index (RWi) of 13.0 kWh/t was selected.
13.6.4 JK SAG MILL COMMINUTION TEST RESULTS
SMC tests in which the drop weight index (DWi) and the JK breakage parameters A and b were determined, were completed on samples from the Phase I and Phase II test programs. The results obtained were reviewed by JKTech Australia. The complete results were reported in the respective comminution test reports. DWi values obtained ranged from 1.23 to 12.55 kWh/m3 and the 75th percentile value was reported as 8.61 kWh/m3. DWi values were also grouped by copper rock type classification and the summary of the results obtained is shown in Table 13.11.
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Table 13.11 Drop Weight Index Results by Rock Type Classification
Rock Type Classification Statistical Outcome (kWh/t) Hypogene Supergene Oxide Minimum 2.44 1.23 1.84 Maximum 12.20 12.55 5.24 Average 7.38 6.43 3.32 75th Percentile 8.69 8.61 4.11 No. of Samples Tested 38 28 6
The Transmin report indicated that the Axb values ranged from 21.02 to 184.09. The Axb value indicates the power required but it is the inverse of the BWi values in that a lower Axb value indicates a harder type of material. The highest values of Axb were for supergene material tested in Phase I. These high values for Axb could be partially attributed to the location of these samples which were selected from the transition zone of the deposit.
The calculated values were reported, categorized and ranked according to the JKTech database. A total of 21 samples were tested but the results of only 15 samples were fully documented in the SMC report. These 15 Axb results reported for the Phase I test work ranged from 27.4 to 184.2 with an average value of 67.6. This average value of 67.6 would categorize the material tested as “moderately soft” and only two of the samples tested were categorized as “hard”. The 51 Phase II sample results obtained indicated that material tested was much harder. These samples had considerably more samples at depth in the deposit, with the Axb values ranging from 21.0 to 124.4 with an average of 39.8. This average Axb value of 39.8 corresponds to a “moderately hard” ranking. Also, 13 results were ranked as “very hard”, 21 as “moderately hard to hard”, 8 as “medium”, and 5 as “moderately soft to very soft” according to the JKTech database.
These results are consistent with other comminution results and have been included in the PDC, although these values were not required as a basis for design at this time.
13.7 F LOTATION T ESTS
The main and most recent flotation test programs were carried out in parallel at the two laboratories, Plenge and G&T, during 2012. The results were summarized and reported in the 2012 Transmin report as individual laboratory reports were not available at the time this report was compiled. The sequencing of test work evaluation differed at both laboratories. Therefore, in order to compare the results, the tests will not be evaluated in sequence but rather by flotation parameter.
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13.7.1 BULK SULPHIDE FLOTATION
Initial bulk sulphide flotation tests of 35 individual samples along with the three composite samples were carried out at the Plenge laboratory only. Potassium amyl xanthate (PAX) as the collector and methyl isobutyl carbinol (MIBC) as the frother, were used in order to generate bulk sulphide concentrates. The samples were ground to a primary size P80 of 250 µm, conditioned with the collector and floated for 10 minutes. Portions of each of the samples were assayed and material balances were developed with the remaining concentrate samples stored for future mineralogical work.
Copper concentrate grades obtained ranged from 0.34 to 7.71% copper with recoveries ranging from 7.22 to 94.4% copper. The mass generated in the rougher concentrate ranged from 4.65 to 18.04%.
Various graphs were generated using these results. The poorest performing samples were classified as oxide or mixed samples. All samples of hypogene or supergene had copper recoveries in excess of 70%.
Figure 13.1 gives the results of copper head grade versus recovery by rock type classification. There is no apparent correlation between recovery and feed grade, or the feed material type with the exception of oxide and mixed mineral types.
Figure 13.1 Copper Head Grade versus Recovery for Rock Type
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Analyses for gold, sulphur and iron were also conducted and recorded. Although this test work program was conducted as a form of variablity testing, intermediate assays do not appear to have been done and therefore flotation kinetics cannot be determined.
13.7.2 FLOTATION PROGRAMS
In order to define the flotation circuit design, initial tests were performed on the three composite samples, followed by variability testing of the individual samples and this was followed by locked cycle tests (LCTs) using the composite samples. There was some variation in the number of samples tested at each laboratory. Each facility tested the three master composite samples, while Plenge tested 35 variability samples, and G&T tested 30 samples for the variability tests.
As mentioned previously, the sequence of flotation parameter testing varied at the two laboratories with grind size evaluation occurring at the G&T facility only.
13.7.3 FLOTATION TESTING – GRIND SIZE EVALUATION
The characterization of this parameter was only carried out at the G&T laboratory. The grind tests were conducted on the composite samples using PAX as a collector and MIBC as a frother. A pH value of 10 was chosen for this initial evaluation. Flotation time was taken to be eight minutes. The three targeted grind sizes were P80 sizes of 210, 150, and 105 µm.
In this methodology, the PAX was used as a collector and therefore a larger mass is anticipated due to the relatively unselective nature of this bulk sulphide collector. Collector addition rates were low and seemingly close to starvation levels. However, some inconsistency in the test work procedure was obtained during the testing since the amount of PAX reagent added varied per test.
Composite ZFC-MC01 showed very fast flotation rates for copper at all grind sizes while the overall gold recovery was highest at the finest grind size. Mass recoveries to the concentrate were inconsistent. The data analysis does not clearly indicate if the gold is following the pyrite or the chalcopyrite, and this observation also holds for other aspects of the test work. The graph of flotation time versus copper rougher recovery indicates that flotation is not complete after eight minutes, particularly at the coarser grinds.
Composite ZFC-MC02 shows slightly slower flotation kinetics although the PAX addition is higher than that in ZFC-MC01 making direct comparison of the results obtained difficult. A lower overall copper recovery was obtained on the ZFC-MC02 sample.
The effect of grind size was shown to have the most effect on the sample ZFC- MC03.
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The Transmin selection of 147 µm as the primary P80 grind size value obtained is based on the decreased iron (pyrite) recovery, increased gold recovery and similar copper recovery as compared with the coarser grind of about 210 µm. Although this conclusion is not completely incorrect, it is not shown that the gold necessarily follows the copper and may in fact follow the iron (pyrite). The gold also displays a slower flotation kinetic rate. The grind size P80 of 147 µm indicated is a good general choice since it was shown that the initial flotation kinetics varied only slightly based on grind size in the size range selected for testing. There would be little advantage to grinding to a finer P80 of 100 µm as the grinding costs would become prohibitive particularly since the apparent advantage is not significant. Further analysis of the test work may indicate that a slightly coarser grind could also achieve the same results as apparently indicated in the bulk mineral flotation tests by Plenge. For purposes of the process design, a P80 value of 150 µm was used in the design criteria.
13.7.4 FLOTATION – COLLECTOR REAGENT EVALUATION
Based on the initial test results obtained, Transmin concluded that the gold did not follow the iron (pyrite), and therefore it would be advantageous to reduce the amount of pyrite floated in the rougher stages prior to cleaning. A collector evaluation test program to reduce the amount of pyrite floating in the rougher concentrate was carried out independently at both laboratories with a variety of collector combinations. All the evaluation tests were based on copper and iron recoveries to rougher flotation. However, it does not appear that collector dosage was a variable that was evaluated and this should be added to a future level of study.
G&T evaluated four reagent schemes with a rougher flotation time of eight minutes. G&T used very small quantities of reagent and obtained over 90% recovery of copper to the rougher concentrate for ZFC-MC01, 89% copper recovery for ZFC-MC02 and approximately 75% copper recovery for ZFC-MC03 with all reagent choices. In choosing the reagent to continue with in the test program, other elements were analyzed for recovery and grade.
The copper recovery versus mass recovery graphs indicate that the mass recovery to the rougher concentrate for ZFC-MC01 had the most significant impact on the overall recovery.
The collector reagent 3418A appeared to enhance the amount of the gold reporting to the concentrate. The collector reagent 3418A was chosen on the basis of giving the best results for composite ZFC-MC03, even though this is the transition/oxide sample. The best gold recovery is shown with this material. As demonstrated for the sample ZFC-MC01, the mass recovered to the concentrate appears to be limiting the amount of copper recovered.
Although the reagent type 3418A was selected, the reagent dosage used in the variability test work was increased in order to ensure that the highest attainable copper and gold recoveries were obtained in the variability testing.
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Plenge evaluated eight different collector reagent schemes and used an overall rougher flotation time of 18 minutes. Because of the much longer flotation times used for these tests, the mass recoveries to the rougher concentrates and the resulting recoveries were significantly higher.
The Plenge tests all showed very fast rougher kinetics with the majority of the recovery occurring within the first two minutes for all three composite samples and all the reagent schemes selected. The final choice of collector reagent was a combination of 15 g/t A3302 and 15 g/t 3418A. This was not necessarily the optimum choice for all the composite samples tested but the choice was made in order to have commonality for the reagents and dosage rate. The overall choice appears to be based on the apparently low iron (pyrite) recovery to the concentrate.
On review, it would appear that other reagent schemes may be better suited to achieving equivalent or better copper and gold recovery in the rougher stage of flotation, and the reduction of iron (pyrite) could be better targeted in the cleaner circuit. The reagent PAX at 40 g/t should not be ruled out as a primary collector as in some cases it appears that the gold is following the iron (pyrite) despite the initial interpretation of results that the pyrite did not contain gold. This aspect remains inconclusive and warrants further investigation.
Kerosene collector addition was also evaluated but only at Plenge. The kerosene addition had various effects of the recoveries of the copper, iron and gold depending on the sample tested. There was an increase in copper and iron recovery in each instance with a corresponding increase in mass recovery to the concentrate. Based on the apparent non-selectivity of the reagent, Plenge decided not to select kerosene as a collector reagent.
13.7.5 FLOTATION – PH EVALUATION
The pH value in the rougher flotation was evaluated at both laboratories with the stated objective of reducing the iron (pyrite) recovery.
During reagent testing at G&T, the reagent scheme was not fixed. The iron (pyrite) was depressed with the increased pH, but gold recovery also appeared to be affected to some extent. However, no correlation was attempted.
The Plenge test results have a more consistent mass pull and results would indicate that a pH of 10.5 for two of the samples tested gave the best results, with a pH of 9.5 for the ZFC-MC03 sample.
According to the report from Transmin, the pH value of 10 was chosen based on the combined results from G&T and Plenge. The conclusion is that the pH value of 10 gives the best recovery results for copper, and gold with the least recovery for iron (pyrite). However, this comment from Transmin is surprising since the LCT program at G&T used a pH value of 10 and gave very poor results, and a pH value of 11 was used in a repeat LCT series of work. The reason(s) for the discrepancy in the results
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obtained by G&T for the pH 10 results were not discussed, and it may have been a procedural problem.
This pH value of 10 is included in the design criteria, and although it is subject to confirmation, it appears to be the optimal value.
13.7.6 FLOTATION – ROUGHER FLOTATION TIME
Rougher flotation times prior to the variability and locked cycle testing were determined based on copper recovery results. The optimal flotation time was taken at the point where the copper partial grade of the rougher concentrate was calculated to be the same as the feed grade.
The initial G&T test work had a rougher flotation time of eight minutes, and it is possible that further copper flotation could occur with extended flotation time.
The Plenge kinetic test work had flotation times of 18 minutes allowing for a clear graphical analysis. A conservative flotation time of 10 minutes was selected, although an analysis of the results indicates that eight minutes may suffice in certain instances. However the additional time up to 10 minutes will allow for the apparently slower floating gold to be recovered and will also accommodate any variability in the mineralized material types.
The ten minute rougher flotation time parameter was included in the design criteria.
13.7.7 FLOTATION – REGRIND
Concentrate regrind size was determined using a three-stage cleaning test procedure. Regrind sizes of 40 µm and 20 µm were targeted during testing.
Transmin interpreted the test work to indicate that a 40 µm regrind is suitable for this application with acceptable copper grades and recoveries being achieved. Transmin also gave the additional statement that a finer regrind size would also work, although this option would be more costly.
The evaluation does not include grade recovery curves which may also indicate that fewer cleaner stages may be required. Recoveries at the same common grade as the various grind sizes were not shown in the graphs. Gold recoveries were not evaluated. The acceptable copper grade was not defined and would be required to determine the necessary number of cleaner stages and regrind size. Grinding of the concentrates to the finer size appears to enhance gold recovery for some of the samples.
Further laboratory flotation testing used a regrind P80 size of 40 µm. The design criteria initially indicated that a regrind size P80 of 20 µm was required. However, a subsequent review of the test work data presented by Transmin indicates that a P80 regrind size of 40 µm may not be conclusive, but might suffice for purposes of the
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study, although the detailed testing across a wider spread of regrind sizes is required for the next phase of the Project.
13.7.8 FLOTATION – CLEANER PH EVALUATION
The cleaner flotation pH testing was done at G&T with a defined procedure of cleaner flotation conditions. A pH value of 10 was chosen in order to obtain acceptable copper concentrate grades.
An evaluation of the results appears to indicate that gold recovery follows the iron (pyrite) recovery in composite samples ZFC-MC02 and ZFC-MC03. However, since the target mineral is copper, any enhancements or use of conditions favouring the recovery of gold should only be undertaken when there is no detrimental effect to the copper grade or recovery, and when additional processing costs are not incurred.
13.7.9 FLOTATION – LOCKED CYCLE TESTING
LCT conditions were defined from the preliminary rougher and cleaner testing discussed in earlier sections and as outlined in the Transmin report. The initial parameters include a primary grind P80 of 150 µm, rougher flotation time of 10 minutes at pH 10, a regrind size P80 of 40 µm, cleaner flotation times of the first cleaner 5 minutes, the second cleaner 4 minutes, the third cleaner 3 minutes, and the cleaner scavenger 4 minutes. All the cleaner stages were conducted at a pH value of 10. However, the test procedures and the reagent schemes differed at both laboratories.
The LCT procedures differed between G&T and Plenge with the recycle streams being recycled and returned to different areas of the flotation circuit within the two test programs as presented in Figure 13.2 and Figure 13.3.
Plenge had six cycles of flotation for each LCT whereas G&T had five cycles. Plenge used 15 g/t A3302 and 15 g/t 3418A collector reagent combination for the LCT work, while G&T used variable amounts of only the 3418A collector reagent.
The initial LCT program completed at G&T used a pH value of 10 and had very poor copper concentrate grade. The test was repeated at the higher cleaner pH value of 11.0. However, other parameters in the test procedure should also have been analyzed. Also, the LCTs were unstable and the recycle streams could be causing problems in this regard. A significant amount of gold was rejected in the first cleaner scavenger tails. This may be related to regrind size, or alternatively, an indication that the iron (pyrite) and gold contents are related.
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Figure 13.2 LCT Procedure – Plenge
Feed
Grinding
Rougher Tail Rougher
Regrinding
Cleaner Scavenger Tail Cleaner Scavenger Cleaner 01
Cleaner 02
Cleaner 03
Cleaner 03 Concentrate
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Figure 13.3 LCT Procedure – G&T
Feed
Grinding
Rougher Tail Rougher
Regrinding
Cleaner Scavenger Tail Cleaner Scavenger Cleaner 01
Cleaner 02
Cleaner 03
Cleaner 03 Concentrate
The G&T LCTs conducted at a pH value of 11.0 were found to give reasonable concentrate grade and recovery values, as shown in Table 13.12. However, these test results from G&T still resulted in an unstable flotation procedure with the copper concentrate grade objective values not attained in one instance, and questionable in another.
The LCT’s conducted at Plenge were more stable and the target grade was made in each case using a pH value of 10 in the cleaner circuit.
Statistical analyses of the results were presented by Transmin for grade and recovery values at a 95% confidence level. A summary of the findings are shown in Table 13.12.
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Table 13.12 Locked Cycle Concentrate Summary
Element Recovery to Final Concentrate Grade Concentrate Cu Fe S Au, Cu Fe S Au Laboratory Composite (%) (%) (%) (g/t) (%) (%) (%) (%) Plenge ZFC-01 29.8 ± 32.8 ± 34.9 ± 4.28 ± 85.6 ± 11.2 ± 17.3 ± 40.8 ± 4.2 2.9 2.5 2.07 2.2 4.8 7.5 9.7 (Cycle D-F) ZFC-02 38.5 ± 22.7 ± 33.2 ± 4.94 ± 81.3 ± 6.68 ± 15.7 ± 37.7 ± 3.9 1.5 1.5 2.03 0.6 1.06 2.23 10.8 Cleaner pH 10 ZFC-03 31.8 ± 29.3 ± 35.6 ± 3.13 ± 67.3 ± 6.10 ± 13.9 ± 26.6 ± 1.5 1.4 1.1 0.38 1.7 0.40 0.6 2.9 G&T ZFC-01 24.2 ± 30.2 ± 36.2 ± 2.88 ± 91.0 ± 15.2 ± 23.8 ± 51.3 ± 5.6 2.8 0.9 0.47 8.0 0.3 4.3 18.0 (Cycle IV-V) ZFC-02 33.8 ± 23.9 ± 34.9 ± 3.88 ± 85.1 ± 9.08 ± 18.1 ± 45.9 ± 16.2 8.8 6.0 2.65 2.7 6.45 8.1 15.0 Cleaner pH 11 ZFC-03 29.4 ± 28.9 ± 36.9 ± 4.16 ± 68.7 ± 7.11 ± 13.7 ± 36.0 ± 12.9 5.7 5.7 7.61 4.4 3.51 5.1 22.1
The results obtained from Plenge along with the test procedure used, gives a much more stable operating flowsheet. It was indicated that the commercial concentrate grade objective was 28% copper and it is apparent that this grade is easily achieved within the test program. Further data analysis could potentially enable a higher recovery of copper and gold at an acceptable copper concentrate grade to be achieved through process parameter changes and optimization of the procedure.
The LCT procedure as given for Plenge was used in the design of the flotation plant. Clearly the Plenge procedure is a more logical method, although in addition to the Plenge procedure, the regrinding of the cleaner scavenger concentrate and the second cleaner tailings is also advocated.
13.7.10 FLOTATION – FINAL CONCENTRATE ELEMENTAL ANALYSIS
No deleterious elements in any significant concentration were found in the copper concentrate produced, and all the impurity elements were found to be below penalty limits.
13.7.11 FLOTATION – VARIABILITY TESTS (ROUGHER)
Two sets of variability testing were done on similar samples using the rougher flotation parameters from the LCT procedures.
G&T completed 30 variability tests of which approximately half of the number of tests had rougher flotation times of 8 minutes and the remaining tests had 10 minutes. Copper recoveries ranged from 9.63 to 97.5%, and rougher concentrate grades
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ranged from 0.34 to 10.1% copper. The results of the feed grade versus recovery based on rock type are given in Figure 13.4.
Figure 13.4 Rougher Recovery versus Copper Feed Grade – G&T Variability Tests
Three of the lowest recoveries were on samples with low feed grades of a rock type associated with the leach cap and are possibly oxidized material. These samples were not considered to belong to the sulphide deposit and were not used in further comparisons in the report.
No correlation is seen to exist between the rougher concentrate copper recovery and the copper feed grade. Further analysis by Transmin showed no correlation for gold, iron (pyrite) or sulphur.
Variability tests were also completed by Plenge on 11 of the same samples as G&T had used, together with some additional variability samples included which allows for a side-by-side comparison of the results obtained by the two testing facilities. The rougher flotation tests were done for a period of 18 minutes with correspondingly higher metal and mass recoveries. Plenge completed 35 tests and the overall results for copper are shown in Figure 13.5. Copper recoveries ranged from 21.3 to 96.6% and rougher concentrate grades ranged from 0.33 to 8.5% copper. The results were grouped by zone, alteration type and rock type. Again, no correlation was found between the copper feed grade and the copper recovery.
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Figure 13.5 Rougher Recovery versus Copper Feed Grade – Plenge Variability Tests
A comparison of a portion of the results obtained from the variability tests from both laboratories, G&T and Plenge, was reported and is shown in Table 13.13. The rougher concentrate recovery was taken at a flotation time of 10 minutes for both laboratories (shown as t ~ 10 minutes in Table 13.13). One significant difference between the test programs was that the collector used for the variability tests by G&T was 3418A at a dosage rate of 15 g/t, while Plenge used a mixture of A3302 and 3418A, each at a dosage rate of 15 g/t.
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Table 13.13 Variability Tests – Comparison of Results
Calculated Head Grade Rougher Recovery at t ~ 10 minutes Cu (%) Fe (%) Au (g/t) Cu (%) Fe (%) Au (%) Sample ID Plenge G&T Plenge G&T Plenge G&T Plenge G&T Plenge G&T Plenge G&T ZF-02 0.35 0.31 4.54 3.77 0.17 0.18 88.1 92.6 42.7 65.5 60.2 75.6 ZF-03 0.73 0.68 3.49 3.17 0.06 0.07 91.3 93.9 41.9 81.0 68.9 72.9 ZF-10 0.72 0.58 3.18 2.65 0.12 0.09 79.4 73.6 36.8 27.6 51.6 69.8 ZF-11 0.48 0.42 4.38 4.09 0.17 0.14 92.6 96.3 49.5 55.2 63.5 81.2 ZF-14 0.79 0.70 2.32 1.92 0.05 0.04 90.3 94.5 41.0 89.2 64.1 80.0 ZF-16 0.29 0.26 5.97 5.07 0.12 0.07 15.1 21.7 13.6 19.4 26.0 65.4 ZF-19 0.40 0.42 2.42 2.29 0.10 0.11 90.6 96.5 74.5 75.0 65.1 82.7 ZF-26 0.39 0.37 3.23 2.84 0.26 0.25 93.7 96.6 47.5 54.4 80.9 89.1 ZF-39 0.32 0.27 5.47 3.36 0.17 0.04 94.2 96.7 73.8 71.1 76.8 57.5 ZF-41 0.55 0.57 3.44 3.16 0.12 0.08 44.9 38.8 19.3 14.4 39.7 51.9 ZF-55 0.54 0.54 4.71 4.69 0.33 0.29 75.1 74.8 28.5 32.7 54.9 60.4 Average 0.51 0.47 3.92 3.36 0.15 0.12 77.8 79.6 42.6 53.2 59.2 71.5
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A comparison of the results obtained indicates the following:
• Mass recoveries after ten minutes of flotation time were not reported and would be interesting for comparison purposes. • Feed grades for copper, iron (pyrite) and gold were typically reported higher by Plenge, as shown in Figure 13.6, for copper by way of example. In the case of iron, the assay differences between Plenge and G&T is significant. • Copper recovery averaged slightly higher in the G&T tests. • Iron recovery was lower at Plenge with a correspondingly lower gold recovery. As mentioned previously, the possible association of gold with iron (pyrite) needs confirmation. Further detailed analyses of the test work conducted, or additional test work, is required if gold recovery is to be targeted.
Figure 13.6 Assay Comparison, Copper Head Grade
It is apparent that some coordination characterization of testing at specific slurry pH values, and assaying methodologies, will be required between the Plenge and G&T laboratories in order to standardize procedures.
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13.8 S ETTLING T EST W ORK (TAILINGS)
G&T carried out settling test work using tailings samples produced from open circuit flotation tests for each of the three composite samples.
Three flocculants were tested on each sample at various dosage rates. Composite sample ZFC-MC03 was found to require slightly more flocculant probably as a result of the oxidized nature of the sample. Either the anionic or high anionic flocculants would be acceptable choices base on the limited test work done.
13.9 E NVIRONMENTAL T ESTING
Acid based accounting (ABA) test work was done on specific samples. Over 50% of the samples tested were determined to be acid producers while 30% of the samples tested remain as uncertain.
13.10 S OLVENT E XTRACTION AND E LECTROWINNING P ROCESSING F ACILITY
A limited amount of test work to recover copper from oxide samples from the Zafranal deposit, was conducted under the direction of Transmin. The recovery of the copper was based on the heap leach process using acid leaching on pads, and SX-EW.
13.11 C ONCLUSIONS
Within the section titled “Conclusions”, Transmin reported the process parameters that were to be chosen for plant design, a summary of which appears in Section 17.0.
13.12 R ECOMMENDED A DDITIONAL T EST W ORK
No recommendations for further test work were identified by Transmin in the report that was issued. However, Transmin and Tetra Tech recommends the following additional test work and evaluations:
• Material handling characterization including specific gravity, bulk density/particle size, etc. • A more detailed grinding evaluation of the Zafranal rock types.
• Confirmation of the regrind P80 size is required; test work should be conducted to highlight the sensitive ranges of the parameters selected.
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• The Plenge flotation procedure used for LCT work requires some modification, particularly with regard to the cleaner scavenger concentrate collected. Confirmatory testing should include confirmation of the revised regrind size, the collector reagent suite and the slurry pH. The G&T flotation circuit has been shown to be unstable and should therefore not be considered in any future test work program. • The disparity in the results obtained between Plenge and G&T at different pH values requires qualification.
• Although the primary grind P80 particle size has been determined to be 150 µm, the P80 of the rougher concentrate produced has not been determined to date and is required for the sizing of the regrind circuit. The potential of using a coarser primary grind also requires additional test work. • Test work is required to finalize the collector reagent suite to be used for Zafranal material. Plenge used a mixture of A3302 and 3418A, while G&T used PAX and then switched to the reagent 3418A only but for reasons which were not disclosed. In order to facilitate comparison of results, the suggested test work should include: A3302 only 3418A only PAX only A3302 and 3418A PAX and 3418A. The potential for using a relatively inexpensive basic collector type augmented by a more costly specific reagent such as 3418A at a lower dosage rate potentially has significant operating cost implications. A consistent grind, pH and dosage rate above starvation levels is required. • Although gold is not a major financial contributor to the value of the concentrate (estimated value about 6% of the total value of the concentrate), it is vital to understand its behaviour in the flotation circuit, and at which stage of the process the gold is lost to tailings. • Copper concentrate needs to be generated for settling and filtration characterization. • Flotation tailings testing for settling and filtration characterization should also be conducted. • A more detailed revision of the primary grinding circuit is recommended. A brief review of the grinding circuit has indicated that the plant feed treatment rate of 80,000 t/d could be exceeded with the present design. Since the available grinding test data is suitable only for this PEA level report, a more detailed assessment is not possible, although a cursory review indicates that the grinding circuit throughput could be increased to about 100,000 t/d based on the present design parameters. A more detailed review using more specific grinding data based on pit depth, and varying the grinding
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circuit parameters such as transfer size, grinding media addition, and the use of a pebble crushing circuit, could result in a substantial increase in the daily throughput rate. It must be noted however, that a detailed review of the downstream processing facilities will also be required.
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14.0 MINERAL RESOURCE ESTIMATES
In February 2011, AMEC-Minproc completed a resource estimate (AMEC-Minproc, 2011) for the Zafranal Main Zone. AMEC-Minproc used a 0.2% copper grade shell to constrain their mineral zone model and stated their resource at a lower threshold of 0.2% copper. Their estimate is summarized in Table 14.1.
Table 14.1 AMEC-Minproc 2011 Zafranal Main Zone Resource Estimate at 0.2% Copper Cut-off
Tonnes Cu Au Class (t) (%) (g/t) Measured 17,000,000 0.93 0.09 Indicated 284,000,000 0.44 0.08 Inferred 51,000,000 0.32 0.06
The basis of the following estimate differs from the AMEC-Minproc estimate in two significant respects: 1) grade shells were not used; the estimate is based upon values for the entire modeled mineralized envelope, and 2) the current estimate pertains not only to the Zafranal Main Zone but also includes a first-time estimate for the Victoria, Sicera Norte and Sicera Sur Zones.
14.1 E XPLORATORY D ATA A NALYSIS
14.1.1 ASSAYS
Tetra Tech received location, survey and assay data for the Zafranal Main, Victoria, Sicera Norte and Sicera Sur mineral zones in Microsoft Access® database format. Table 14.2 summarizes the number of drillholes and samples for each zone.
Table 14.2 Zafranal Property Drillhole and Sample Quantities
No. of Metres No. of No. of Metres No. of Zone Core Holes (m) Samples RC Holes (m) Samples Main Zone 192 67,682 33,860 73 22,953 15,637 Victoria 31 10,765 5,379 9 2,729 1,662 Sicera Norte 47 14,504 7,381 10 3,232 2,861 Sicera Sur 3 1,018 448 34 11,456 10,575 Total 273 93,969 47,068 126 40,370 30,735
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The data was imported into Gemcom GEMS™ 6.3 and checked for integrity (logical errors). No errors were found.
As the data are comprised of both drill core and RC assays it was considered possible that there were statistically meaningful differences between the two populations so the data were separated and T-tests (assuming equal variances) were performed. No statistically-meaningful differences were found between the two populations so they were combined for the purpose of the resource estimate that follows.
14.1.2 CAPPING
Capping is the process of reducing high values within a sample population that are regarded as statistically anomalous with respect to the population as a whole to avoid the influence these values would have on the statistical characteristics of the population if left at their full value. The risk in including statistically high values in a resource estimate is that their contribution to the estimated grade will be disproportionate to their contribution to the tonnage and therefore the grade of the resource as a whole will be overstated.
Cumulative frequency curves for gold and copper assay values for all four mineral zones were constructed and examined for the presence of outliers. No unequivocal evidence of outliers was found so no capping of either copper or gold values was carried out.
14.1.3 COMPOSITES
Compositing of samples is done to overcome the influence of sample length on the contribution of sample grade. About 95% (47,281) of all samples, both drill core and reverse-circulation, from the Zafranal Main Zone are equal to 2 m or less in length; therefore all Zafranal Main samples were composited to 2 m. The Zafranal Main sample population represents about 64% of the entire sample population and, as most of the samples from the other three zones were collected in 2 m or shorter lengths (Victoria = 94%; Sicera Norte = 93%; Sicera Sur = 99.8% (88% of the Sicera Sur samples were from RC drilling and all were 1 m in length)) all samples for all zones were composited to a length of 2 m.
14.2 B ULK D ENSITY
AQM provided 878 bulk density measurements of samples of drill core collected at 40 m intervals downhole. These samples were correlated with the corresponding copper solubility zones that form the basis of the resource estimate and a mean bulk density was established for each zone; these are presented below in Table 14.3. Note the values for Sicera Norte were adopted from measurements made at Sicera Sur.
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Table 14.3 Zafranal Mineral Zone Bulk Density Measurements
Density Mineral Zone (g/cm3) Zafranal Main Hypogene 2.66 Zafranal Main Leached 2.36 Zafranal Main Oxide 2.39 Zafranal Main Supergene 2.48 Victoria Hypogene 2.69 Victoria Leached 2.36 Victoria Mixed 2.63 Victoria Supergene 2.31 Victoria Transitional 2.73 Sicera Norte Hypogene 2.67 Sicera Norte Mixed 2.58 Sicera Norte Supergene 2.40 Sicera Norte Transitional 2.44 Sicera Sur Hypogene 2.67 Sicera Sur Leached 2.36 Sicera Sur Mixed 2.58 Sicera Sur Oxide 2.39 Sicera Sur Supergene 2.40 Sicera Sur Transitional 2.44
14.3 G EOLOGICAL I NTERPRETATION
AQM supplied geological solids representing lithology, alteration, and copper solubility zones. As AQM requested that the resource estimate be based upon copper solubility and as copper solubility appears to be more important than lithology or alteration with respect to the distribution of mineralization, only the solubility solids were used for the resource estimate.
Copper solubility was determined for all samples containing more than 0.2% copper and is based upon the percent of total copper (CuT) that was liberated in three sequential digestions: 1) sulphuric acid (CuS); 2) cyanide (CuCN); and 3) nitric and hydrochloric acid (CuNH). Solubility was categorized as (in idealized order from bottom to top): hypogene, transitional, supergene, mixed, oxide and leached. The Hypogene Zone contains less than 20% of total copper that is soluble in sulphuric acid and cyanide, i.e. 80% or more of total copper is soluble only in nitric and hydrochloric acid. The Transitional Zone contains more than 20% total copper that is soluble in sulphuric acid and cyanide but less than 50% that is cyanide soluble alone, and the cyanide solution component is also higher than the sulphuric acid soluble component. The Supergene Zone contains more than 50% of total copper that is cyanide soluble. The Oxide Zone contains more than 30% of total copper that is
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soluble in sulphuric acid but where the sulphuric acid soluble component is higher than the cyanide soluble component. The base of the Leached Zone represents the threshold at which the ratio between sulphur content and iron content is 20% or less. This ratio is supplemented where possible by the cyanide solubility limit for defining the supergene zone. An additional category, “Mixed”, contains copper oxides and secondary copper sulphides that are between 20% and 70% cyanide-soluble but where the sulphuric acid soluble component is greater than 20%.
Metallurgical test work of the various rock types indicated that the geometallurgical characteristics of the hypogene and transitional zones were very similar in the Main Zone deposit and as a result these two zones were combined for the resource calculation. Similarly, the geometallurgical characteristics of the oxide and mixed zones were very similar in the Main Zone deposit and were also combined for the resource calculation. As little or no metallurgical test work was carried out on the other deposits, all rock types were reported separately for each deposit. Only the Sicera Sur Zone reports the full set of solubility zones; the Victoria Zone is comprised of leached, mixed, supergene, transitional, and hypogene; and Sicera Norte contains only mixed, supergene, transitional and hypogene.
Perspective views of the copper solubility zones in each of the four mineral occurrences are shown in Figure 14.1 to Figure 14.3.
Figure 14.1 Perspective View of the Zafranal and Victoria Mineral Zones
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Figure 14.2 Perspective View of the Sicera Norte Mineral Zones
Figure 14.3 Perspective View of the Sicera Sur Mineral Zones
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14.4 S PATIAL A NALYSIS
Variograms were modelled using Sage2001 software on the basis of copper composites. Separate variograms were estimated for the hypogene and non- hypogene portions of the Zafranal Main and Victoria Zones. However, the Sicera Norte and Sicera Sur Zones were modelled as single variograms for the entire dataset because each of the zones contains relatively little data and it was considered improbable that meaningful variograms could be obtained if the approach used was the same as for Zafranal Main and Victoria Zones. Table 14.4 is a compilation of the parameters for the various variograms and respective search ellipses. The same parameters were used for the interpolation of both copper and gold.
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Table 14.4 Zafranal Property Variograms and Search Ellipse Parameters
Model Uncapped Composites
Zafranal Main Hypogene Structure Weight C1 C1 C1 C1 C2 C2 C2 C2 Axis Rotation Axis Range Axis Rotation Axis Range C0 0.238 Z 3 X 70 Z 39 X 440 C1 0.362 Y 27 Y 113 Y -24 Y 740 C2 0.391 Z 87 Z 60 Z 56 Z 127 ZMH Search Ellipse Axis Rotation Axis Range Z 39 X 130 Y -24 Y 740 Z 56 Z 440 Zafranal Main Non-Hypogene Structure Weight C1 C1 C1 C1 C2 C2 C2 C2 Axis Rotation Axis Range Axis Rotation Axis Range C0 0.259 Z -115 X 10 Z 7 X 150 C1 0.392 Y 7 Y 10 Y -87 Y 160 C2 0.349 Z 35 Z 105 Z 76 Z 400 ZMNH Search Ellipse Axis Rotation Axis Range Z 7 X 150 Y -87 Y 160 Z 76 Z 395 table continues…
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Model Uncapped Composites
Victoria Hypogene Structure Weight C1 C1 C1 C1 C2 C2 C2 C2 Axis Rotation Axis Range Axis Rotation Axis Range C0 0.291 Z -7 X 75 Z 59 X 200 C1 0.228 Y 9 Y 50 Y 24 Y 115 C2 0.481 Z -83 Z 15 Z -38 Z 500 VH Search Ellipse Axis Rotation Axis Range Z 59 X 200 Y 24 Y 115 Z -38 Z 500 Victoria Non-Hypogene Structure Weight C1 C1 C1 C1 C2 C2 C2 C2 Axis Rotation Axis Range Axis Rotation Axis Range C0 0.209 Z -10 X 50 Z -58 X 350 C1 0.547 Y 45 Y 115 Y -73 Y 500 C2 0.217 Z 0 Z 40 Z 87 Z 165 VNH Search Ellipse Axis Rotation Axis Range Z -58 X 345 Y -73 Y 500 Z 87 Z 165 table continues…
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Model Uncapped Composites
Sicera Norte All Zones Structure Weight C1 C1 C1 C1 C2 C2 C2 C2 Axis Rotation Axis Range Axis Rotation Axis Range C0 0.368 Z 19 X 330 Z -45 X 125 C1 0.478 Y 38 Y 40 Y 13 Y 500 C2 0.154 Z -2 Z 105 Z 39 Z 400 VH Search Ellipse Axis Rotation Axis Range Z -45 X 125 Y 13 Y 100 Z 39 Z 200 Sicera Sur All Zones Structure Weight C1 C1 C1 C1 C2 C2 C2 C2 Axis Rotation Axis Range Axis Rotation Axis Range C0 0.287 Z 54 X 245 Z -52 X 370 C1 0.454 Y -19 Y 40 Y 61 Y 150 C2 0.259 Z -36 Z 45 Z 25 Z 110 VNH Search Ellipse Axis Rotation Axis Range Z -52 X 370 Y 61 Y 150 Z 25 Z 107
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14.5 R ESOURCE B LOCK M ODEL
The resource block model was developed in the PSAD 56 datum as shown in this section, and subsequently transformed into WGS84 for the design of the mine plan. Because of their wide geographic separation, separate block models were constructed for each mineral deposit although all models were constructed with the same block size (x = 15, y = 15, z = 15 m). The block size approximates one selective mining unit (SMU).
Table 14.5 below lists the various parameters of each block model.
Table 14.5 Zafranal Property Block Model Parameters
Coordinates Size Rotation Origin (°) Dimensions Number (m) (°) Zafranal Main Block Model Parameters X 792750 Columns 190 15 0 Y 8223800 Rows 70 15 - Z 2930 Levels 75 15 - Victoria Block Model Parameters X 795200 Columns 65 15 0 Y 8223880 Rows 70 15 - Z 2930 Levels 60 15 - Sicera Norte Block Model Parameters X 785200 Columns 60 15 0 Y 8227960 Rows 45 15 - Z 2140 Levels 40 15 - Sicera Sur Block Model Parameters X 786600 Columns 65 15 0 Y 8224800 Rows 70 15 - Z 2330 Levels 45 15 -
14.6 I NTERPOLATION P LAN
Each model was interpolated in a single pass using the search ellipses described in Table 14.4. The interpolation process was constrained by the number of samples per block (minimum 8, maximum 40) and the maximum number of samples per hole (8) which at a minimum allowed a grade to be interpolated into a block on the basis of one hole. For all the models the interpolation process was restricted to using only samples from the zone for which values were being estimated.
Interpolations were carried out using ordinary kriging (OK).
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14.7 M INERAL R ESOURCE C LASSIFICATION
Resources within the Zafranal Main and Victoria Zones were classified as Measured, Indicated, and Inferred. Resources within the Sicera Norte and Sicera Sur zones were classified as Indicated or Inferred.
The Measured category is commonly used with the implicit meaning that no greater understanding of these resources could be obtained by additional drilling and is usually applied to resources defined by blast hole sampling or data of similar density; as applied to the Zafranal Main and Victoria Zones that have been defined to date only by exploration drilling, this term refers to that portion of those deposits for which a) there is abundant data and b) for which the grade is considered unlikely to change meaningfully from that already estimated if additional data is acquired.
Resource blocks classified as Measured required that the interpolated grade was based upon a minimum of six drillholes with a maximum mean sample-to-block distance of 100 m.
Resource blocks classified as Indicated required that the interpolated grade was based upon a minimum of four and a maximum of five drillholes and a maximum mean sample-to-block distance of 200 m.
Resource blocks classified as Inferred include all blocks for which a grade was interpolated but did not qualify as either Measured or Indicated and which are located within a mean sample-to-block distance of 300 m.
14.8 M INERAL R ESOURCE T ABULATION
The resources for each of the four mineral zones are tabulated below for a cut-off grade of 0.2% copper. Table 14.6 is a synopsis of the kriged estimates for all four zones. Summaries of the kriged estimates for each zone are given in Table 14.7 to Table 14.9.
Table 14.6 Zafranal Property Kriged Resource Synopsis at 0.2% Copper Cut-off
Tonnes Cu Au (t) (%) (g/t) Kriged: All Zones Measured Zafranal Main Zone Measured Total 107,907,000 0.43 0.08 Victoria Zone Measured Total 13,184,000 0.31 0.03 Kriged: Total Measured 121,091,000 0.42 0.08 Kriged: All Zones Indicated Zafranal Main Zone Indicated Total 319,042,000 0.38 0.08 Victoria Zone Indicated Total 54,731,000 0.25 0.04 Sicera Norte Zone Indicated Total 34,005,000 0.26 0.02 table continues…
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Tonnes Cu Au (t) (%) (g/t) Sicera Sur Zone Indicated Total 28,322,000 0.30 0.03 Kriged: Total Indicated 436,100,000 0.35 0.06 Kriged: Total Measured + Indicated 557,191,000 0.36 0.07 Kriged: All Zones Inferred Zafranal Main Zone Inferred Total 12,665,000 0.29 0.04 Victoria Zone Inferred Total 31,470,000 0.25 0.03 Sicera Norte Zone Inferred Total 6,170,000 0.25 0.02 Sicera Sur Zone Inferred Total 3,432,000 0.31 0.03 Kriged: Total Inferred 53,738,000 0.27 0.03 Note: Totals may not agree because of rounding.
Table 14.7 Zafranal Main Zone Kriged Resource Estimate at 0.2% Copper Cut- off
Tonnes Cu Au (t) (%) (g/t) Zafranal Main Measured Hypogene 48,638,892 0.29 0.08 Leached 5,056,850 0.28 0.13 Oxidized 2,073,769 0.27 0.14 Supergene 52,137,860 0.58 0.08 Measured Total 107,907,371 0.43 0.08 Zafranal Main Indicated Hypogene 212,892,550 0.28 0.07 Leached 12,755,709 0.34 0.10 Oxidized 2,711,902 0.28 0.14 Supergene 90,682,265 0.62 0.08 Indicated Total 319,042,425 0.38 0.08 Zafranal Main Inferred Hypogene 12,314,893 0.29 0.04 Leached 70,953 0.38 0.05 Oxidized 0 0.00 0.00 Supergene 279,370 0.31 0.03 Inferred Total 12,665,216 0.29 0.04
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Table 14.8 Victoria Zone Kriged Resource Estimate at 0.2% Copper Cut-off
Tonnes Cu Au (t) (%) (g/t) Victoria Measured Hypogene 1,530,959 0.27 0.04 Leached 90,037 0.24 0.06 Mixed 3,564,329 0.33 0.03 Oxide 0 0.00 0.00 Supergene 3,070,399 0.40 0.02 Transitional 4,928,300 0.26 0.04 Measured Total 13,184,024 0.31 0.03 Victoria Indicated Hypogene 33,608,361 0.26 0.04 Leached 180,928 0.24 0.04 Mixed 2,665,596 0.25 0.02 Oxide 0 0.00 0.00 Supergene 152,649 0.61 0.03 Transitional 18,123,360 0.25 0.03 Indicated Total 54,730,893 0.25 0.04 Victoria Inferred Hypogene 23,718,937 0.27 0.03 Leached 0 0.00 0.00 Mixed 31,563 0.24 0.02 Oxide 499,481 0.26 0.03 Supergene 0 0.00 0.00 Transitional 7,220,212 0.23 0.03 Inferred Total 31,470,193 0.25 0.03
Table 14.9 Sicera Norte and Sicera Sur Zones Kriged Resource Estimates at 0.2% Copper Cut-off
Tonnes Cu Au (t) (%) (g/t) Sicera Norte Indicated Hypogene 31,815,134 0.25 0.02 Mixed 0 0.00 0.00 Supergene 0 0.00 0.00 Transitional 2,189,701 0.29 0.01 Indicated Total 34,004,835 0.26 0.02 Sicera Norte Inferred Hypogene 5,422,871 0.25 0.02 Mixed 134,174 0.24 0.01 Supergene 531,834 0.27 0.01 table continues…
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Tonnes Cu Au (t) (%) (g/t) Transitional 81,288 0.36 0.02 Inferred Total 6,170,168 0.25 0.02 Sicera Sur Indicated Hypogene 17,447,158 0.30 0.02 Leached 3,284,467 0.30 0.03 Mixed 1,622,293 0.25 0.03 Oxide 3,698,234 0.32 0.04 Supergene 100 0.56 0.04 Transitional 2,269,853 0.27 0.02 Indicated Total 28,322,106 0.30 0.03 Sicera Sur Inferred Hypogene 779,943 0.31 0.03 Leached 656,215 0.30 0.02 Mixed 58,573 0.25 0.02 Oxide 1,457,521 0.26 0.03 Supergene 211,122 0.79 0.04 Transitional 268,911 0.25 0.01 Inferred Total 3,432,285 0.31 0.03
14.9 B LOCK M ODEL V ALIDATION
The block models were validated visually by comparing the fit of the model to the enclosing solids.
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15.0 MINERAL RESERVE ESTIMATES
A mineral reserve has not been estimated for the Project as part of this PEA.
A mineral reserve is the economically mineable part of a Measured or Indicated Mineral Resource.
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16.0 MINING METHODS
16.1 I NTRODUCTION
AQM commissioned NCL to prepare the mining study for the PEA of the Project, which contains four mineral zones of interest: Main Zone, Victoria, Sicera Norte and Sicera Sur, all containing similar porphyry-style copper mineralization.
16.2 I NITIAL I NFORMATION
The following information was provided for the development of the study:
• Surface Topography: Digital topographical drawings in WGS84 coordinate system with the UTM 18S horizontal projection, for all of the interest area including the Main Zone, Victoria, Sicera Norte and Sicera Sur, were provided by AQM. • Block Models: The resource block models used for the pit designs were prepared by Gregory Mosher, P.Geo. of Tetra Tech in March 2012, as four independent models, in the PASAD56 coordinate system which was later transformed to the WGS84 coordinate system by Aspi-System.
NCL received the block model information, in WGS84 coordinate system, as ASCII files and loaded them into Gemcom™ software, as four independent models.
BLOCK MODEL GRADE-TONNAGE TABLES
The grade-tonnage tables for the four deposits currently under study are provided in Table 16.1 to Table 16.4.
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Table 16.1 Grade-tonnage Table for Zafranal Main
Measured Indicated Measured+Indicated Inferred Cut-off Grade Tonnage Cu Au Tonnage Cu Au Tonnage Cu Au Tonnage Cu Au (%) ('000 t) (%) (g/t) ('000 t) (%) (g/t) ('000 t) (%) (g/t) ('000 t) (%) (g/t) 0.80 10,830 1.056 0.124 23,233 1.064 0.118 34,063 1.061 0.120 - 0.000 0.000 0.50 29,113 0.786 0.106 57,779 0.803 0.104 86,892 0.797 0.105 628 0.562 0.034 0.40 41,223 0.686 0.099 84,291 0.690 0.098 125,514 0.689 0.098 1,637 0.486 0.035 0.30 63,523 0.566 0.092 147,183 0.541 0.091 210,707 0.549 0.091 4,411 0.396 0.037 0.20 108,051 0.433 0.080 320,517 0.378 0.076 428,568 0.392 0.077 12,712 0.292 0.038 0.15 156,110 0.353 0.075 571,147 0.287 0.066 727,256 0.301 0.068 42,898 0.206 0.040 0.10 218,457 0.288 0.070 1,038,231 0.214 0.055 1,256,688 0.227 0.058 196,239 0.139 0.041 0.00 308,130 0.219 0.058 1,778,235 0.149 0.042 2,086,365 0.159 0.044 496,366 0.089 0.031
Table 16.2 Grade-tonnage Table for Victoria
Measured Indicated Measured+Indicated Inferred Cut-off Grade Tonnage Cu Au Tonnage Cu Au Tonnage Cu Au Tonnage Cu Au (%) ('000 t) (%) (g/t) ('000 t) (%) (g/t) ('000 t) (%) (g/t) ('000 t) (%) (g/t) 0.80 89 0.897 0.052 40 0.900 0.080 129 0.898 0.061 - 0.000 0.000 0.50 942 0.616 0.045 251 0.639 0.084 1,193 0.621 0.053 219 0.539 0.088 0.40 2,124 0.518 0.040 1,030 0.484 0.071 3,155 0.507 0.050 613 0.475 0.074 0.30 5,129 0.415 0.034 7,030 0.353 0.051 12,159 0.379 0.044 5,665 0.346 0.044 0.20 13,255 0.311 0.032 55,075 0.253 0.035 68,330 0.264 0.034 31,700 0.257 0.031 0.15 18,459 0.272 0.030 105,973 0.214 0.030 124,432 0.223 0.030 59,099 0.218 0.027 0.10 25,988 0.229 0.029 177,905 0.177 0.026 203,894 0.184 0.026 106,966 0.175 0.023 0.00 33,761 0.192 0.027 371,850 0.113 0.020 405,611 0.120 0.021 339,613 0.087 0.015
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Table 16.3 Grade-tonnage Table for Sicera Sur
Measured Indicated Inferred Cut-off Grade Tonnage Cu Au Tonnage Cu Au Tonnage Cu Au (%) ('000 t) (%) (g/t) ('000 t) (%) (g/t) ('000 t) (%) (g/t) 0.80 - - - 37 0.948 0.068 55 1.532 0.057 0.50 - - - 519 0.606 0.041 213 0.805 0.049 0.40 - - - 1,917 0.481 0.041 303 0.698 0.046 0.30 - - - 9,993 0.363 0.031 650 0.501 0.036 0.20 - - - 24,294 0.300 0.025 2,565 0.316 0.031 0.15 - - - 24,962 0.297 0.025 2,978 0.295 0.029 0.10 - - - 25,009 0.296 0.025 2,999 0.294 0.029 0.00 - - - 25,009 0.296 0.025 2,999 0.294 0.029
Table 16.4 Grade-tonnage Table for Sicera Norte
Measured Indicated Inferred Cut-off Grade Tonnage Cu Au Tonnage Cu Au Tonnage Cu Au (%) ('000 t) (%) (g/t) ('000 t) (%) (g/t) ('000 t) (%) (g/t) 0.80 - - - - 0.000 0.000 - 0.000 0.000 0.50 - - - - 0.000 0.000 13 0.526 0.013 0.40 - - - 252 0.434 0.018 36 0.490 0.013 0.30 - - - 5,825 0.333 0.021 824 0.333 0.018 0.20 - - - 23,346 0.251 0.017 4,339 0.246 0.018 0.15 - - - 27,415 0.236 0.016 5,240 0.230 0.017 0.10 - - - 35,064 0.214 0.015 6,494 0.211 0.016 0.00 - - - 35,556 0.212 0.015 7,047 0.199 0.015
PIT SLOPE ANGLES
Pit slope configurations were developed based on the results of stability analysis test work undertaken on the Main Zone Pit area. This analysis was provided in the report entitled “Zafranal Pit Slope Design, Conceptual Study” (Slope Design) dated May 30, 2011 and prepared by Knight Piésold. This report contains recommendations for the different pit domains, as detailed in Table 16.5 and Figure 16.1.
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Table 16.5 Inter-ramp and Overall Slopes Recommended for Zafranal Main Pit
Bench Total Inter-ramp Minimum Overall Design Pit Height Slope Slope Height Berm Width Slope Sector (m) (°) (°) (m) (m) (°) SD1 417 47 65 to 70 10.20 8.6 43 SD2 500 45 60 to 65 10.20 8.6 42 SD3 482 45 65 to 70 10.20 8,6 42 SD4 270 43 60 to 65 10.20 8.6 41 SD5 226 40 60 to 65 10.20 8.6 38 SD6 307 45 65 to 70 10.20 8.6 43
Figure 16.1 Inter-ramp and Overall Slope Angle by Domain
Given the irregular presence of weak rock mass in the deposit area, the maximum inter-ramp slope heights are limited to 200 m. It was assumed that 35 m wide haul ramps and/or step-outs will be incorporated into the pit walls to flatten the overall slopes.
16.3 P IT O PTIMIZATION
The optimized pit outlines for the four deposits were derived using Whittle™ Four-X optimization software. This analysis was carried out using the March 2012 resource block model, as well as technical and economic data recommended for the pit optimization by the QP and provided by AQM.
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The metallurgical analyis used in the Whittle™ evaluation takes into account two lines of processing; milling and flotation for sulphide-mineralized material including hypogene, supergene, and transitional and ROM leaching for all types of mineralized material, with the exception of hypogene, including oxide, leached, mixed, supergene, and transitional. Economic evaluation by block during the optimization process was modelled using the sulphide-mineralized material and oxide, leached and mixed materials were evaluated considering only the ROM leach process. Blocks with negative values were classified as waste.
The pit optimization runs considered the resource classification categories of Measured, Indicated, and Inferred.
16.3.1 TECHNICAL AND ECONOMIC PARAMETERS
The following base data and parameters were used in the pit optimization process.
MINING COSTS
The mining costs were estimated by NCL according to a preliminary study conducted in 2011. The costs of various unit operations are indicated in Table 16.6. The cost of $0.948/t moved was assigned to waste blocks located at the reference elevation of the mine exit to which was added $0.041/t for each uphill bench and $0.015/t for each downhill bench. For mineralized blocks, $0.337/t was added to the mining cost to represent the incremental cost of haulage to the primary crusher.
Table 16.6 Mining Costs
Cost/t ($) Comments Drilling 0.093 - Blasting 0.146 - Loading 0.205 - Hauling 0.291 Waste block at reference elevation Ancillary 0.185 - Support 0.028 - Total 0.948 Plus bench increment
SLOPE ANGLES
Global pit slope angles used for the Zafranal Main Zone correspond to those shown previously in Table 16.5. For the Victoria, Sicera Norte and Sicera Sur Zones, an average overall slope angle of 41.5° was used.
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MILLING/FLOTATION PROCESS
• Metallurgical Recovery: Copper recovery: 87% Gold recovery: 60% hypogene, 50% supergene, and 50% transitional • Process Cost: $6.33/t including transport • Selling Cost: Copper $0.36/lb Gold $6.0/oz.
ROM LEACH PROCESS
• Metallurgical Recovery: Copper recovery: 55% oxide, 55% leached, 55% mixed, 70% supergene, 50% transitional • Process Cost: $5.3/t including solvent extraction-electrowinning cost and transport.
METAL PRICES
• copper $2.70/lb • gold $1,100/oz.
ROYALTY
• 1.65% of the revenue.
16.3.2 PIT OPTIMIZATION OUTPUT
Based on the technical and economic data parameters described previously, a pit optimization run was performed independently for each block model using a series of metal prices. Table 16.7 to Table 16.10 lists the results, by respective mineralized zone.
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Table 16.7 Whittle™ Run Results – Zafranal Main Zone
Design Mineralized Cu Au Price Rock Material Strip Grade Grade ($/lb) ('000 t) ('000 t) Ratio (%) (g/t) 0.54 34,538 8,088 3.27 1.289 0.124 0.70 60,194 16,783 2.59 1.121 0.124 0.92 130,421 39,383 2.31 0.927 0.124 1.24 330,084 105,074 2.14 0.722 0.093 1.46 395,050 144,723 1.73 0.634 0.093 1.57 425,507 165,214 1.58 0.598 0.093 1.84 500,238 216,887 1.31 0.528 0.093 2.00 574,980 260,582 1.21 0.485 0.093 2.21 678,683 324,227 1.09 0.437 0.093 2.43 785,655 389,920 1.01 0.400 0.093 2.59 854,989 438,676 0.95 0.377 0.075 2.65 879,800 454,435 0.94 0.371 0.075 2.70 904,470 472,707 0.91 0.363 0.075 2.86 978,726 528,616 0.85 0.343 0.075 2.92 1,008,763 549,081 0.84 0.337 0.075 3.02 1,059,684 587,517 0.8 0.325 0.075 3.19 1,142,883 650,515 0.76 0.309 0.075 3.24 1,163,517 669,734 0.74 0.304 0.075 3.40 1,236,922 730,407 0.69 0.291 0.075
Table 16.8 Whittle™ Run Results - Victoria Zone
Design Mineralized Cu Au Price Rock Material Strip Grade Grade ($/lb) ('000 t) ('000 t) Ratio (%) (g/t) 1.57 4,693 994.0 3.72 0.548 0.044 1.73 6,356 1,522.0 3.18 0.510 0.040 1.89 9,349 2,558.0 2.65 0.453 0.037 2.00 11,589 3,452.0 2.36 0.420 0.037 2.16 16,126 5,375.0 2.00 0.373 0.037 2.21 19,702 6,640.0 1.97 0.356 0.034 2.38 32,214 10,885.0 1.96 0.325 0.031 2.48 39,578 13,952.0 1.84 0.308 0.031 2.54 46,384 16,656.0 1.78 0.297 0.031 2.65 57,546 20,826.0 1.76 0.283 0.031 2.70 60,093 22,104.0 1.72 0.279 0.031 2.81 71,863 26,879.0 1.67 0.266 0.031 2.97 98,223 35,749.0 1.75 0.250 0.031 3.02 98,499 36,481.0 1.70 0.248 0.031 table continues…
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Design Mineralized Cu Au Price Rock Material Strip Grade Grade ($/lb) ('000 t) ('000 t) Ratio (%) (g/t) 3.08 120,405 42,609.0 1.83 0.241 0.031 3.19 135,388 47,672.0 1.84 0.236 0.031 3.24 152,251 52,539.0 1.90 0.232 0.031 3.35 166,146 57,670.0 1.88 0.227 0.031 3.40 172,045 59,743.0 1.88 0.225 0.031
Table 16.9 Whittle™ Run Results – Sicera Norte Zone
Design Mineralized Cu Au Price Rock Material Strip Grade Grade ($/lb) ('000 t) ('000 t) Ratio (%) (g/t) 1.51 2 2 0 0.366 0.025 1.73 219 140 0.56 0.332 0.025 1.89 1,808 970 0.86 0.310 0.022 1.94 3,535 1,805 0.96 0.305 0.022 2.00 3,708 1,896 0.96 0.303 0.022 2.21 12,293 5,685 1.16 0.286 0.019 2.38 25,847 9,846 1.63 0.282 0.019 2.59 33,783 12,634 1.67 0.273 0.019 2.65 35,590 13,240 1.69 0.271 0.019 2.70 36,949 13,654 1.71 0.270 0.019 2.75 54,206 18,689 1.9 0.259 0.019 2.86 61,433 20,925 1.94 0.255 0.019 2.97 66,537 22,315 1.98 0.253 0.019 3.02 70,316 23,270 2.02 0.251 0.019 3.19 80,885 25,956 2.12 0.247 0.019 3.40 93,143 28,766 2.24 0.243 0.016
Table 16.10 Whittle™ Run Results – Sicera Sur Zone
Design Mineralized Cu Au Price Rock Material Strip Grade Grade ($/lb) ('000 t) ('000 t) Ratio (%) (g/t) 1.30 7 7 0.00 0.473 0.072 1.51 1,592 474 2.36 0.508 0.047 1.73 2,887 1,027 1.81 0.454 0.044 1.94 3,403 1,450 1.35 0.417 0.040 2.00 3,918 1,771 1.21 0.398 0.040 2.16 6,134 2,535 1.42 0.379 0.037 2.21 9,871 3,873 1.55 0.352 0.034 table continues…
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Design Mineralized Cu Au Price Rock Material Strip Grade Grade ($/lb) ('000 t) ('000 t) Ratio (%) (g/t) 2.38 13,595 4,673 1.91 0.350 0.034 2.48 15,802 5,131 2.08 0.348 0.034 2.59 19,517 5,877 2.32 0.343 0.031 2.70 23,965 7,052 2.40 0.329 0.031 2.81 27,108 7,764 2.49 0.324 0.031 2.86 28,865 8,060 2.58 0.323 0.031 3.02 30,890 8,447 2.66 0.320 0.031 3.08 34,363 8,866 2.88 0.320 0.031 3.19 38,393 9,473 3.05 0.317 0.031 3.29 39,306 9,667 3.07 0.316 0.031 3.46 41,435 9,958 3.16 0.314 0.028
16.3.3 MINING INVENTORY
Pit outlines corresponding to copper design price of $2.70/lb were selected for the design of the open pits. Table 16.11 to Table 16.14 shows the resource inventory by mineralization type and category for the selected pit outlines.
Table 16.11 Zafranal Main Zone Whittle™ Pit Mineral Resource
Tonnage Cu Au ('000 t) (%) (g/t) Measured Hypogene 60,898 0.265 0.068 Leached 6,793 0.254 0.130 Oxidized 2,443 0.260 0.144 Supergene 56,368 0.553 0.074 Measured Total 126,502 0.392 0.076 Indicated Hypogene 228,694 0.256 0.072 Leached 15,742 0.309 0.100 Oxidized 3,144 0.268 0.137 Supergene 98,014 0.589 0.075 Indicated Total 345,594 0.353 0.074 Measured + Indicated Hypogene 289,592 0.258 0.071 Leached 22,535 0.292 0.109 Oxidized 5,587 0.265 0.140 Supergene 154,382 0.576 0.075 Measured + Indicated Total 472,096 0.363 0.075 table continues…
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Tonnage Cu Au ('000 t) (%) (g/t) Inferred Hypogene 362 0.167 0.230 Leached 74 0.202 0.019 Oxidized 0 0.000 0.000 Supergene 175 0.288 0.026 Inferred Total 611 0.206 0.146
Table 16.12 Victoria Zone Whittle™ Pit Mineral Resource
Tonnage Cu Au ('000 t) (%) (g/t) Measured Hypogene 347 0.261 0.041 Leached 161 0.220 0.056 Mixed 3,821 0.318 0.029 Oxide 0 0.000 0.000 Supergene 2,972 0.400 0.025 Transitional 4,380 0.248 0.034 Measured Total 11,681 0.309 0.030 Indicated Hypogene 3,684 0.235 0.037 Leached 146 0.214 0.053 Mixed 877 0.247 0.022 Oxide 0 0.000 0.000 Supergene 93 0.734 0.044 Transitional 4,994 0.242 0.031 Indicated Total 9,794 0.244 0.033 Measured + Indicated Hypogene 4,301 0.237 0.038 Leached 307 0.217 0.055 Mixed 4,698 0.305 0.027 Oxide 0 0.000 0.000 Supergene 3,065 0.410 0.026 Transitional 9,374 0.245 0.033 Measured + Indicated Total 21,475 0.280 0.032 Inferred Hypogene - - - Leached - - - Mixed - - - Oxide 499 0.257 0.034 Supergene - - - Transitional 130 0.216 0.019 Inferred Total 629 0.248 0.031
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Table 16.13 Sicera Norte Zone Whittle™ Pit Mineral Resource
Tonnage Cu Au ('000 t) (%) (g/t) Indicated Hypogene 10,102 0.266 0.019 Mixed 0 0.000 0.000 Supergene 0 0.000 0.000 Transitional 1,638 0.305 0.012 Indicated Total 11,740 0.271 0.018 Inferred Hypogene 1,397 0.252 0.016 Mixed 0 0.000 0.000 Supergene 440 0.276 0.015 Transitional 77 0.355 0.015 Inferred Total 1,914 0.262 0.016
Table 16.14 Sicera Sur Zone Whittle™ Pit Mineral Resource
Tonnage Cu Au ('000 t) (%) (g/t) Indicated Hypogene 2,083 0.372 0.028 Leached - - - Mixed 205 0.295 0.022 Oxide 2,750 0.331 0.037 Supergene - - - Transitional 303 0.262 0.016 Indicated Total 5,341 0.342 0.032 Inferred Hypogene 98 0.452 0.056 Leached - - - Mixed - - - Oxide 1,158 0.259 0.031 Supergene 55 1.532 0.056 Transitional 400 0.173 0.016 Inferred Total 1,711 0.291 0.030
Table 16.15 shows the total mining inventory for the Project.
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Table 16.15 Total Mineral Inventory from Whittle™ Pits
Tonnage Cu Au Total In-pit Resources (t) (%) (g/t) Measured 138,183,000 0.385 0.07 Indicated 372,469,000 0.347 0.07 Measured + Indicated 510,652,000 0.358 0.07 Inferred 4,865,000 0.260 0.04 Waste 509,960,000 - - Total 1,025,477,000 - -
16.4 M INING P HASES
The optimized Whittle™ Pits at metal prices of $2.70/lb copper and $1,100/oz gold were selected as the ultimate pit outlines for the four deposits, but only the Main Zone and Victoria Zone were considered for the PEA. The Main Zone pit development considered six operative mining phases based on the sequence of Whittle™ Four-X nested pit shells, while only a single operative phase was designed for the Victoria pit. Sicera Norte and Sicera Sur deposits were not considered in the development scenario given the level of exploration and distance from the Main Zone. The optimized Whittle™ Pits for the Main Zone and Victoria were modified to incorporate pit access roads, safety berms and ensure adequate working areas. The mine design resulted in the configuration of pits and dumps as shown in Figure 16.2.
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Figure 16.2 General Mine Configuration
16.5 M INING S CHEDULE
The mining schedule was designed to reflect a concentrator feed rate of 80,000 t/d for 365 d/a.
16.5.1 PLANNING CRITERIA
The mine schedule was prepared according to the plant ramp-up projection, as listed below:
• Month 1: 36% throughput design • Month 2: 57% throughput design • Month 3: 63% throughput design • Month 4: 70% throughput design • Month 5: 80% throughput design • Month 6: 83% throughput design • Month 7: 87% throughput design • Month 8: 92% throughput design
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• Month 9: 96% throughput design • Month 10: 96% throughput design • Month 11: 97% throughput design • Month 12: 100% throughput design.
The sulphide-mineralized material with grade greater than the copper cut-off grade (COG) was sent to the flotation plant. In-pit marginal hypogene, with a copper grade greater than 0.15% and lower than the COG, was sent to a low-grade stockpile located near the crusher and reprocessed in the flotation plant at the end of the mine life. Oxide, leached, and marginal supergene, transitional and mixed mineralized material (copper grade greater than 0.15% and lower than the COG) were sent to a ROM leach pad.
Planning periods considered were as follows:
• start: January Year -2 • pre-stripping and production, Year -2: quarterly (Q3 Year -2 through Q4 Year 2) • from Year 3 to the end of mine life: yearly.
All available classified resources (Measured, Indicated and Inferred), were considered as plant feed material.
16.5.2 PRODUCTION PLAN
A mine production schedule was developed showing mineralized material tonnages, metal grades, and waste material by period, throughout the mine life. The distribution of mineralized material and waste contained in each of the mining phases was used to develop the schedule, thereby assuring that criteria for adequate mineral exposure, mining accessibility, and consistent material movement were maintained throughout the planning period.
Table 16.16 shows the 80,000 t/d mining plan summary for flotation feed by planning period and Table 16.17 shows the 80,000 t/d mining plan summary for leach feed and low grade flotation feed to stockpile by planning period.
The outputs of the proposed mining plans are shown in Figure 16.3 to Figure 16.5.
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Table 16.16 80,000 t/d Mining Plan Summary By Planning Period
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Table 16.17 80,000 t/d Mining Plan Summary By Planning Period
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Figure 16.3 Total Tonnage Mining Plan
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Figure 16.4 Concentrator Feed
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Figure 16.5 Leach Facility Feed
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16.6 M INE E QUIPMENT F LEET
The mine equipment fleet was selected on the basis of the schedule of material movement, the geometry of the planned open pits and the compatibility of individual equipment within the overall fleet. The following units were selected:
• main equipment electric drill diesel drill (311 mm diameter) rope shovel 3 front end loader (FEL) (17 m ) trucks (300 t) • ancillary equipment bulldozer 500 hp wheeldozer 500 hp grader 16 m 3 water truck 85 m .
The following mine operation characteristics were assumed for the fleet estimation process:
• 363 working days per year • 2 shifts per day • 12 hours total per shift
16.7 S CHEDULE OF M INE E QUIPMENT
Table 16.18 summarizes the total mine fleet requirements for the defined production plan.
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Table 16.18 Mining Equipment Requirements
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17.0 RECOVERY METHODS
17.1 I NTRODUCTION
The Zafranal concentrator has been designed to process a nominal 80,000 t/d, of copper-gold porphyry plant feed material from an open pit operation. The concentrator has been designed to produce a marketable copper concentrate of 28% copper containing about 3 g/t gold.
An SX-EW facility to recover copper from the dump leaching of low-grade oxide and sulphide material from the open pit will also be part of the overall processing plant.
17.2 S UMMARY
The metallurgical processing procedures selected for Zafranal were based on the results of metallurgical testing performed at G&T in Kamloops, BC, Canada and Plenge in Lima, Perú, along with the mining resources set out by NCL. The processing procedures selected for the design will produce a saleable, high-grade copper concentrate containing gold as a by-product.
The treatment plant will consist of a primary stage that includes crushing, SAG and ball mill grinding circuit, which is then followed by a flotation process to recover and upgrade copper from the feed material. As shown in the simplified flowsheet (Figure 17.1), the flotation concentrate will be thickened and filtered and sent to the concentrate stockpile for subsequent trucking and shipping to smelters.
The final flotation tailings will be disposed of using high density thickened slurry deposited into a TMF. This will allow for greater control of water management through recovery and re-use of water prior to tailings deposition. Process water will be recycled from the tailings thickener overflow and will be supplemented with process water recovered from the overflow of the concentrate thickener. Fresh water will primarily be used for gland service and reagent preparation.
The process plant will consist of the following unit operations and facilities:
• ROM plant feed material receiving and primary crushing • coarse feed material stockpile and reclaim • a SAG and ball mill grinding circuit incorporating cyclones for classification • SAG mill with potential for adding a pebble crushing circuit • copper rougher flotation
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• flotation concentrate regrinding • three-stage cleaner flotation • copper concentrate thickening, filtration, and dispatch • tailings thickening and disposal to a tailings storage.
The simplified flowsheet is shown in Figure 17.1.
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Figure 17.1 Simplified Process Flowsheet – Zafranal Project
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17.3 P LANT D ESIGN
17.3.1 MAJOR DESIGN CRITERIA
The major criteria used in the design of the concentrator are outlined in Table 17.1.
Table 17.1 Major Design Criteria
Criteria Unit Value Overall Plant Feed t/d 80,000 Operating Year d 365 Crushing Utilization % 55 Grinding and Flotation Utilization % 92 Bond Crushing Work Index, design kWh/t 7.85 Bond Rod Mill Work Index, design kWh/t 13.0 Bond Ball Mill Work Index, design kWh/t 12.1 Bond Abrasion Index, design g 0.280 Specific Gravity of Plant Feed Material - 2.70 Moisture Content of Plant Feed Material % 2.0 Primary Crushing Circuit Throughput Rate t/h 6061 SAG Mill Feed Size, 80% Passing µm 165,000 Pebble Crushing Circulating Load % 25.0 Ball Mill Feed Size, 80% Passing µm 2,000 Ball Mill Product Size, 80% Passing µm 150 Ball Mill Circulating Load % 250 Milling and Flotation Process Rate t/h 3,623 Pebble Crushing Rate t/h 906 Regrind Mill Product Size, 80% Passing µm 40 Head Grade, Design % Cu 0.837 Head Grade, Design Au, g/t 0.130 Copper Concentrate Grade % 28.0 Gold Grade in Copper Concentrate g/t 3 Anticipated Recovery, Design (Copper) Cu, % 87.7 Anticipated Recovery, Design (Gold) Au, % 49.0
The design parameters are based on test work results obtained by G&T and Plenge from the test programs performed 2012. Data from the SGS Lakefield Research Ltd. (SGS Lakefield) test work completed during 2010 was also incorporated, as was information provided by Transmin.
The grinding mills were sized based on the Bond Work Index data obtained from the test work. The regrind mill was sized in consultation with the vendor.
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The flotation cells were sized based on the optimum flotation times as determined during the laboratory test work. Typical flotation cell design parameters have been used in the design of the flotation circuit.
17.3.2 OPERATING SCHEDULE AND AVAILABILITY
The crushing and processing plants will be designed to operate on the basis of two 12-hour shifts per day, for 365 d/a.
The primary crusher availability will be 75% and utilization will be 55%. The grinding, and flotation circuit availabilities and running times will be 92%. This will allow for a potential increase in crushing rate, and will allow sufficient downtime for the scheduled and unscheduled maintenance of the crushing and process plant equipment.
17.4 P ROCESS P LANT D ESCRIPTION
17.4.1 PRIMARY CRUSHING
A single 1,524 mm by 2,870 mm gyratory crusher will be located in proximity to the open pit, approximately 700 m from the main pit exit. The crushing circuit will reduce the ROM plant feed material from a nominal top size of 1,000 mm to a crushed product size P80 of 165 mm in preparation for the grinding process. The primary crushing circuit will contain the following main items of equipment:
• ROM feed hopper/surge bin • coarse rock breaker • gyratory crusher • apron feeder • conveyor belts • belt scale • dust ventilation system.
Haulage trucks with nominal 300 t capacity will bring ROM material to the dry crushing plant. The material will be dumped directly from the trucks into the feed hopper for the crushing stage. The feed hopper will have an equivalent surge capacity of just over one truck-load. The area will be equipped with a coarse rock breaker in the event that oversize material makes its way into the feed hopper. The ROM material will exit the feed hopper and will feed the primary gyratory crusher. The gyratory crusher will operate at a utilization factor of 55% and will process 6,060 t/h of material. The gyratory crusher will reduce the plant feed material rock size from minus 1,000 mm to a product size P80 of about 165 mm, with the option of crushing finer to 125 mm. The crushed material will be deposited onto conveyor
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belts feeding the coarse material stockpile, located 250 m above and 1,125 m from the primary crusher.
The primary crushing and conveyor drop points will be equipped with dust ventilation systems to control fugitive dust that will be generated during crushing and conveyor loading, and the transportation of the crushed material.
17.4.2 COARSE MATERIAL STOCKPILE AND RECLAIM
The coarse material stockpile will be a production surge facility which will allow for a steady feed of material to the milling circuit.
The major equipment and facilities in this area includes:
• coarse material stockpile • reclaim apron feeders • a conveyor belt • belt scale.
The primary crusher will have reduced the ROM material to a P80 size of 165 mm, and this material will be conveyed to the coarse material stockpile.
The coarse material stockpile will have a live capacity of 54,000 t. The material will be reclaimed from this stockpile by three apron feeders with a combined nominal rate of 3,623 t/h. The apron feeders will feed a conveyor belt which will in turn will feed the SAG mill. The conveyor belt will be equipped with a belt scale.
17.4.3 GRINDING AND CLASSIFICATION
The concentrator will have a single grinding circuit comprised of a 40” diameter wrap around SAG mill and two 28” diameter wrap around ball mills. This will be designed as a two-stage operation with the SAG mill in closed circuit with a pebble crusher, and the two ball mills in closed circuit with the classifying cyclones. The SAG mill will be equipped with pebble ports to remove pebbles finer than 65 mm. The SAG mill circuit will not have pebble crushers installed initially. However, a pebble crushing circuit could be installed at later date. The grinding will be conducted as a wet process at a nominal rate of 3,623 t/h of material. The grinding circuit will include:
• a SAG mill • two ball mills • two pebble crushers (potentially added in the future) • a cyclone feed pumpbox • two sets of cyclone feed slurry pumps • two cyclone clusters
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• mass flow meter • cyclone overflow product particle size analyzer • sampler system.
The material from the coarse material stockpile will be reclaimed under controlled feed rate conditions using apron feeders. These feeders will discharge the material onto a conveyor belt feeding the SAG mill. A belt scale will control the feed to the SAG mill. Water will be added to the SAG mill feed material to assist the grinding process. The SAG mill will operate at a critical speed of 75%.
Should future grinding needs require pebble crushers, the SAG mill discharge end will be equipped with 65 mm pebble ports to remove the critical size material. The mill discharge will be screened by the mill trommel and the SAG mill discharge screen. The trommel and screen underflow will be discharged into the cyclone feed pumpbox. Each ball mill will be fed from an independent cyclone cluster and will be in closed circuit with the cyclone feed pumpbox. The product from each ball mill will be discharged into the cyclone feed pumpbox combining with the SAG mill discharge to become the cyclone feed. The slurry in the cyclone feed pumpbox will be split and the slurry will be pumped in two portions and will report separately to two cyclone clusters for classification. The cut size for the cyclones will be a P80 of 150 µm, and the circulating load to the ball mill circuits will be 250% with the cyclone underflow returning to the ball mills as feed material.
The new feed to each ball mill will be 1,812 t/h and the combined total of the two mills (3,623 t/h) will constitute the feed rate to the copper flotation circuit. The ball mills will operate at a critical speed of 75%. Dilution water will be added to the grinding circuit as required.
The cyclone overflow from both classification circuits will be discharged into the first copper rougher flotation cell in each rougher flotation line. The pulp density of the cyclone overflow slurry will be approximately 36% solids.
Provision will be made for the addition of lime to the cyclone feed pumpbox for the adjustment of the pH of the slurry in the grinding circuit prior to the flotation process.
Grinding media will be added to the mills in order to maintain the grinding efficiency. Steel balls will be periodically added to each mill using a ball charging kibble.
17.4.4 FLOTATION AND REGRIND CIRCUITS
The milled pulp will be subjected to flotation to recover the targeted minerals into a high-grade copper concentrate containing gold as a by-product. Tank style flotation cells will be used throughout the majority of the flotation circuit with final cleaning and upgrading of the concentrate occurring in column cells.
The flotation circuit will include the following equipment:
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• rougher flotation collection box • flotation reagent addition facilities • rougher flotation tank cells • one concentrate regrind mill, Isamill type • one classification cyclone cluster for the regrind mill • cleaner flotation collection box • first cleaner flotation tank cells • first cleaner scavenger flotation tank cells • second cleaner flotation tank cells • third cleaner flotation column cells • pumpboxes and standpipes • slurry and concentrate pumps • sump pumps • particle-size analyzer for the regrind mill • sampling system.
The two cyclone overflow streams from the grinding circuit will feed the rougher flotation circuit collection box by gravity flow from each of the ball mill grinding circuit classification cyclone clusters. The slurry will be monitored for P80 particle size, and flotation feed samples will be taken periodically for process control and metallurgical accounting.
The collection box will mix the two product streams in preparation for sampling at the design feed rate of 3,623 t/h. Flotation reagents will be added to the collection box as defined through testing. The flotation reagents added will be the collectors, PAX, the xanthate allyl ester Aero 3302 (A3302), and the dithiophosphinate Aerophine 3418A (3418A) together with the frother, MIBC. Provision will be made for the addition of collector reagents to the grinding circuit, as well as the staged addition of the reagents in the rougher circuit and in the cleaner stages of the flotation circuit.
The rougher flotation circuit will consist of two flotation trains which will operate in the following manner. The two cyclone overflow slurry streams will flow into the collection box; the combined slurry stream will be sampled and will then be split into two streams in the distribution box with each stream feeding one of the rougher flotation tank cell lines. Air injection will facilitate the flotation process. The copper minerals will be selectively floated into a rougher concentrate away from the other minerals and the gangue present in the slurry. The rougher concentrate will constitute approximately 11.0% mass of the plant feed. Each rougher tailings stream will be sampled automatically prior to discharge into the rougher tailings pumpbox for
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process control and metallurgical accounting purposes. These streams will be discharged to the tailings thickener.
The rougher concentrate streams will be combined and form a single feed which will continue to the regrind and cleaner flotation sections of the flotation circuit for further processing.
To completely liberate the fine-sized grains of the copper minerals from the gangue constituents and to enhance upgrading of the copper concentrate, regrinding and staged cleaning will be incorporated in the cleaner flotation circuit. A single stage of regrinding and three stages of flotation cleaning have been chosen so that a final product of acceptable grade and recovery will be achieved.
The rougher concentrate will be pumped to the regrind densification cyclone cluster. The regrind cyclone cluster will operate as densification cyclones and will deliver the feed to the regrind Isamill at the required density of 50% solids.
The Isamill product will discharge the finely milled material into the cleaner collection box where it will be recombined with the regrind densification cyclone overflow, other product streams and reagents, prior to being fed into to the head of the first cleaner flotation cells. The rougher concentrate regrind Isamill will have a design treatment rate of about 200 t/h, and will have a circulating load of approximately 10% as specified by the vendor.
The regrind circuit discharge will be combined with the second cleaner tailings and the first cleaner scavenger concentrate and this will constitute the feed to the first cleaner flotation stage. This first cleaner stage will be directly followed by the first cleaner scavenger flotation stage. Tailings from the first cleaner scavenger flotation stage will be directed to the tailings thickener.
The first cleaner concentrate will be combined with the third cleaner tailings as the feed to the second cleaner flotation stage. The second cleaner concentrate will report to the third cleaner flotation stage as flotation feed. The concentrate from the third cleaner flotation stage will be the final copper concentrate with a design copper grade of 28% copper. The third cleaner copper concentrate product will feed directly to the copper concentrate thickener for dewatering. The second cleaner tailings will be returned to the cleaner collection box. The third cleaner tailings will be the feed to the second cleaner stage together with the first cleaner concentrate. Provision will be made for the copper concentrate thickener overflow water to be re-used in the grinding and flotation circuit as process water providing this does not have a deleterious effect on the flotation of the copper and gold minerals as a result of residual reagents.
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17.4.5 CONCENTRATE HANDLING
The cleaner flotation concentrate will be thickened, filtered, and stored prior to shipment to the smelter. The concentrate handling circuit will have the following equipment:
• concentrate thickener • concentrate thickener underflow slurry pump • concentrate filter feed tanks • concentrate filter press feed pumps • concentrate filter press • filter press washing and filtrate handling equipment • dewatered concentrate storage and dispatch facility.
The concentrate produced will be pumped from the final cleaner flotation stage to the concentrate thickener. Flocculant will be added to the thickener feed well to aid the settling process. The thickened concentrate will be pumped to the concentrate filter feed tanks using the thickener underflow slurry pump. The underflow density will be 60% solids. The concentrate filter feed tanks will be agitated tanks that will serve as the feed surge for the two concentrate filters. The concentrate filters will be pressurized filter press units. Since filtration will be a batch process, the concentrate filter feed tanks will also act as surge tanks for the filtration operation. The filter presses will dewater the concentrate to produce a final concentrate with a moisture content of about 9%. The filtrate will be returned to the concentrate thickener. The filter press solids will be discharged directly onto the concentrate stockpile. The dewatered concentrate will be stored in a designated storage facility and will periodically be loaded into trucks for dispatch off the property.
The thickener overflow solution from the concentrate thickener will be collected in the process water tank for recycling within the mill circuit.
17.4.6 TAILINGS THICKENER
The plant tailings will be thickened to maximize the recovery of water prior to final deposition. The tailings circuit will include the following equipment:
• tailings thickener feed tank • thickener unit • final tailings sampler.
The rougher flotation tailings, combined with the first cleaner scavenger tailings, will be the final plant tailings.
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Flocculant reagent will be added to aid the settling of the tailings in the thickener. The flocculant will be added as a dilute solution at an addition rate which will be determined during future testing.
The thickener overflow solution will be directed to the process water tank for re-use in the plant as general process water. The solids will be thickened to a pulp density of 65% solids and will then be gravity fed to a tailings pond facility for final deposition.
17.4.7 REAGENT HANDLING AND STORAGE
The various chemical reagents will be added to the process slurry streams to facilitate the recovery of copper during the flotation process. The preparation of the various reagents will require:
• a bulk handling system • mix and holding tanks • metering pumps • a flocculant preparation facility • a lime slaking and distribution facility • eye-wash and safety showers • applicable safety equipment.
Various chemical reagents will be added to the grinding and flotation circuit to modify the mineral particle surfaces and enhance the flotability of the valuable mineral particles into the copper-gold concentrate product. The reagents will be prepared reasonably close to the point of usage. The reagent preparation section will be under a roof to protect the reagents and equipment from the environment. Most reagents will be received in bulk in palletized bags, chemtainers, drums or bulk bags. The reagent preparation section will contain strategically located safety showers and eyewash stations. Each reagent preparation area will be bunded to contain any spillage which may arise during the preparation stage with each bunded area served by a sump pump for the cleaning up and control of any spillage arising. Fresh water will be used in the making up or the dilution of the various reagents that will be supplied in powder/solids form, or which require dilution prior to the addition to the slurry. These solutions will be added to the addition points of the various flotation circuits and streams using metering pumps.
At present, test work is ongoing in order to specifically identify the collector reagents which will be used in the flotation plant. The three flotation collector reagents which have been tested are PAX, A3302, and 3418A. All three of these reagents will be mentioned in the report, although it is probable that only two of these will be used operationally at any time.
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LIME
The design is based on using hydrated lime as the pH modifier for the various unit processes. Hydrated lime will be delivered in 40 t trucks, and will be added to the grinding and flotation circuits. The hydrated lime will be off-loaded into a storage silo. A screw conveyor will add the required amount of hydrated lime to the hydrated lime tank. The hydrated lime slurry strength will be 20%. The hydrated lime will then be distributed to the addition points via a closed loop piping system.
POTASSIUM AMYL XANTHATE
PAX will be used in the flotation circuit as the collector reagent for the flotation of copper and gold minerals. The PAX will arrive at the plant in bags. The reagent will be made up in a mixing tank to the required solution strength of 20%, and then transferred to the holding tank, from where the solution will be pumped to the addition points in the circuit using metering pumps.
AERO 3302
The A3302 will be used in the flotation circuit as the collector reagent for the copper minerals and the gold. The A3302 will arrive in drums and will be pumped to points of addition as a 100% solution strength. Since this reagent is not soluble in water, the A3302 will likely be added to the grinding circuit. The A3302 reagent belongs to the xanthate family of reagents, and is a xanthate allyl ester.
AEROPHINE 3418A
The 3418A will be used in the flotation circuit as the collector reagent for the flotation of copper and gold minerals. The 3418A will arrive in drums and will be transferred to a holding tank prior to distribution within the plant. The 3418A will be pumped directly to the flotation circuit as a 100% concentration strength solution using metering pumps. The 3418A reagent is a dithiophosphinate-type collector.
METHYL ISOBUTYL CARBINOL
The reagent MIBC will be used in the flotation circuit to provide the froth phase of the flotation process. MIBC will arrive in bulk containers and will be transferred to a holding tank prior to distribution within the plant. MIBC will be pumped directly to the flotation circuit as a 100% concentration strength solution using metering pumps.
FLOCCULANT
Flocculant will be used in the flotation concentrate thickener and tailings thickener as an aid in the settling process. The flocculant will be prepared at the required concentration in a proprietary vendor-supplied flocculant preparation facility. Flocculant will be delivered in bulk bags. A screw conveyor will deliver the correct amount of dry flocculant powder to be mixed with water prior to delivery into the flocculant mix tank. The flocculant will be allowed to hydrate in the mix tank before
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being transferred to the holding tank where it will be made up to the required dosing strength. A metering pump will transfer the required amount of flocculant from the holding tank to the points of addition at the flotation concentrate thickener and the tailings thickener where final dilution will occur.
Flocculant will be prepared in the standard manner as a dilute solution with 0.20% solution strength. This will be further diluted in the thickener feed well prior to mixing with the slurry.
17.4.8 ASSAY AND METALLURGICAL LABORATORY
The assay laboratory will be equipped with the necessary analytical instruments to provide all routine assays for the mine, the concentrator, and the environment departments. The most important of these instruments includes:
• an AAS • an inductively coupled plasma-spectrometer (ICP-MS) • a Leco furnace, or equivalent.
The metallurgical laboratory will undertake all the necessary test work to monitor metallurgical performance and, more importantly, to improve process flowsheet unit operations and efficiencies. The laboratory will be equipped with laboratory crushers, ball and stirred mills, particle size analysis sieves, flotation cells, filtering devices, balances, and pH meters.
17.4.9 WATER SUPPLY
Two separate water supply systems, one for fresh water and one for process water, will be provided to support the process operations.
FRESH WATER SUPPLY SYSTEM
Fresh and potable water will be supplied to a fresh/fire water storage tank which will be delivered by pipeline as desalinated water from a desalination plant situated on the coast. Fresh water will be used primarily for the following:
• fire water for emergency use • cooling water for mill motors and mill lubrication systems • gland service for the slurry pumps • reagent make-up • potable water supply.
The potable water from the fresh water source will be treated and stored in the potable water storage tank prior to delivery to the various service points.
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PROCESS WATER SUPPLY SYSTEM
Some process water generated in the flotation circuit as concentrate thickener overflow solution will be re-used in the process circuit via the process water tank. Reclaimed water will also be pumped from the tailings thickener overflow to the process water tank for distribution to the points of usage.
17.4.10 AIR SUPPLY
Separate air service systems will supply air to the following areas:
• low-pressure air for flotation cells will be provided by air blowers • high-pressure air for the concentrate filter press operation will be provided by dedicated air compressors • instrument air will come from the plant air compressors and will be dried and stored in a dedicated air receiver.
17.4.11 ON-LINE SAMPLE ANALYSIS
The plant will rely on the on-stream, or in-stream, analyzer for process control. The analyzer will determine the metal content in the various flotation streams in the circuit. A sufficient number of samples will be taken so that the circuit can be balanced by analytical resultant and calculation as required. Specific samples that will also be taken for metallurgical accounting purposes will be the flotation feed to the circuit, the final tailings, and the final concentrate produced. These samples will be collected on a shift-basis and will be assayed in the assay laboratory. On-stream particle size monitors will determine the P80 particle size of the primary cyclone overflow and the regrind circuit products.
17.5 S OLVENT E XTRACTION-E LECTROWINNING P ROCESSING F ACILITY
17.5.1 INTRODUCTION
The heap leach SX-EW plant has been designed to produce a nominal 10,000 t/a of high quality copper cathodes from oxide and secondary sulphide plant feed material with copper grade in the range of 0.15 to 0.5%, with an average grade of 0.23% total copper. The plant feed material will be processed using acid solution dump leaching and SX, and the copper will be recovered using an EW operations process.
17.5.2 SUMMARY
The treatment plant will consist of the initial curing by acid impregnation of the leach piles which will be followed by the dump leaching stage consisting of 94 modules, with the resulting pregnant leach solution (PLS) being pumped to the SX circuit for copper recovery. The high concentration copper-rich electrolyte will be processed in the EW circuit. Two rows of cells will produce high quality copper cathodes. The
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low-cost ROM dump leach process will not require any crushing; rather the dump piles will be formed by the direct dumping of the ROM material from trucks on the leach pads.
The process plant will consist of the following unit operations and facilities:
• dump leaching pad • solution ponds • SX circuit • EW circuit • water supply facilities, reverse osmosis (RO) plant.
17.5.3 PLANT DESIGN
MAJOR DESIGN CRITERIA
The leach SX-EW plant has been design to produce 10,000 t/a of high quality copper cathodes. The major plant design criteria are included in Table 17.2.
Table 17.2 Major Design Criteria for the Leach SX-EW Plant
Criteria Unit Value Monthly Production Capacity t copper 840 Operating Year d/a 360 Oversize Factor 1.09 Plant Availability % 95.0 Average Plant Feed to Leach Pads t/d 19,853 Head Grade, Design % Cu 0.15 to 0.50 Particle Size of Plant Feed to Leach Pads mm 100% < 13 Bulk Density of Plant Feed t/m3 1.60 Dumping/Lift Height m 6.00 Copper Recovery % 60 Instantaneous Irrigation Rate L/h/m2 14.8 Total Leaching Period d 541 PLS to SX Concentration g/L Cu 2.76 Raffinate from SX Concentration g/L Cu 0.28 PLS to SX Flow Rate m³/h 431 Raffinate from SX Flow Rate m³/h 433 Organic Flow Rate m³/h 431 Charged Electrolyte Flow Rate m³/h 97.8 Number of O/A Rate Extraction Stages - 1.00 Number of O/A Rate Re-extraction Stages - 4.41 Extractant Concentration % v/v 11.7 table continues…
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Criteria Unit Value Efficiencies of SX Stages % 94.0 Number of Extraction Stages - 2.00 Number of Re-extraction Stages - 1.00 Number of Cells for EW - 62.0 Current Efficiency % 92.0 Number of Rectifiers - 2.00 Design Current A 17,601 Discharged Electrolyte from EW Concentration g/L Cu 95.5 Note: O/A = organic/aqueous, v/v = volume to volume
17.6 P ROCESS P LANT D ESCRIPTION
17.6.1 DUMP LEACHING
Fresh leach feed material from the mine will be delivered to the leach pads by dump trucks. The total leaching period is 541 days, with an intermediate leach solution (ILS) spray period of 506 days and a raffinate and water spray period of 34.5 days.
The surface of the heap will be ripped between the lifts to re-establish porosity. The plant feed will be piled up to a height of 6 m and will be irrigated by modules. Each module will hold 94,159 t with a surface area of 9,808 m2. A total of 94 modules will be in operation. In the first step, namely the curing stage, the leach plant feed will be covered with a dilute sulphuric acid solution (10 kg/t strength) as a pre-treatment. It will then be sprayed with previously re-acidified raffinate ILS from the PLS recirculation pond at a flow rate of 6,981 m3/h, and will then be sprayed with previously re-acidified raffinate at a flow rate of 3,855 m3/h. The leached tailings heap will then be washed with water from the water treatment facility at the flow rate of 91 m3/h.
The dump leaching facility consists of the following units:
• a curing solution pond which will receive raffinate solutions (raffinate and discharged electrolyte) and mix these solutions • a PLS recirculation pond consisting of two chambers to have independent acid dosage in each pond in order to allow irrigation of leach pads with different acid strengths • a PLS to SX pond consisting of two chambers of sufficient capacity for the inlet chamber to ensure a guaranteed minimum final sedimentation time for the solids • a raffinate pond with a truncated pyramid pond coated with plastic sheet where raffinate will be collected and re-acidified
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• an emergency pond sized for the worst case scenario, namely the coincidence of maximum rainfall, an intermediate dump level with the plant feed undergoing treatment, and with abandoned leached tailings.
17.6.2 SOLVENT EXTRACTION
A two-stage PLS pond has been incorporated into the process design to assist with the settling of any fine colloidal particles before these reach the SX facility.
The solvent extraction circuit will consist of four stages of mixer-settlers with a highly flexible hydraulic system to accommodate the necessary configurations and to achieve high overall system availability. It consists of a pond of loaded organic solution and a loaded organic washing stage. The net efficiency of the mixing stage will be 94% and will operate 24 h/d.
The mixer-settlers will be arranged in a multi-stage arrangement with process pipelines lying beneath the passageways and at base level of the mixers. The settling tanks will be uncovered, and each settling tank will be provided with a distribution board and picket fences to coalesce the primary streams, a system to refill the organic and aqueous phases, piping to facilitate recirculation of organic and aqueous phases in the extraction stage and the aqueous phase in the re-extraction. The instrumentation and means of operational adjustment of the recirculation and the thickness of the organic layer, together with a system for removal of the sediments, will also be provided.
The basic design consists of four mixer-settler equipment units for all the corresponding streams for an overall feed rate of 431 m3/h at a concentration of 2.76 g/L copper. In its basic configuration, each train consists of two extraction stages with a O/A rate of 1, and one organic washing stage and one re-extraction stage with a O/A rate of 4.41; however, these trains can be switched to alternative configurations in order to meet fault conditions and maintenance that require replacement or removal of some of the service equipment and re-configuration of hydraulic lines.
A loaded organic tank and post-settlers for all aqueous phases as well as a retention filter for organic in the rich electrolyte are included to ensure low organic consumption. The organic used in extraction will be a mixture of extractant and diluent. The extractant will be a Ceto-oxima reagent used at 11.7% v/v in a high grade kerosene diluent. The purpose of the diluent will be to reduce the viscosity of the mixture with the extractant to practical levels.
The SX plant facilities will have a sediment treatment plant to maintain the circulating organic within specifications in order to prevent the formation of crud. The same installation allows the treatment and removal of crud should it form in the organic. The operation of this plant is completely manual and directly controlled by the operators.
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A raffinate pond has been included with a design flow rate of 433 m3/h. Raffinate will be pumped from the pond to the dump leaching facility.
The plant’s design envisages the installation of an electrolyte post-settler and a filter system to recover the organic entrained by the re-extraction of charged electrolyte. The separator will be installed immediately downstream of the stage R-1 settlers (also connected to L-1 stages in alternative configurations), thus minimizing the difference in elevation and the distance between the spillway of aqueous phase and the post-settler. The Terral filter system recovers the organic through high-density activated carbon beds, from which the retained organic is removed by flotation in low-pressure inducted air. It is fed by gravity from the post-settler then discharged to the retention tank for delivery to the EW tank house.
17.6.3 ELECTROWINNING CIRCUIT
The copper EW tankhouse includes two rows of electrolytic cells fed with rich and circulating electrolyte. Only one of the lines includes cleaning cells that are fed with a high concentration rich electrolyte while the balance of the cells in the other row are fed with circulating electrolyte.
The copper deposition is accomplished by passing current through the cells via an arrangement of installed anodes and cathodes. This current is controlled by an external rectifier which provides the necessary voltage for current to flow in series through the cell arrangement. The current is transferred from one cell to another by intermediate walls in contact, which simultaneously connect the anodes of one cell and with the cathodes of the next cell.
Each electrode will have a submerged area of 1 m2. An average current density of 275 A/m2 (maximum 300 A/m2) is employed, which should be achievable in the designed operating environment. This current density results in a requirement for 64 cells of 33 cathodes each, 12 of which are cleaning cells. A design current efficiency of 92%, yields a design current of 17,600 A divided between two rectifier units. The total design cathode copper production is 10,080 t/a, based on 28,000 t/d and an operating rate of 360 d/a.
The discharged electrolyte is sent to the SX plant for upgrading to rich electrolyte after passing through a reductive column, or section of the electrolyte tank filled with copper scrap to reduce the high redox potential of the discharged electrolyte, which left untreated could eventually degrade the organic in contact at stage R-1.
17.7 W ATER S UPPLY – R EVERSE O SMOSIS P LANT
Industrial water will be pumped to the RO plant. The reject water will be used to feed the raffinate ponds and to wash the leached tailings. The clean water will be used in processes including the washing of organic and cathodes. Water purification equipment will also convert some of the clean water into potable water for human consumption.
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18.0 PROJECT INFRASTRUCTURE
18.1 I NTRODUCTION
As a comprehensive greenfield project, the Project will require the development of supporting infrastructure.
These items include:
• a process plant or concentrator that will include crushing, grinding, flotation, regrinding, concentrate filtration, concentrate thickener, tailings thickener facilities and assay laboratory • a warehouse, maintenance shop, administration offices, and supporting infrastructure • a permanent operations camp • a geomembrane-lined TMF, and two structural earth dams, two waste rock dumps and a low grade sulphide stockpile • a SX-EW facility and heap leach pads • a network of access and on-site roads • a fresh water supply and distribution system, including the option of a seawater desalination plant • power supply and distribution, including a 220 kV power transmission line from Arequipa, a 220 kV substation at the plant site, and 66 kV and 25 kV power distribution lines • other infrastructure including truck shop, temporary construction camp, and laydown areas.
The overall site layout is presented in Figure 18.1. The detailed site layout is presented in Figure 18.2.
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Figure 18.1 Overall Site Layout
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Figure 18.2 Site Layout
Note: Coordinate Reference System WGS-84, UTM Zone 18
As shown in Figure 18.3, the Project is located in rugged and mountainous terrain, surrounded by ridges, relatively steep slopes, gullies, and canyons. There are two large, terraced alluvial plains on the Property. The locations of the alluvial plains are controlled by narrow, steep walled canyons.
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Figure 18.3 Site View from Concentrator Location
Site elevation ranges from about 1,400 m at the base of the tailings embankments on the alluvial plains to about 2,800 m near the Main Zone pit. The process plant, the dump leach pad, the truck shop, the concentrator/administration/maintenance shop facility, and the waste dumps will be located adjacent to the open pit. The TMF and associated infrastructure will be located approximately 10 km due south in a natural basin, as shown in Figure 18.2. Mineralized material will be crushed and processed at the process plant in order to produce a final concentrate. Tailings will be delivered via pipeline to the TMF from the process plant.
The desalination plant will be located on the Pacific Coast, approximately 115 km from the process plant, and pipelines and pumping stations will be used to bring the water to site.
18.2 P ROCESS P LANT AND A NCILLARY F ACILITIES
The overall process plant facility will be 230 m long by 90 m wide and house the crushing, grinding, flotation, regrinding, filtration and concentrate thickener areas. Equipment will be supported by concrete foundations and structural steel and accessed with ladders or stairs with access/maintenance platforms where required. Heavy or large equipment will be serviced by overhead cranes.
Figure 18.4 illustrates the plant site layout.
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Figure 18.4 Plant Site Layout
18.2.1 FACILITIES DESCRIPTIONS
CRUSHING FACILITY
The crushing building will be made of concrete construction, with multiple levels, housing the primary crusher and apron feeder. The structure will utilize mechanically stabilized earth (MSE) walls on three sides, stick-built steel framing, and will be enclosed up to the dump pocket. ROM material will be reduced in size, then discharged into the dump pocket on the upper level of the facility for transport to the grinding process. Interior steel platforms will be provided to support equipment for ongoing operations and maintenance.
The area will also be equipped with a dust control system, to collect the fugitive dust.
COARSE MATERIAL STOCKPILE AND RECLAIM
This coarse material stockpile will be a production surge facility that will have a live capacity of 54,000 t. Feed will be reclaimed form the stockpile by three apron feeders.
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GRINDING FACILITY
The grinding facility will consist of a SAG/ball mill combination circuit in a two-stage operation. The 40 ft by 22 ft SAG mill will be designed to operate in close-circuit with two pebble crushers but will initially operate without the crushers and in open-circuit with two 28 ft by 42 ft ball mills. The SAG mill will be equipped with pebble ports. The grinding will be conducted as a wet process.
This facility will be housed in an open-air structure and serviced by an overhead crane, with the equipment installed in a bunded area with sloped concrete floor feeding to a sump pump.
FLOTATION FACILITY
The flotation plant will include 12 rougher flotation cells having a cell size of 300 m3 with 300 kW each of installed power that will be used to recover the targeted minerals into a high-grade copper concentrate containing by-product gold. The cells will be equipped with particle-size analyzers. They will be mounted on a bunded, sloped concrete floor feeding to a sump pump for recovery, in case of spillage. The building will be open-air and serviced by a 100 ft tower crane.
The regrind facility will consist of a single IsaMill, combined with a hydrocyclone package which will also be mounted upon a bunded, sloped concrete floor feeding to a sump pump in an open-air setting and serviced by an overhead crane.
CONCENTRATE THICKENER
The concentrate thickener will have a diameter of 27 m (88.6 ft). The thickener will be set upon a bunded, sloped concrete floor to sump pump. This facility will be open to atmosphere and will contain a dry chemical makedown system required for the production of flocculant used in the thickening process.
CONCENTRATE FILTRATION FACILITY
Two plate-and-frame filter presses will be mounted on an elevated steel structure that will also consist of a concentrate load-out facility.
TAILINGS THICKENER
The single high-rate tailings thickener will have a diameter of approximately 80 m and will be supported on a 86 m diameter mat foundation and steel support as required. The thickener will be installed upon a open-air, bunded and sloped concrete pad to sump pump
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ASSAY LABORATORY
An assay laboratory will be equipped with the necessary analytical instruments to provide all routine assays for the mine, the concentrator, and the environment departments. The most important of these instruments includes:
• an AAS • an ICP-MS • a Leco furnace, or equivalent • fire assay furnaces.
The laboratory will be equipped with laboratory crushers, ball and stirred mills, particle size analysis sieves, flotation cells, filtering devices, balances, and pH meters.
18.3 W AREHOUSE, M AINTENANCE, AND C ONCENTRATE O PERATIONS F ACILITIES
18.3.1 MAINTENANCE SHOP AND WAREHOUSE
The maintenance shop and warehouse will be a 40 m long by 25 m wide enclosed steel structure complete with concrete foundations. An overhead crane is required to service the maintenance area.
This building will house a parts warehouse and some offices for maintenance staff. It will also have a 25 m by 25 m fenced storage compound adjacent.
18.3.2 CONCENTRATOR OPERATING FACILITY
The concentrator operating facility will be a 30 m long by 25 m wide enclosed steel structure complete with concrete foundations. This facility will house offices for 30 people, and include a reception area, a mill dry for 100 people and all required amenities.
18.3.3 FUEL STORAGE AND DISTRIBUTION
Diesel fuel for the mining equipment and ancillary facilities will be supplied from storage tanks, located near the haul road to the crusher. The diesel fuel storage tanks will have a capacity sufficient for approximately three days of operations. Diesel storage will consist of aboveground tanks and a containment pad, complete with loading and dispensing equipment conforming to regulations.
18.4 T AILINGS M ANAGEMENT F ACILITY
The TMF will consist of a tailings delivery pipeline that will route thickened slurry by pump and high-density polyethylene (HDPE) pipelines for storage behind two
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structural earth dams. The starter embankments will be lined with geomembrane. Conceptual level design for the tailings and water management components are based on an estimated total of 425 Mt of processed mineralized material. The TMF will be located approximately 10 km south of the plant site and a total of 17 Mm3 of borrow material will be required for embankment construction. Seepage recovery will include toe and embankment drains to allow the capture and collection of seepage in seepage collection pond. It is anticipated that some process water and runoff will be impounded, but much of this water will be absorbed into the beaches and lost to evaporation. Given the low levels of rainfall and high evaporation rates in the region of the Project, water reclamation was not considered for the Project.
The conceptual level study developed by Knight Piésold for tailings and water management includes the following components:
• development of the design basis and operating criteria for the TMF, including confirmation of the hazard potential classification • seismic risk assessment to determine the extreme seismic events that could occur in the region • review of available climate and hydrology data to determine appropriate storm runoff events for input into the design of the various facilities • TMF volumetric (depth-area-capacity) calculations, embankment design and construction staging • construction material descriptions from local borrow sources • seepage control and collection measures, including grout curtains and geomembrane liners • tailings delivery layouts and assessments of pumping requirements and tailings distribution methods • seepage recovery assessments and development of a water balance for the TMF • conceptual (scoping) level cost estimates for the TMF (capital and operating cost estimates).
The geotechnical conditions at the TMF are based on general knowledge of the area and no specific geotechnical investigations were completed.
18.5 L EACH F ACILITY
The leach facility is comprised of lined leach pads, collection ponds and tanks, solvent extraction building, electrowinning building, water treatment building, a small maintenance shop and an operations control centre adjacent to the electrowinning building. The leach facility will be separate from the concentrator and most of its industrial buildings will be constructed of engineered post-and-beam steel structure with a steel roof and no walls. The foundations will consist of concrete spread
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footings, grade walls along the structure perimeters, a slab-on-grade floor, and containment curbing. The floor surfaces will be acid protected, and will have localized areas that are sloped toward sump pumps for cleanup operations.
The storage ponds for curing solutions, enriched solutions for irrigation recirculation, enriched feed solution for the solvent extraction plant, raffinate, and emergency ponds will be constructed in the form of truncated pyramids and lined with polymer concrete or with welded plastic sheets. These ponds will provide regulation capacity between the leaching and solvent extraction operations. The emergency pond accommodates overflows from the leach pads generated mainly by pumping suspension due to power failure and rainwater catchment.
Suitable material will be transported to the leach pad in large mining trucks that unload directly onto the permanent lined pad.
Mineralized material is continuously heaped for irrigation in modules whose areas are defined by the planned irrigation system although only sufficient area is irrigated at any one time to ensure the production of a nominal 10,000 t/a of cathode.
The leach pad will eventually cover an area approximately 1,500 m by 500 m, and reach 100 m in height for a final capacity of 87 Mt.
18.6 R OADS
18.6.1 OVERVIEW
Approximately 372 km of roads will need to be constructed or upgraded, including:
• 65 km from the plant site to Pedregal de Majes (main access road) • 1 km access road to the explosives preparation plant • 38 km from Corire to the plant site • 6 km from the Corire road to the TMF • 8 km from the TMF to connect with the road to Pedregal de Majes • 6 km maintenance road for TMF piping and embankments • 13 km of external pit haul roads • 115 km maintenance road for water supply power line and pipeline from the coast • 120 km maintenance road for the main power line from Socabaya Substation near Arequipa.
The main access road will be paved and connect the plant site with the Pan American Highway and will be used for transporting personnel, supplies and product to the Matarani Port. The Corire road will be gravel/clay-topped and primarily used
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for transporting employees that reside in the area to the mine site, as well as supplying locally grown produce to the camps. The tailings line will follow the Corire road for approximately 4 km from the plant site, then branch off on a 6 km gravel road to the TMF. An 8 km gravel road will connect the TMF to the main access road. A 1 km gravel branch road off the main access road will access the explosives plant.
Additionally, approximately 13 km of external pit haul roads will be constructed during the LOM.
A total of 6 km of maintenance roads will service the TMF and another 241 km of maintenance roads will be constructed on- and off-site to service the water supply pipeline and power line and the power line from the Socabaya Substation.
18.6.2 EXISTING ACCESS ROADS
Existing gravel-topped access roads join the Project site with the nearby communities of Pedregal de Majes and Corire. Approximately 39 km of the proposed main access road to Pedregal de Majes exists and 5 km has been paved by the government. The 34 km unpaved portion will require upgrading of its structural base and paved with a chip seal topping. A road from Corire to CMZ´s second exploration camp at Ganchos exists but the Project contemplates constructing a new gravel/clay- topped road in an alternative river canyon that will connect Corire with the proposed plant site.
18.6.3 NEW ACCESS ROADS
A total of 26 km of the proposed main access road will need to be constructed and paved with chip seal topping. The Corire road will also require 38 km of new construction and topping with gravel/clay material. The 6 km branch road to the TMF as well as the 8 km connection to the main access road will also be constructed and topped with gravel.
18.6.4 SITE ROADS
The 13 km external pit haul roads will be developed over the mine life, with surface treatment for dust suppression.
A total of 241 km of new gravel roads will be developed for maintenance on and off site including:
• 6 km around the TMF for piping and embankments • 115 km for power lines and water pump stations. • 120 km for main power line from Socabaya Substation.
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18.7 F RESH W ATER S UPPLY AND D ISTRIBUTION
AQM is actively talking with authorities and local communities to find alternative sources of water that would provide benefits to some local communities in exchange for the use of local water sources by the Project. For this PEA, the fresh water supply for the Project is assumed to come from a desalination plant on the Pacific Coast, approximately 115 km from the proposed plant site. The water will be produced, as required, for drinking and industrial water respectively, based on the following parameters:
• desalination plant: reverse osmosis (RO) 8 MW power demand 3 production: 586 L/s or 18.5 Mm /a. • water pipeline: 24 MW power demand 115 km, 2,660 m pressure head 5 pump stations, 3 in-line (2 operating and 1 stand-by) 4 h tank reservoir each station. • potable water plant: 15 kW power demand. 3 5 L/s or 158,000 m /a located at the process plant with distribution piping to mine facilities and the SX-EW facility.
18.7.1 SEAWATER DESALINATION PLAN
The desalination plant is designed for a capacity of 51.137 m3/d. Assuming a conversion rate of 45%, the seawater flow rate required to achieve the desired output is 5.110 m3/h. An open intake solution has been adopted for seawater intake.
The intake works include construction of a buried pumping chamber and the installation of vertical, axially split-casing centrifugal pumps, variable frequency drive equipped, that will supply the flow rate required to feed the plant at the required pressure, thus optimizing energy consumption.
For pre-treatment, a double filtration system is envisaged, consisting of coarse screening by a 40 mm intake screen, and filtering by Beaudrey-type rotating filter screens with a 6 mm slot size.
An ultrafiltration system will use pressurized hollow-fiber membranes to screen out solids suspended in the seawater flow prior to entry into the desalination process. The desalination system has been designed on the basis of RO technology using
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spiral wound membranes. Membranes are made of polyvinyl-pyrollidone (PVP)/polyethersulfone (PES).
Water produced by the desalting plant will be stored in a closed tank in amounts sufficient to meet at least four hours worth of supply requirements. The tank will be made of carbon steel with a capacity of 9,000 m3. Overall power consumption of the desalination plant will be approximately 8 MW.
18.7.2 PUMPING AND CONVEYING THE WATER TO THE MINE
The impulsion system consists of a ND 700 mm steel and glass reinforced plastic (GRP) pipeline, 114.7 km long, running from the desalination plant to the Project’s processing plant. The water delivery point is at an elevation of 2,660 masl. The flow rate will be 586 L/s, carried by 5 pumping stations. The pumping system will be 3+1R (in standby mode), using horizontal multi-stage pumps. The total power demand of the pumping system will be approximately 24 MW.
18.7.3 DRINKING WATER PLANT
The Project includes a drinking water plant to be located near the Zafranal Mine, which will produce drinking water at a flow rate of 5 L/s. In order to meet the quality standards for drinking water, the design considers a plant sized to treat 18 m3/h flow of water taken directly from a desalinated water tank and pumped to the drinking water plant installed within a maritime container. The modular drinking water plant will consist of a cartridge filter (as a pre-RO safety system), an RO skid and a remineralization system. The estimated power consumption of the drinking water plant is estimated at 15 kW.
18.8 P OWER S UPPLY AND D ISTRIBUTION
Knight Piésold completed a conceptual level study update for the power delivery and site power distribution systems, based on a total connected power of 140 MW and 135 MW of peak power demand broken down as follows:
• 8 MW for the desalination plant at the coast • 24 MW for the five pump stations along the approximately 114.5 km long desalinated water pipeline from the coast to the plant site • 80 MW for the plant site, including 2 MW to supply power to the TMF • 15 MW for sulphide extraction/electrowinning at the heap leach facility • 11 MW for the mine • 2 MW for contingency.
Power delivery to site will require transmission lines to be installed linking the Project site to the local electrical supply grid. It is estimated that main line power supply will be required at the mine site six months after the start of pre-production stripping
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activities. It is estimated that construction of the main power line and substation will take a year. Power supply will be delivered from the 220 kV Socabaya Substation located in Arequipa and will require the installation of an approximately 120 km long circuit to a new 220 kV substation at the Project site from which power would be re- distributed throughout the plant site as required.
Distribution lines include:
• 115 km, 66 kV for the fresh water supply system • 45 km, 25 kV for the TMF.
Power during construction of the main project components will be provided by the various contractors via portable, diesel powered generators. The installation of the power delivery and site power distribution systems will be completed in time to support mine operations.
18.9 O THER I NFRASTRUCTURE
Examples of other infrastructure include:
• A waste rock dump will cover an area approximately 1,500 m by 1,000 m and will have a capacity of 490 Mt. • A permanent, 233-person camp will be constructed to house some on site contractor personnel and mine operator staff during the pre-mine period and as well as in the production phase of the Project. • A temporary 900-person construction camp during the peak of construction. • Laydown areas including:
construction materials warehousing permanent warehousing low grade mineral stockpile. • Fuel stations including: communication towers, server room and related IT service facilities on-site network of service roads main gate access facility buildings for administration, laboratory and emergency services. • A truck shop which will include vehicle washing and repairing facilities for all types of site vehicles. This facility will cover an area of 400 m long by 200 m wide and consist of a series of concrete pads and prepared gravel surfaces, steel structures will support overhead cranes required in the maintenance areas and will be a steel building with roof and walls. The building will be supported by concrete spread footings with concrete grade walls along its perimeters. Sumps and trenches will be constructed to collect waste water
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and waste oils in the maintenance bays. The building will consist of an exterior wash bay, repair bays, warehouse area, welding area, machine shop, emergency vehicle parking, first aid room, electrical room, compressor room and a lubricants storage room. The mine dry and a dining facility will also be located in the truck shop area.
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19.0 MARKET STUDIES AND CONTRACTS
The final products to be produced by the Project are copper concentrate and copper cathode. Copper concentrate will likely be transported to Asian smelters located in Japan, China or India, while copper cathode will likely be transported to Europe. A more precise marketing plan and terms of sale of final products will be prepared during the next phase of the Project.
Smelting, refining and transportation terms for copper concentrate as well as copper cathode were prepared by Ayrmin Consultorias y Representacions Ltda (Ayrmin) located in Santiago, Chile. A summary of their recommendations is presented in Table 19.1.
Table 19.1 Commercial Terms for Sale of Products
Parameter Unit Estimate Remarks Realization Charges Smelter Copper Deduction % 3.40 Minimum deduction 1% unit Payable Gold in Copper Concentrates % 90 - Minimum Gold Content in Copper Au g/t in Cu 1 If content below 1 g/t min Concentrates concentrate deduction 0.4 g/t, no refining charges Concentrate Selling Cost Marketing Costs – Concentrates $/t concentrates 12.00 Includes representation fee Transport – Concentrates – Truck to Port $/t (wet) concentrates 1.00 Equivalent to $15.00/t considering 150 km to port Port and Insurance Charges Concentrate $/t (wet) concentrates 20.00 - Transport Losses (Plant to Loading Port) % of wet tonnes of 0.3 - concentrates Shipping Concentrate $/t (wet) concentrates 65.00 - Smelting Charges – Concentrate (Dry) $/t concentrates 63.50 - Refining Charges – Payable Copper $/lb Cu 0.0635 - Price Participation Trigger $/lb Cu 0.00 - Refining Charges – Payable Gold $/oz Au 4.00 If content below 1 g/t no refining charges applicable Cathode Selling Costs Marketing Costs $/t cathode 5.00 - Transport Cathodes via Truck to Port $/ton per 10 km 1.00 Equivalent to $15.00/t considering 150 km to port Port and Insurance Costs $/t cathode 10.00 - Shipping Costs to Rotterdam $/t cathode 67.00 - Annual Cathode Premium (Grade A LME) $/t cathode 90.00 -
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20.0 ENVIRONMENTAL STUDIES, PERMITTING, AND SOCIAL OR COMMUNITY IMPACT
20.1 E NVIRONMENTAL S TUDIES
Environmental and social baseline studies for the Project have been conducted in order to obtain environmental certifications and permits for the exploration programs of the Project. These baseline studies covered physical, biological, and socio- economic assessments of the areas to be explored, as per the Ley del Sistema Nacional de Evaluación de Impacto Ambiental (Ley del SEIA) or National Environmental Impact Assessment System Law.
All the environmental certifications and permits for the exploration activities of the Project required the support of local inhabitants in defined areas of direct influence (ADI). This support was accomplished through informative workshops and meetings with local authorities and community representatives, guided visits to the Project area, participatory environmental monitoring programs and an ongoing formalization program of the miners working in some of the mining concessions of the Project.
AQM has an EMP up and running as a result of the potential impacts identified in the EIAsd and the conditions set forth in the drilling permits granted by the Ministry of Energy and Mines.
The drilling permit currently in place is valid until August 1, 2013 and allows for drilling from up to 226 platforms on the Zafranal Main, Victoria, Sicera Norte and Sicera Sur Zones. A modification to the EIAsd drilling permit to allow drilling additional areas on the Property is in progress at the time this report was prepared by Tetra Tech.
The formal environmental baseline studies leading to an EIA for the development and the operation of the mine itself will be carried out in parallel with the prefeasibility study for the Project.
A more detailed environmental management and closure plan for the Project will be developed in the prefeasibility study. However, a cost estimate for an EMP has been included as part of the capital and operating costs.
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20.2 E NVIRONMENTAL P ERMITTING
Aside from reconnaissance and prospecting activities, all other mining activities in Perú must comply with the Sistema de Concesión Minera or Mining Concessions System, governed by the Ley General de Minería or General Mining Law. This system establishes that mining exploration and exploitation can only be legally carried out by means of a valid mining concession title issued by the Instituto Geológico Minero Metalúrgico (INGEMMET) or the Metallurgical Mining Geological Institute. Mining concessions are granted for specific areas (established by UTM coordinates) and for either metallic or non-metallic minerals. Titleholders of mining concessions acquire ownership of all minerals extracted within the concession area, considering the type of mineral concession granted.
20.2.1 ENVIRONMENTAL IMPACT ASSESSMENTS
Estudios de Impacto Ambiental or EIAs are required prior to carry out mineral exploration and exploitation in Perú, including project components that are not mining related (electrical transmission and port installation facilities, for example). The national authority that evaluates the EIA is the Ministerio de Energía y Minas (MINEM) or Ministry of Energy and Mines through the Dirección General de Asuntos Ambientales Mineros (DGAAM) or the General Directorate for Environmental Mining Matters. In addition, the evaluation of the EIA takes into account the opinion of the Autoridad Nacional de Agua (ANA) or the National Water Authority. The preparation of the EIA must follow Articles 47, 48, 49 and 68 of the Ley del SEIA regulations. These articles explain that the content of the EIA needs to include a feasibility design for the Project, mandatory public hearings and documented community participation prior to presenting the EIA, and comply with all requirements from the MINEM. The approval process takes at least eight months.
20.2.2 OTHER PERMITS
In addition to the EIA, other permits are required throughout the stages of the life of the mine including exploration, development, production, and closure. These stages are listed below with the main permits required for each stage:
• Exploration:
mining concession surface rights surface water/groundwater use license sanitary authorization approving waste water treatment system and discharge archeological evaluation registration as a direct consumer of liquid fuels (fixed or mobile facilities)
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• Development and Construction: . mining concession . surface rights EIA (modifications during the life of the mine may apply) mine closure plan approval (modifications during the life of the mine may apply) Certificate of Non Existence of Archaeological Remains (CIRA) surface water/groundwater use license authorization for water constructions authorization for water studies sanitary authorization, approving wastewater treatment system and discharge sanitary authorization for drinking water treatment system registration as a direct consumer of liquid fuels (fixed or mobile facilities) bi-annual authorization for explosives use (global authorization) license to operate explosives shack user’s certificate for controlled chemical substances and products operation license for radioactive equipment authorization for the exploitation of construction materials – quarries • Production: . the above mentioned permits beneficiation concession authorization to initiate mining exploitation posting of financial assurance for closure • Closure:
final closure plan (due two years before final closure).
20.2.3 MINE CLOSURE PLAN
The Mine Closure Plan must be submitted within one year after EIA approval. Posting of the corresponding financial assurance for closure must also be completed before the start of production and within the first 12 business days of the following year in which the closure plan is approved, as stated by the Regulaciones para el Cierre de Minas or Mine Closure Regulations. The final closure plan shall be submitted two years prior to closure of operations.
20.2.4 ENVIRONMENTAL LEGISLATION
The main Peruvian environmental laws and regulations that apply to new mineral activities are listed below:
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• Supreme Decree No 014-92-EM – General Mining Law • Law No 28611 – General Environmental Law • Law No 28245 – National Environmental Management System Law • Supreme Decree No 008-2005-PCM – Regulations on National Environmental Management System Law • Law No 27446 – National Environmental Impact Assessment System Law • Supreme Decree No 019-2009-MINAM – Regulations on the National Environmental Impact Assessment System Law • Law No 29325 – National Environmental Enforcement System Law • Law No 28090 – Mine Closure Law • Supreme Decree No 033-2005-EM – Regulations on Mine Closure Law • Supreme Decree No 016-93-EM – Environmental Protection Regulations for Mining-Metallurgical Activities and Amendments (Supreme Decree No 059- 93-EM, Supreme Decree No 058-99-EM, Supreme Decree No 022-2002- EM, Supreme Decree No 078-2009-EM) • Supreme Decree No 020-2008-EM – Environmental Regulations for Mineral Exploration • Law No 29338 - Water Resources Law • Supreme Decree No 001-2010 – AG – Water Resources Law Regulations • Supreme Decree No 002-2008-MINAM – National Environmental Standards for Water Quality • Supreme Decree No. 010-2010-MINAM – Maximum Permissible Levels for Liquid Effluents in Mining-Metallurgical Activities • Supreme Decree No. 003-2010-MINAM – Maximum Permissible Levels for Effluents in Domestic and Municipal Waste Water Treatment Plants • Supreme Decree No. 003-2008-MINAM – Environmental Standards for Air Quality • Ministerial Resolution No, 315-96-EM – Maximum Permissible Levels of Elements and Compounds Present in Gaseous Emissions Resulting from Mining-Metallurgical Units • Supreme Decree No. 085-2003-PCM – Regulations of National Environmental Standards for Noise Quality • Law No 27314 – General Solid Waste Law • Supreme Decree 057-2004-PCM – Solid Waste Regulations • Law No 28296 National Cultural Heritage and Archaeological Investigations • Law No 26834 – Protested Natural Area Law
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• Supreme Decree Nº 038-2001-AG – Regulations on Law of Protected Natural Areas and amendments • Law No 29785 – Indigenous People Public Consultation Law • Supreme Decree No 001-2012-MC – Regulations on Indigenous People Public Consultation Law • Supreme Decree 028-2008-EM – Regulation on Citizen Participation in the Mining Subsector • Ministerial Resolution No 304-2008-MEM/DM – Provisions on the Citizen Participation Process in the Mining Sector
20.2.5 PERUVIAN ENVIRONMENTAL GUIDELINES
Peruvian mining environmental guidelines provided by the MEM covers monitoring protocols for air and water quality, water management, acid mine drainage management, evaluation of impacts resulting from mining-metallurgical activities, mine closure and others.
20.2.6 APPLICABLE INTERNATIONAL STANDARDS
A number of Canadian and international sources provide guidance to the design, operation and closure of TMFs. The design will be compliant with the Mining Association of Canada’s 2011 Guide to the Management of Tailings Facilities, and with the tailings management strategies outlined in the International Finance Corporation (IFC) and World Bank’s Environmental, Health and Safety Guidelines for Mining (IFC and World Bank 2007). The recommended tailings management strategies include:
• consideration of zero discharge tailings facilities. • independent (third-party) assessment of risks to surface water and groundwater and users of those resources downstream of the facility together with mitigation of impacts, including lining of the facility. • monitoring of the tailings facility, including the physical structure and water quality in surrounding areas. • on-land disposal in a system that can isolate material from oxidation or percolating water, such as a tailings impoundment with dams and subsequent dewatering and capping. • on-land disposal alternatives will be designed, constructed and operated according to internationally recognized geotechnical safety standards. • design specification should take into consideration the probable maximum flood event and the required freeboard to safely contain it (depending on site specific risks) over the planned life of the TMF, including the post-closure phase
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Given the TMF’s location upstream of the community of Corire, AQM may wish to consider engaging an external expert during the detailed design stage to review the TMF design, as recommended by IFC Guidance Note 5, which accompanies the IFC Performance Standard 5 for Community Health and Safety and states the following (IFC 2012):
When structural elements or components, such as dams, tailings dams or ash ponds are situated in high-risk locations, and their failure or malfunction may threaten the safety of communities, the client will engage one or more external experts with relevant and recognized experience in similar projects, separate from those responsible for the design and construction, to conduct a review as early as possible in project development and throughout the stages of project design, construction, operation and decommissioning.
20.3 E N VIRONMENTAL M ANAGEMENT P LAN
The EMP for the mine should be administrated by a manager and include a team of environmental technicians to perform all monitoring activities at the mine, including the TMF. Quarterly and annual groundwater seepage reporting is required.
20.3.1 POTENTIAL ENVIRONMENTAL IMPACTS
The Project will be subject to an environmental assessment in the future. Based on the work to prepare this conceptual level PEA, the following potential impacts have been identified regarding the Zafranal TMF:
• Risk of an unplanned breach of an embankment causing release of tailings to the environment. • Seepage of groundwater:
Tailings pore water has the potential to seep through the embankment and from the base of the TMF into the groundwater. Containment of the tailings pore water within the TMF will be necessary given the absence of low permeability materials in the foundation. However, the tailings will provide an effective barrier against seepage into the foundation as the tailings blankets the impoundment and consolidates to a less permeable state. The process plant will recycle water to the greatest extent possible, prior to discharging tailings to the TMF. The TMF will be designed as a zero-discharge facility due to the catchment size and high evaporation rate. The facility will also accommodate extreme rainfall conditions. • Dust and noise:
There are no communities or activities near the TMF that are expected to be affected by dust or noise emissions from the TMF operations. Nevertheless, standard dust and noise suppression measures will be implemented during operations to provide for a safe and healthy
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workplace. Tailings will be discharged into the TMF from off-takes in the tailings distribution pipeline along the crest of the embankments and at selected locations around the tailings basin. Discharge will be rotated around much of the TMF and managed in order to ensure efficient filling and pond development. These processes will help minimize dust generation during operations. At closure, the TMF will be capped so as to minimize the long-term impact of dusting from the TMF. • Storm containment and unplanned embankment failure:
The TMF will be designed as a zero-discharge facility under all conditions during operations, and therefore discharge of surface water to the downstream environment is not a concern. Embankment stability will be monitored using geotechnical instrumentation (piezometers, etc.), settlement monuments and traditional surveying methods. Additional safeguards that may be employed, if necessary, include buttressing and re-grading of slopes.
20.4 T AILINGS M ANAGEMENT
The TMF alternative discussed in this report is situated approximately 26 km upstream from the community of Corire, which is the nearest community to the TMF. The location of this community influenced the hazard potential classification for the design of the TMF embankments and its potential for seepage. There are no communities or activities close to the Project that are sensitive to dust and noise.
With regards to tailings water management, free water from discharged tailings and seasonal runoff will collect in a small pond in the south end of the tailings impoundment, from which the water will evaporate. No operational spillway will be required for the impoundment during operations; however, a spillway will be constructed at closure. Sufficient freeboard to contain extreme precipitation events will be maintained during mine operations. The Project is designed to recycle as much water as practical. The TMF will be a zero-discharge facility; the tailings embankments will incorporate a geomembrane liner and grout curtains to minimize seepage to the environment.
20.4.1 TAILINGS DRAINAGE
The TMF embankments will be provided with drainage measures; including internal drains which will be provided to collect internal seepage and consolidation water. An upstream toe drain will be constructed underneath the starter embankments to ensure adequate drainage. Seepage through the impoundment will be directed to this drain and be discharged into seepage collection ponds downstream of the embankments. An embankment drain will be installed during the first embankment raise to assist in the consolidation of the tailings mass. The collected seepage will be directed to the seepage collection ponds. Collected seepage will be pumped back into the impoundment while a portion of the collected seepage will evaporate from the collection pond. Seepage estimates for the tailings impoundment will be
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carried out with finite element modelling using appropriate computer programs. Water is critical for operations, and there may be a small amount of excess of water in the tailings impoundment from time to time.
20.4.2 TAILINGS MONITORING
An operation, maintenance and surveillance manual (OMS) will be developed prior to operation of the TMF. The OMS will consider the above measures and apply the guidance provided by the Mining Association of Canada’s Developing an Operation, Maintenance and Surveillance Manual for Tailings and Water Management Facilities, 2003. The environmental monitoring components will consist of the following elements:
• Monitoring of water levels and collection of seepage (if any) from the downstream groundwater monitoring wells on a quarterly basis during operations. Groundwater samples will be analyzed for general chemistry and metals (total and dissolved). Groundwater quality results will be compared against Perú’s Maximum Permissible Levels (MPLs) for Liquid Effluents in Mining-Metallurgical Activities. Sampling frequency will be adjusted depending upon the presence of groundwater seepage in the wells and the groundwater chemistry results. Exceedance of the MPLs will be flagged and reviewed in detail, and corrective actions will be undertaken. • Sampling of tailings supernatant on a quarterly basis and testing for general chemistry and metals, for comparison purposes with groundwater seepage. • Visual inspection by site environmental staff regarding unusual seepage or excessive dust emissions. • Visual inspections by TMF operators; with any unusual conditions regarding environmental concerns reported.
Monthly geotechnical inspections will be undertaken and monthly reports shall be submitted to the General Manager or Mine Manager. Reporting shall include:
• a summary of activities undertaken during the reporting period • any material deviations or non-compliances to this Plan de Administración Ambiental • planned activities during the next reporting period • any other issues of concern
Annual geotechnical inspections will be carried out by a qualified geotechnical engineer. A summary of TMF performance, including the results of geotechnical inspections and water quality results will be presented in an annual report.
A decision will be made regarding the required monitoring post-closure as the Project approaches closure.
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20.4.3 TAILINGS FACILITY CLOSURE
Free water in the tailings impoundment at closure will be allowed to evaporate. The tailings surface will be covered to promote shedding of water and to reduce infiltration at closure. The embankment drains will continue to operate for a period of time, and the water quality will be monitored. The seepage will be evaporated in specially designed evaporation ponds. A spillway will be constructed at closure to ensure long-term stability. A conceptual level closure plan for the TMF has been prepared by Knight Piésold.
Three principal objectives for the TMF closure have been considered:
• long-term physical and chemical stability of the Project area needs to be secured in order to protect community health and safety • natural recovery of disturbed areas must be optimized to a landscape and habitat consistent with adjacent land use • long-term maintenance and monitoring requirements need to be minimized.
20.5 S ULPHIDE T AILINGS M ANAGEMENT
Preliminary geochemical characterization of the Zafranal mineralized material body was carried out by pHase Geochemistry Inc. (2012), with the following conclusions:
• Sulphur content in the drill core assay database (mineralized material and waste) averaged 2.5%, which is typical of porphyry deposits. Acid rock drainage (ARD) and metal leaching (ML) are important issues for these types of deposits. • Based on ABA results, 70% of the samples are classified as potentially acid generating (PAG) with 6% classified as uncertain, 10% as low potential and 14% as non-acid generating (NAG). Low paste pH values indicated that a number of samples were already acidic. • There is no definitive correlation between rock type or mineral zone and ARD potential. • Elevated copper was reported over a wide distribution of sulphur contents, reflective of copper grades with the various mineral zones (oxide, supergene, hypogene). Bismith, selenium and occasionally cobalt, chromium, molydbdenum, silver, arsenic, cadmium and magnesium were elevated relative to crustal abundances.
Ongoing testing is underway to further define the potential for ARD and ML. On the basis of the testing completed to date, it is conservatively assumed that any seepage generated will exceed the Peruvian MPLs. Most of this seepage will be collected during operation and recycled to process, but groundwater wells will be installed downgradient of the seepage collection ponds to detect any seepage that bypasses
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the ponds. The wells will also have the capacity to pump this seepage into the seepage collection ponds.
It will be necessary to monitor the chemical composition of the tailings supernatant, seepage accumulating in the seepage collection pond and groundwater further down gradient to ensure that concentrations are maintained at or below MPLs. Exceedance of the MPLs in the groundwater monitoring wells located downstream of the seepage collection pond will trigger adaptive management measures, such as additional seepage interception/collection.
20.6 M INE C LOSURE AND R ECLAMATION
An updated mine closure plan and reclamation will be required for the Project. The mine closure plan should include information; such as:
• justification for the closure plan considering technical, environmental and legal aspects • objectives and how they will be met • photo evidence and details of the environmental situation prior to commencing closure activities • schedule of activities • the progressive reclamation of the site during the life of the operation • the design of tailings disposal areas • the reclamation and re-vegetation of the surface disturbances wherever practicable • a cost estimate of the work required to close and reclaim the mine • a plan for ongoing and post-closure monitoring and reporting at the site.
20.6.1 PROGRESSIVE RECLAMATION
Opportunities to progressively reclaim the TMF should be undertaken wherever practically possible. This will allow selected closure activities to be completed during operations, rather than at the end of the mine life. The tailings embankments will be designed for long-term stability, but may require buttressing and slope re-grading prior to closure. It may also be possible to progressively apply the closure cover (cap) over portions of the tailings beach to minimize the amount of dusting before closure.
20.6.2 FINAL CLOSURE
Long-term physical and geochemical stability, protection of the downstream environment and management of surface water are requirements for the TMF at closure. The following sections describe the measures that will be undertaken for
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final closure of the Project based on the conceptual design for tailings and water management.
GEOTECHNICAL STABILITY OF TMF EMBANKMENTS
The North and South Embankments are expected to remain stable at the constructed 2H:1V slopes. The embankments will be subject to annual geotechnical inspections throughout the life of the Project, and at closure will be assessed by a geotechnical engineer to evaluate stability and the need for any remedial measures at closure.
CLOSURE SPILLWAY
A closure spillway will be constructed to discharge flows from the TMF catchment to the environment downstream of the South Embankment. The spillway will pass the appropriate event to protect the embankments from overtopping and breaching. A preliminary location for a closure spillway has been identified and will involve drilling and blasting of an approximate 120 m long channel in bedrock. The excavated section will connect to an existing gully which will convey flows to the valley downstream of the South Embankment, seepage collection pond and groundwater monitoring wells.
A portion of the bedrock excavated to construct the spillway will be used as riprap to line the runoff channels that will be excavated across the tailings impoundment to direct flows off the TMF to the spillway. The remaining excavated bedrock will be used for armouring where required and will otherwise be disposed of nearby.
TAILINGS COVER
Runoff channels will be constructed across the impoundment surface to direct runoff from the catchment areas upstream of the TMF, across the TMF surface to the closure spillway. Three runoff channels have been envisaged to convey 45%, 45% and 10% of the TMF catchment area runoff, respectively. The channels will be constructed by excavating into the tailings and using the excavated tails to contour the impoundment to encourage runoff to the channels. The channels will then be lined with geosynthetics, sand and gravel and riprap.
The impoundment will be capped with a 300 mm thick layer of a local NAG sand and gravel (or other suitable material) cover material to prevent surface exposure, reduce infiltration, manage runoff and prevent dusting. It is expected that minor settlement will occur during the first few years of the closure period due to the consolidation of the tailings. The tailings will be tested and modelled during future studies to estimate potential seepage rates and consolidation characteristics during operations and post- closure.
It is expected that the 300 mm thick cover will wick water upward due to the large amount of evaporation at the site. This will prevent seepage from entering the
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tailings mass following closure. The tailings cover will be modelled during future studies to confirm this.
SEEPAGE CONTROL
A number of aspects of the TMF design will reduce the amount of seepage reporting downstream of the South Embankment. Tailings will be deposited to the TMF at 65% solids, the TMF will be designed as a zero-discharge facility and runoff will be ponded and allowed to evaporate during operations. The presence of a supernatant pond during operations is not expected until the very end of operations due to the high solids content of the tailings and the high evaporation rate.
The embankments will be lined with a geomembrane and the foundation and abutments will be grouted (as required), so no significant seepage downstream of the embankments is expected. During operations, any seepage that does occur will be collected in the toe and embankment drains and conveyed to the seepage collection ponds for recycling back into the impoundment. Any seepage that bypasses the collection systems will be intercepted in the groundwater monitoring wells and pumped back into the seepage collection ponds. At closure, consolidation of tailings will continue and some seepage is expected for a period of time post-closure.
The post-closure seepage is expected to reduce over time because the discharge of tailings will cease, water in the tailings pond will be consumed and the cover material will be applied to shed water and prevent infiltration. Since the mine is located in an arid zone, rainfall is minimal and occurs infrequently.
This conceptual closure plan has conservatively assumed that a small volume of seepage will occur for years after closure, as the tailings continue to consolidate. Depending upon the estimated rate/volume of seepage expected (estimated from consolidation modelling), the seepage collection pond located downstream of the South Embankment will be used as an evaporation pond, or a larger evaporation pond will be constructed to evaporate any remaining seepage. Shallow dykes will be constructed around the evaporation pond to divert runoff from entering the evaporation pond. It is possible, based on geochemical characterization conducted on the ore to date, that the seepage from the TMF may be of poor quality, exceeding the MPLs for liquid effluent. The volume of seepage, however, is expected to be minimal and will reduce to zero over time. More detailed study is required of both the volume and quality of seepage. Additional measures may be required to neutralize the poor quality seepage if the volume of seepage is estimated to be high. Any seepage into the evaporation pond, as well as downstream groundwater monitoring wells, will be monitored for the presence and quality of seepage for up to five years post closure.
REMOVAL OF ANCILLARY INFRASTRUCTURE
Ancillary infrastructure associated with the TMF includes access roads, tailings and water pipelines, pump stations and a landfill. The tailings and water pipelines will be
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drained and removed. The flow control box/pump station will be demolished and removed from site. The areas will be re-graded to restore the natural drainage. The access roads associated with the TMF will remain in place at closure to facilitate monitoring. It is assumed that a closure landfill will be located elsewhere on the mine site. This landfill will be capped and contoured to promote drainage and prevent erosion.
MONITORING
Closure and post-closure monitoring will consist of inspections for physical stability and groundwater (seepage) monitoring. The site will be subject to annual geotechnical inspections for a number of years post-closure to visually inspect the embankments and the success of closure measures to achieve long-term physical stability.
20.7 S OCIAL AND C OMMUNITY C ONSIDERATIONS
The socio-economic baseline studies will:
• identify communities that may potentially be affected by the development of the Project • identify potential positive and adverse effects of the Project on local communities • advise on further study requirements.
20.7.1 PROJECT LOCATION AND AREA OF DIRECT INFLUENCE
The Project is located in southern Perú, about 150 km by road (90 km straight-line distance) northwest of the city of Arequipa, east of the Majes River basin on the border between the District of Huancarqui in the Province of Castilla and the District of Lluta in the Province of Caylloma at an altitude ranging from 2300 to 2800 masl.
Principal access to the Property originates in Pedregal de Majes near the Pan American Highway and extends 5 km via paved road and 51 km of gravel road in direction of the town of Lluta, then 12 km westward on a gravel road to the principal exploration camp adjacent to the Main Zone, approximately 68 km from Pedregal de Majes.
Access to the Property is also possible via a gravel road from Corire to the satellite exploration camp at Ganchos and ultimately the principal exploration camp adjacent to the Main Zone, approximately 20 km from Corire.
Because of the steep topography of the Project area, its desertic climate and scarcity of water, there are no populated towns or communities close to the Project area.
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The ADI as established for the Project’s exploration stage corresponds to the districts of Huancarqui, Uraca, Aplao (in the province of Castilla) and Lluta (in the province of Caylloma) and the towns of Corire, El Pedregal, Cantas and Pitis in the department of Arequipa. The ADI includes the Project concessions that are located in the Huancarqui district and to a lesser extent in the Lluta district as well as other local geographical features listed as follows:
• Majes River basin: Aplao, Huancarqui and Uraca • Access road: Corire – Punta Colorada bridge – El Pedregal – Pitis – Cantas – Ganchos Camp – Zafranal Camp.
One major interest group includes the members of the Board of Majes Valley and its committees (specifically Aplao, Huancarqui, Uraca, Cantas, Pitis, Pedregal), as well as associations of shrimp fishermen. Another important interest group is the Virgen de la Asunta which has incorporated most of the informal miners working in some of the Project mining concessions. AQM has been working with the Virgen de la Asunta to formalize the mining activities of its members in response to the recent enforcement of legislation outlawing informal mining in Perú.
The Project’s ADI could be modified according to the final location of its onsite and offsite infrastructure.
20.7.2 CONSULTATION WITH COMMUNITIES
The implementation of an effective community engagement program is fundamental to the successful permitting of mining projects. As part of a comprehensive community engagement program, consultation should be initiated at an early stage of the Project. Consultation will include addressing concerns on the environmental and social-economic impacts of the Project activities during its different stages. AQM has initiated an early dialogue with the public and local authorities as part of stakeholder engagement process which involved handling communications and disclosing project information. Three liaison offices were set up near the Project area in order to receive feedback and provide information to the public in a proper and timely manner.
Consultation and the development of a working relationship with local communities typically involve the development of a series of agreements that lay the groundwork for ongoing discussion. These include:
• memorandums of understanding • protocol agreements • community consultation/participation agreements.
As project exploration and development proceeds, other agreements will become necessary, including:
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• socio-economic benefits agreements • environmental monitoring agreements • training agreements • accommodation and impact benefit agreements.
20.7.3 COMMUNITY RELATIONS PLAN
Since the beginning of the Project exploration activities, AQM has been able to establish and maintain an effective relationship with communities in the area of influence of the Project by keeping them informed about project activities, having them participate in environmental monitoring programs, hiring temporary local labour and supporting social development initiatives to improve educational infrastructure and programs, local infrastructure and to promote small businesses. The Project has been widely accepted by the public.
The Community Relations Plan (CRP) for the Project addresses the actions necessary to establish and maintain relations with the key authorities, institutions and local organizations to keep them informed about the Project activities, mitigate exploration impacts, and support community development programs.
The main CRP is designed to reinforce local capacity building and promote active participation of local and provincial authorities and engage leaders and representatives from Majes Valley organizations in activities aimed at improving transparency of AQM’s activities related to the Project. These actions are being carried out according to the program structure of the CRP, which integrates community development, community relations, public relations and communications. The main activities selected for the CRP supported by AQM include the following:
• consultation and disclosure • education, health and nutrition • environmental management • local employment and local economy • participatory monitoring program • infrastructure • development and capacity building of local institutions • promotion of local culture and customs.
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21.0 CAPITAL AND OPERATING COSTS
21.1 C APITAL C OST E STIMATES
21.1.1 INTRODUCTION
The initial capital cost for the Project was estimated at US$1,520 million with an expected accuracy range of ±35%. All dollar figures presented in this section are stated in US dollars, unless otherwise specified.
The estimate was developed by Tetra Tech, with input from the following consultants:
• CESEL – material take-offs for concrete, structural steel, platework, electrical and instrumentation • Knight Piésold – TMF, water management and power supply • NCL – mine development • ABEIMA – desalination plant and fresh water supply • SociedadTerral S.A. (Terral) – SX-EW plant and leach area • AQM – Owner’s costs.
The capital cost estimate consists of four main parts:
• direct costs • indirect costs • Owner’s costs • contingency.
The capital cost summary and its distribution by area is shown in Table 21.1.
Table 21.1 Capital Cost Summary
Labour Construction Cost Material Equipment Equipment Total Hours ($) ($) ($) ($) ($) Direct Costs Open Pit Mining 3,264,828 33,853,935 89,652,778 34,082,832 210,099,771 367,689,316 Process Plant 2,265,690 19,575,564 57,581,043 16,054,396 147,783,856 240,994,860 table continues…
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Labour Construction Cost Material Equipment Equipment Total Hours ($) ($) ($) ($) ($) Leach Area 182,393 1,575,876 13,239,939 3,705,163 18,578,436 37,099,415 Tailings and Water Management 25,932 224,049 49,318,009 164,128 5,402,184 55,108,370 Infrastructure 606,387 5,239,184 58,021,637 41,971,267 10,107,649 115,339,737 Fresh Water Supply 147,294 1,272,619 61,360,570 20,453,523 121,448,522 204,535,234 Power Supply and Distribution 8,978 77,566 49,675,910 60,725 441,000 50,255,201 Indirect Costs Project Indirects 41,760 4,066,099 173,894,873 - - 177,960,973 Owner's Costs Owner's Costs - - 67,777,915 - - 67,777,915 Project Contingency Contingencies - - 202,912,212 - - 202,912,212 Total 6,543,262 65,884,892 823,434,886 116,492,035 513,861,418 1,519,673,233
21.1.2 ESTIMATE BASE DATE AND VALIDITY PERIOD EXCHANGE RATE
Tetra Tech prepared this preliminary assessment estimate with a base date of Q3 2012. No escalation beyond Q3 2012 is applied to the estimate.
The estimate was prepared in Canadian dollars and converted into US dollars using a currency exchange rate of Cdn$1.00 to US$0.977.
21.1.3 ESTIMATE APPROACH
The capital cost estimate was developed based on the following:
• in-house databases; reference works were used for the equipment • preliminary material quantity estimates was provided by CESEL for civil, concrete, steel, architectural, platework and pipelines • Knight Piésold provided the material quantities for the construction of the tailings facilities • inputs for the site preparation components was provided by AQM/CESEL, with labour included in the cost • inputs for the mining components were provided by NCL • power supply and distribution costs were developed by Knight Piésold and Tetra Tech • electrics, instrumentation, piping, and heating, ventilation, and air- conditioning (HVAC) were provided by CESEL
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• estimated installation hours was based on in-house experience and cost book references • project development schedule.
All equipment and material costs are based on FCA manufacturer plant (Incoterms® 2010) and will be exclusive of spare parts, taxes, duties, freight, and packaging.
The freight costs and spares costs are covered in the indirect section of the estimate as an allowance, based on information provided by AQM.
Tetra Tech assumed the construction man-hours/work week to be 10 h/d with a 3- week-on/1-week-off rotation, with a land accessible construction camp.
Owner estimated and provided the Owner’s costs and customs and importation duties payable on goods shipped to Perú.
21.1.4 COST BASIS BY PROJECT
AREA BULK EARTHWORKS INCLUDING SITE PREPARATION, ACCESS, AND HAUL ROADS
Bulk earthwork quantities were generated from preliminary layout designs. Excavation of top soil and allowance for rock excavation was based on assumptions made at the time of the estimate preparation. Structural fill pricing was based on aggregates being produced at site utilizing a portable crushing and screening plant. The mobilization and set-up costs of the aggregate plant are included in the unit rates. The actual cost of aggregate production is included in the earthwork unit rates. Earthwork quantities do not include any allowance for bulking or compaction of materials; these allowances are included in the unit prices. Inputs for all of the above were provided by CESEL to Tetra Tech.
Tetra Tech has made the following assumptions:
• Topsoil, 300 mm average, is to be stripped and stockpiled on-site. • A total of 5% of excavated material is deemed to be unsuitable. • An average of 50% of the excavated material is deemed to be excavation in rock, of which 50% is rippable rock and the balance requires drilling and blasting. Surplus excavated material is stockpiled at a location within 5 km of the site. • Allowable ground bearing pressure is assumed to be minimum 400 kPa at the plant site location. Equipment foundations may require greater ground bearing capacity (to be confirmed by selected vendors and a geotechnical engineer in the next phase of the Project).
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• The mine site primary crusher is assumed to be located adjacent to the pit, on rock. • Rock slope is assumed to be cut at a 1:1 slope. • Allowable ground bearing pressure for structures located at the mine site is assumed to be a minimum of 600 kPa. • The borrow pit for the construction of the tailings dam was assumed to be 10 km from site. • The possible need for soil remediation, or special sub-surface measures, or the need for piling is excluded.
Access roads are based on 6 m width with chip sealor dust suppressant surfacing material or gravel.
CONCRETE
Concrete quantities were provided by CESEL. Typically all concrete is based on a 28-day compressive strength of 30 MPa. Tetra Tech used a concrete price of $641/m3, to supply and deliver to the point of placement at site. Concrete unit rates include for formwork, reinforcing steel, placement, and finishing of concrete.
STRUCTURAL STEEL
Structural steel quantities are based on quantities developed and provided by CESEL.
Craneage is included for all tonnages at a rate of $250/t.
PLATEWORK AND LINERS
Preliminary quantities for platework and metal liners for tanks, launders, pump- boxes, and chutes were estimated by CESEL.
MECHANICAL
The equipment estimate was prepared based on the project process flow diagrams and equipment list. The mechanical pricing is based on information provided by CESEL, or the in-house database or reference books.
DUST COLLECTION AND FIRE PROTECTION
Fire protection is included as a percentage of the equipment costs, based on recent project history
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PIPING AND VALVES
Piping and valve costs were estimated by CESEL.
ELECTRICAL
The power supply, distribution and sub-stations costs were provided by Knight Piésold.
The other electrical costs were provide by CESEL.
INSTRUMENTATION
Instrumentation was estimated by CESEL.
Plant control system costs are based on recent similar projects., including taxes will be supplied.
BUILDINGS
The estimate for the engineered steel framed, pre-engineered, and modular buildings is based on information provided by CESEL, or based on recent similar projects.
An in-house data base and experience with similar recent projects was used as a base for the cost estimate. The major structures and buildings are identified from general arrangement drawings.
These structures and buildings include:
• crushing building • grinding • flotation and regrind • reagents • maintenance building • truck shop and mine dry • administration • assay and metallurgical laboratory • warehouse • temporary construction and permanent operations camps • emergency response building, including medical clinic • gatehouses.
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PLANT MOBILE EQUIPMENT
The estimate for plant mobile equipment was based on information from comparable operations.
PERMANENT ACCOMMODATION AND CONSTRUCTION CAMPS
There is a construction camp included in the estimate.
TAXES AND DUTIES
The estimate was prepared with the inclusion of Peruvian taxes and duties as provided by AQM.
CONSTRUCTION INDIRECTS
Allowances have been included for construction indirects, based on a percentage of the direct costs, except for the power supply. Tailings and leaching were included as a fixed amount.
SPARES
An allowance for commissioning and capital spares was included based on the value of the process equipment.
INITIAL FILLS
An allowance for initial fills was included.
FREIGHT AND LOGISTICS
The amounts used in the estimate were provided by AQM and include taxes. The mining mobile equipment costs include freight and commissioning.
COMMISSIONING AND START-UP
An allowance for commissioning has been included in the indirect costs.
EPCM
Engineering, procurement and construction management (EPCM) was included, based on a percentage of the relevant works or as provided by the consultants
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VENDORS
An allowance for vendor representatives has been included in the Indirect costs.
OWNER’S COSTS AND PERMIT ALLOWANCES
Owner has provided a schedule of costs which are included in that section.
EXCLUSIONS
The following will not be included in the capital cost estimate:
• force majeure • schedule delays such as those caused by: major scope changes unidentified ground conditions labour disputes abnormally adverse weather conditions • receipt of information beyond the control of the EPCM contractors • cost of financing (including interests incurred during construction) • schedule acceleration costs • working capital • cost of this study • sunk costs.
ASSUMPTIONS
The following assumptions have been made in the preparation of this estimate:
• All material and installation subcontracts will be competitively tendered on an open shop, lump sum basis. • Site work is continuous and is not constrained by the owner or others. • Skilled tradespersons, supervisors, and contractors are readily available. • The geotechnical nature of the site is assumed to be sound, uniform, and able to support the intended structures and activities. Adverse or unusual geotechnical conditions requiring piles or soil densification have not been allowed for in this estimate.
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CONTINGENCY
The contingency allowance included in the estimate amounts to $202,912,212 based on an overall average 18% of direct costs.
21.1.5 OWNER’S COSTS
The Owner has provided an allowance for Owner’s costs, which is $67,777,915 plus 20% contingency as shown in Table 21.2, and included in the capital cost estimate for the Project.
The Owner’s cost estimate was developed to cover the costs associated with the pre-production operating activities as well as recruitment, training, transportation, accommodation, and catering of the workforce and contractors where appropriate. For example, concentrator and leach facility employees will be hired six months in advance of the start-up of their respective installations to ensure adequate training.
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Table 21.2 Details of Owner’s Costs
Mine Mill SX-EW G&A Costs Costs Costs Costs Construction Security Total Description ($) ($) ($) ($) ($) ($) ($) Employee Recruitment 672,300 475,606 136,077 136,469 - - 1,420,453 Employee Pre-operations Wages (Training) - 2,920,733 767,510 2,697,994 - - 6,386,237 Employee Catering Costs 1,844,195 196,433 88,487 407,507 - - 2,536,622 Employee Transportation 1,423,823 193,438 86,242 261,050 - - 1,964,553 Employee Housing Allowance 2,420,000 1,640,000 640,000 390,000 - - 5,090,000 Construction Management Costs - - - - 17,025,900 - 17,025,900 Mine Contractor Costs - 78,750 - - - - 78,750 Security - - - - - 1,142,874 1,142,874 Subtotal 6,360,318 5,504,960 1,718,316 3,893,021 17,025,900 1,142,874 35,645,389 Security Operations - - - - - 390,000 390,000 Camp Construction Costs 4,280,000 2,080,000 520,000 1,240,000 760,000 440,000 9,320,000 Project Permitting Costs - - - 2,418,500 - - 2,418,500 Owner’s Water Supply Costs - - - - 750,000 - 750,000 Owner’s Power Supply Costs - - - 1,660,100 964,500 - 2,624,600 Social Responsibility Management - - - 4,390,426 - - 4,390,426 Project Insurance Costs - - - 9,239,000 - - 9,239,000 External Consultants 1,000,000 1,000,000 500,000 500,000 - - 3,000,000 Subtotal 11,640,318 8,584,960 2,738,316 23,341,047 19,500,400 1,972,874 67,777,915 Contingency 20% 2,328,100 1,717,000 547,700 4,668,200 3,900,100 391,600 13,555,700 Total 13,968,418 10,301,960 3,286,016 28,009,247 23,400,500 2,367,474 81,333,615
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21.1.6 SUSTAINING CAPITAL
Any costs associated with work and investments that are scheduled to start after Year 1 are included in the sustaining capital costs and are not in the initial capital cost estimate. A detailed sustaining capital cost and schedule was developed for the Project and was applied to the Project’s financial model for inclusion in the Project economic evaluation.
21.1.7 ELEMENTS OF COSTS
LABOUR RATES AND PRODUCTIVITY
The labour rate was developed with consideration to the following:
• vacation and statutory holiday pay • fringe benefits and payroll burdens • overtime and shift premiums • small tools • consumables • personal protection equipment • contractor’s overhead and profit.
Based on available information, Tetra Tech has assumed that the labour source is as follows:
• 50% locally • 25% regionally • 25% out of town.
A productivity factor of 1.35 was to be applied to the labour portion of the estimate to allow for the inefficiency of long work hours, climatic conditions, and due to the 3- week-in/1-week-out rotation. This was based on in-house data supplied by contractors on previous similar projects in Perú.
21.1.8 MINING CAPITAL COST ESTIMATE
The mining capital cost estimate includes investment and operating costs associated with the construction period and considers the following items:
• pre-production costs • initial investment, replacement spare components and contingencies associated with the mine equipment fleet
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• other investment such as dispatch, training sets, etc • haul road development.
A summary of pre-production mining costs is presented in Table 21.3.
Table 21.3 Summary of Mining Costs
Costs Pre-production ($ million) Mining Operations 151.7 Mining Equipment 210.1 Other Investments 5.9 Subtotal of Direct Costs 367.7 Indirect Costs (Spares) 5.8 Contingency 38.0 Total Pre-production 411.5
PRE-PRODUCTION COST
Pre-production cost includes the mine cost associated with the removal of all the material before the startup of the processing plants. The estimated pre-production tonnage is 117 Mt and it is expected to take approximately two years to remove this material and expose sufficient mineral to initiate production. Details of the pre- production costs are shown in Table 21.4.
Table 21.4 Mining Pre-production Unit Costs
Costs Costs Pre-production (US$ million) (US$/t mined) Loading 18.6 0.16 Hauling 69.8 0.60 Drilling 9.2 0.08 Blasting (Contract Service) 17.3 0.15 Ancillary 20.9 0.18 Support 2.7 0.02 Indirects 13.2 0.11 Subtotal 151.7 1.30 Contingency (20%) 30.3 0.26 Total Unit Costs 182.0 1.56
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MINING FLEET INVESTMENT
The mining fleet investment has been developed in accordance with the mine development schedule and considered the useful life of each type of equipment. The full purchase prices of the mining equipment were obtained from estimations prepared by NCL, based on quotations received in 2011 for similar projects located in Perú and reference prices supplied by providers in 2012. The prices included freight to the site and commissioning.
A factor of 5% of the total cost of initial major equipment was added to the overall cost of equipment to fund the purchase of additional minor equipment.
The following main components are included in the initial investment: one shovel bucket, one loader bucket, two truck bodies, and five truck tires, plus 6% of total initial mining equipment first period cost for spare parts. In addition, 5% contingency was added to the total mining equipment cost to account for omissions. Table 21.5 outlines these details.
Table 21.5 Initial Mining Equipment Fleet
Costs Description Units US$ million Main Equipment Electric Drill (311 mm diameter) 3 14.2 Diesel Drill (311 mm diameter) 1 3.1 FEL (17 m3) 2 9.2 Electric Rope Shovel (56 m3) 2 53.3 Total Hauling Units (300 t) 17 101.8 Ancillary Equipment Bulldozer 1 500 hp 5 6.8 Wheeldozer 1 500 hp 3 3.3 Motorgrader 1 300 hp 3 2.7 Water Truck HD 150 hp 3 4.3 Support Equipment Support Drill Titon 1 1.5 Backhoe 1 0.2 Lube Truck 1 0.4 Support Truck 1 0.4 Mobile Crane 1 1.3 Lowboy Truck 110 t 1 4.1 Tires Handler 1 0.6 Lightning Plant 10 0.1 Cable Reeler 1 0.5 Motivator 1 1.5 Forklift 7 t 1 0.1 table continues…
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Costs Description Units US$ million Rock Breaker Hammer 1 0.2 Soil Compactor 1 0.2 Other Minor Equipment (5%) - 0.4 Subtotal Equipment Fleet 61 210.2 Extra Bucket for Electric Shovel 1 1.9 Extra Bucket for FEL 1 0.1 Tires Truck 5 0.2 Truck Extra Body 2 0.6 Components (6% Major Equipment) - 3.1 Subtotal Spares 9 5.8 Contingency (5%) - 10.8 Total Equipment Fleet and Spares - 232.7
OTHER MINING INVESTMENTS
Other mining investments were required to support the mining operation including a dispatch control system, an anti-collision system for haul trucks, a operator training system, a truck fuelling station and an explosives preparation plant. These investments are detailed in Table 21.6.
Table 21.6 Other Mining Investments
Costs Description Units US$ million Anti-collision System 1 1.6 Training System 1 1.2 Dispatch Equipment Control 1 2.2 Pit Fuelling Station 1 0.5 Explosives Plant 1 0.5 Subtotal of Other Investments 1 5.9 Contingency (15%) - 0.9 Total of Other Investments - 6.8
NCL designed the main external haul roads around the pit and platforms for the crushing station and truck maintenance shop, at an estimated cost of $34.7 million. These investments were not included in the mining capital cost however they are included under site preparation costs.
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21.1.9 PROCESS PLANT CAPITAL COST ESTIMATE
The process plant will consist of a primary stage that includes crushing and a SAG and ball mill grinding circuit, which is then followed by a flotation process to recover and upgrade copper and gold from the feed material. The capital cost of this facility is detailed in Table 21.7.
Table 21.7 Process Plant Capital Cost
Costs Description (US$ million) Crushing and Stockpile 49.2 Grinding 135.1 Flotation and Regrind 42.1 Thickening and Filtration 9.4 Reagents 2.4 Process Building 2.9 Subtotal Direct Costs 241.1 Indirect Costs 62.9 Contingency 48.2 Total Pre-production 352.2
Indirect costs include construction indirects, spares, first fills, freight, logistics and customs duties, commissioning, EPCM, and vendor assistance. A 20% contingency has been applied to total direct costs.
21.1.10 TAILINGS MANAGEMENT FACILITY CAPITAL COST ESTIMATE
Conceptual level design and cost estimates for the Project tailings and water management components are based on a total of 425 Mt of processed mineral, including 74 Mt of low grade mineral that will be processed at the end of the mine milling life.
Unit rates used in the cost estimates were based on typical recent and relevant rates for similar projects. The costs for Phase 1 construction include allowances for mobilization and demobilization at 4% and 2% of the total construction costs, respectively. Engineering costs have been included and are estimated to be 10% of the total capital costs. A 25% contingency is included in the initial and sustaining capital cost estimates. Table 21.8 details the capital cost estimate for the TMF.
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Table 21.8 Tailings Management Facility Capital Costs
Costs Description (US$ million) Mobilization and Demobilization 3.1 South Embankment Construction 29.2 Water Management System 1.3 Tailings Transport 20.7 Miscellaneous 0.1 Subtotal 54.4 Engineering (10%) 6.0 Contingency (25%) 14.0 Total Capital Cost 74.4
Knight Piésold has estimated the cost of closure for the TMF at $33.5 million and this estimate has been included in the Project´s sustaining capital.
21.1.11 LEACH FACILITY
Terral, in collaboration with Transmin, produced a conceptual study for the proposed leach facility at an estimated total capital cost of $61.85 million. Table 21.9 outlines the capital costs for the leach facility.
Table 21.9 Leach Facility Capital Cost
Costs Area (US$ million Agglomeration - Leaching 23.54 Solvent Extraction 1.73 Electrowinning 6.97 Ponds 0.76 Services 2.26 Inventories 1.84 Subtotal Direct Costs 37.10 Insurance and Freight 1.86 Commissioning and Start-up 1.11 EPCM 5.57 Overhead 5.48 Profit 5.11 Subtotal Indirect Costs 19.13 Direct Costs 37.10 Indirect Costs 19.13 Contingencies 5.62 Total Capital Cost 61.85
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The costs for the Project include the areas of leaching, solvent extraction, electrowinning, ponds, general facilities and services and initial inventory, with a ±30% margin of error. Investment amounts have been determined on the basis of prices prevailing under market conditions in Chile, which may vary from those in Perú.
CALCULATION CRITERIA
The initial investment includes costs for process equipment, assembly works, civil work, piping, electric power and instrumentation, initial inventory, detailed engineering, insurance and freight, general facilities and services, and utilities in the various area of the proposed leach facility.
EQUIPMENT
The equipment considered in the project has been estimated on the basis of the quotes complied in Terral’s database for equipment featuring similar characteristics, which were then scaled using factor-based estimation methodologies, such as William’s and Lang’s methods.
ERECTION WORK
Erection work for the various processes was estimated at 30% of the total value of each item. This factor considered the quality of human resources and their average performance levels from Terral’s experience for this type of the Project.
CIVIL WORKS
For this item, volumes were estimated on the basis of average earthmoving data for projects of this type. The same applied to structures and concrete.
The unit prices set forth reflect the values available in Terral’s databases. Additionally, suppliers were asked for quotes regarding for pools liners and the tank house.
PIPING AND PONDS
The volumes considered were for piping in the various areas for this size of the Project, and the unit prices were those available on Terral’s database. This approach was taken with regard to HDPE and carbon steel pipes and fittings in general.
The volumes determined by the Project and the base prices of domestic suppliers were used to size and price the ponds. In parallel, factorial methods were used to confirm the relevant cost estimates.
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INSTRUMENTATION AND ELECTRICAL POWER
A medium level of instrumentation was selected, which was estimated at a conservative 12% of the Project’s total direct costs. In addition, for the purposes of estimating the instrumentation equipment, Terral relied on past experience to assigned appropriate factors. These varied between 5 and 40% of the equipment costs for each individual process stage.
With regard to electric power consumption and cost estimation, experience with similar plants enabled the identification of all wiring, switchboards, panels and electrical rooms that should be required for the Project.
INITIAL INVENTORY
Initial inventories were based on the mass balance, the preliminary dimensioning of the equipment, and process needs. Terral assumed a minimum three months stock, and used the following factors for calculation purposes:
For extractant, stock is a function of equipment (mixer-decanters, organic pond and 8% for lines) and losses due to effluent generation.
For solvent, stock is a function of equipment (mixer-decanters, organic pond and 8% for lines) and evaporation
For guar and cobalt sulphate, stock is a function of output.
INDIRECT COSTS
To calculate investment associated to indirect costs, we applied a percentage to direct costs on the basis of Chilean standards, which are:
• insurance and freight: 5% of direct costs • commissioning and start-up: 3% of direct costs • EPCM: 15% of direct costs • general facilities and services: 12% of direct costs • utilities: 10% of direct costs, insurance and freight, commissioning and start- up, EPCM, and general facilities and services.
CONTINGENCIES AND INCIDENTALS
The total investment cost includes as “contingencies and incidentals” the equivalent of 10% of investment associated to direct and indirect costs.
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21.1.12 POWER SUPPLY
Knight Piésold was retained by AQM to complete a conceptual level study update for the power delivery and site power distribution systems, based on a total connected power of 140 MW with a peak load requirements of 135 MW. As shown in Table 21.10, the estimated capital cost of the power supply and distribution is $59.2 million.
Table 21.10 Power Supply and Distribution Capital Cost
Costs Description (US$ million) Mine Substation 5.12 220 kV Transmission Line 28.33 25 kV Distribution Lines 2.96 Diesel Plant and Pump Station 13.10 Subtotal of Direct Costs 49.51 Indirect Costs 4.95 Contingencies 4.74 Total Capital Cost 59.20
The preferred option was a connection to the Socabaya 220 kV substation, near the city of Arequipa and approximately 120 km from the Project site. Knight Piésold developed a conceptual level study cost estimate for the power delivery and site power distribution systems. Power during construction of the main project components will be provided by the various contractors via portable, diesel-powered generators.
21.1.13 WATER SUPPLY
ABEIMA completed a preliminary conceptual design and an indicative cost estimate for the Project’s water supply system that will provide seawater intake and treatment by a desalination plant, storage of product water, brine discharge into the sea, pumping 586 L/s of treated water up to the mine, and water treatment plant for drinking purposes. Table 21.11 outlines the total capital cost of the Project water supply.
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Table 21.11 Water Supply Capital Costs
Costs Description (US$ million) Desalination Plant 57.5 Pipeline and Pumping 143.6 Potable Water 0.8 Distribution 2.5 Subtotal 204.4 Indirects 42.9 Contingencies 27.5 Total Capital Cost 274.8
21.2 O PERATING C OST E STIMATES
21.2.1 SUMMARY
The total operating cost for the Project is estimated to be US$8.29/t of plant feed material processed including mining, processing, leaching, G&A, and plant services. The unit costs, as summarized in Table 21.12, are based on an annual treatment rate of 29,200,000 t/a, or a production rate of 80,000 t/d and 365 days of operation. All dollar figures presented in this section are stated in US dollars, unless otherwise specified.
Table 21.12 Operating Cost Summary
Unit Cost Area ($/t processed) Mining 2.56 Milling 4.23 Tailings and Water Management 0.07 G&A 0.49 Leaching 0.94 Total Operating Cost 8.29
The total operating unit cost for the Project per pound of copper produced is estimated to be US$1.26, as shown in Table 21.13.
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Table 21.13 Mine Unit Operating Cost Summary
Unit Cost Cost Item Unit ($) Mining $/t mined 1.40 $/t milled 3.08 Milling $/t milled 5.10 Tailings Management $/t milled 0.08 G&A $/t milled 0.60 Subtotal Flotation $/t milled 8.86 Subtotal Leach $/t leached 5.54 Total Flotation and Leach $/t processed 8.29 Subtotal Flotation $/lb Cu Produced 1.21 Subtotal Leach $/lb Cu Produced 1.82 Total Flotation and Leach $/lb Cu Produced 1.26
21.2.2 MINING OPERATING COST ESTIMATE
The operating cost estimate is based on price information for consumables, maintenance and labour costs gathered by NCL from other similar projects in the area.
RELEVANT CONSUMABLE PRICES
Table 21.14 shows consumable pricing used in the calculation of equipment operating costs.
Table 21.14 Relevant Consumables Prices
Description Unit Price Fuel $/L 0.994 Power $/kWh 0.070 Lube $/L 1.725 Grease $/kg 3.93 ANFO $/kg 0.5744 Tires FEL $/ea 61,785 Truck $/ea 35189 Wheeldozer $/ea 10,291 Grader $/ea 59,83 Water Truck $/ea 14,630 Drill Steel Bit $/ea 3,500 Main Rod $/ea 6,400 table continues…
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Description Unit Price Adapter $/ea 1,200 Bit – Rod Adapter $/ea 2,800 Guide $/ea 2,000 Starter Road $/ea 6,100 Shock Absorber $/ea 22,000 Note: ANFO = ammonium nitrate/fuel oil
MAINTENANCE
A full maintenance and repair contract (MARC) was considered for the operating cost estimates. Costs were estimated from recent quotations for similar projects.
LABOUR
An absenteeism factor of 12% was used in the calculation of the workforce required in the mine. This factor yielded a total of 4.48 operators per effective equipment (discounting mechanical availability and utilization), which is applied to the truck, support and ancillary fleets. For drill rigs and loading equipment the 4.48 operators factor is applied to the total fleet.
The mine indirect manpower includes personnel allocated to mine management, mine operations, maintenance services, technical services and mine geology.
The cost of personnel for direct maintenance was considered as it is part of the MARC cost option.
Labour costs were calculated using the yearly cost per labour category equal to an average of salaries from two comparatively similar Peruvian mining operations. Workers profit sharing was not incorporated into the wage structure but was included in the economic evaluation of the Project.
GENERAL AND ADMINISTRATIVE EXPENSES
Indirect expenses related to the mine site include:
• office supplies • safety equipment • rental of light vehicles • diesel for light vehicles • software • training • consultants
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• infill drilling • tools • mineralized material control • other minor expenses.
Mine drainage has not been included in the indirect figures, due to the dry weather in the mine area. However, this exclusion must be analyzed in more detail in future evaluations as some sporadic rains occur in the region which could require channels to divert rain water.
BLASTING SERVICE
The mine will contract out this service including the supply of mix trucks and 18 trained personnel who will carry out the delivery of the explosive mix to the drillhole and the blasting operation. The fixed annual cost of this service is estimated from a reference quote at $1,275,000 not including consumables. The minimal quantities of water detected in the exploration diamond drillholes supports the conclusion that blasting will be performed using 100% ANFO. An allowance for the explosive plant infrastructure has been added to the capital cost estimate.
MINING OPERATING COST SUMMARY
The operating cost of the total mine movement was calculated based on all information described in this section, including the cost of plant feed re-handle. Table 21.15 summarizes the mining operating costs per activity for the production period and displays those same costs as unit costs in US$/t mined.
Table 21.15 Mining Operating Cost Summary
Production US$ million US$/t mined Loading 187.8 0.20 Hauling 636.3 0.68 Drilling 85.8 0.09 Blasting (Contract) 131.6 0.14 Ancillary 155.3 0.17 Support 18.0 0.02 Indirects 79.3 0.08 Subtotal 1,294.1 1.38 Contingency 18.6 0.02 Total Unit Costs 1,312.7 1.40
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21.2.3 PROCESS PLANT OPERATING COST ESTIMATE
The process operating costs for the Zafranal concentrator includes crushing, grinding, flotation, concentrate thickening and filtration to produce a saleable copper concentrate.
OPERATING COST SUMMARY
Table 21.16 gives the overall estimated cost summary for the processing facility and the G&A costs, and is based on 80,000 t/d with a mill availability of 92% and 365 operating days per year.
The operating costs have been determined using the operating plant complement required to run and maintain the plant facilities.
The manpower complement has been estimated on the basis of having four crews under equal rotation time and two 12-hour shifts per day.
The annual operating costs for the process facilities and tailings handling sections is estimated to be $148.9 million, or $5.10/t of plant feed material treated at the processing rate of 80,000 t/d.
Table 21.16 Concentrator Annual Operating Cost Summary
Cost Manpower ($) $/t milled Labour Costs Operations Staff 4 341,785 0.012 Operations Labour 76 2,220,466 0.076 Maintenance Labour 67 2,570,636 0.088 Metallurgical Laboratory 17 698,614 0.024 Subtotal – Labour Costs 164 5,831,501 0.200 Power Costs Crushing - 2,398,299 0.082 Grinding - 28,542,055 0.977 Flotation & Regrind - 4,348,176 0.149 Air and Water Supply - 1,582,061 0.054 Tailings - 160,589 0.005 Miscellaneous - 305,638 0.010 Auxilary Plant Utilities - 1,164,232 0.040 Subtotal – Power Costs - 38,501,050 1.319 Supplies and Contractors Plant Supplies - 41,129,024 1.409 Maintenance Supplies - 14,000,000 0.479 Reagents - 28,155,017 0.964 table continues…
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Cost Manpower ($) $/t milled Desalinated and Potable Water - 18,900,000 0.647 Miscellaneous - 1,471,000 0.050 Contract Services – Assays 18 957,828 0.033 Subtotal – Supplies and Contractors Costs 18 104,612,869 3.583 Total Operating Costs 182 148,945,420 5.101
The annual operating cost estimate includes the following:
• the staffing and maintenance manpower base salaries, including a burden of 71%, is based on information as supplied by Transmin • power consumption is based on the estimated power usage with the cost of power provided by Transmin • reagent consumption values, and the cost of reagents has been based on prices as supplied by Transmin, and from suppliers • estimated maintenance costs • the staffing and manpower complement based on information originally provided by Transmin and revised by Tetra Tech.
21.2.4 TAILINGS MANAGEMENT FACILITY OPERATING COST ESTIMATE
The tailings delivery pipeline is approximately 9 km long and will be 26” Schedule 40 steel pipe for the 80,000 t/d case. The emergency discharge pipeline is about 400 m long and is selected as 30” DR9 HDPE pipe for the 80,000 t/d case. This pipeline will operate by gravity. The cost of operating the tailings management system has been detailed in Table 21.17 along with the schedule for sustaining capital expenditures.
Table 21.17 Tailings Management Facility Sustaining and Operating Costs
Total Description (US$ million) Mobilization and Demobilization 5.47 Embankment Construction 90.39 Water Management System 0.43 Miscellaneous 0.30 Subtotal Sustaining Capital 96.58 Engineering (10%) 9.66 Contingency (25%) 24.15 Total Sustaining Capital 130.39
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Cost Cost Operating Cost (US$ million) US$/t milled Labour 16.80 - Water Management 3.99 - Tailings Transport 11.21 - Miscellaneous 1.44 - Total Operating 33.44 0.08
21.2.5 GENERAL AND ADMINISTRATIVE COSTS
Table 21.18 outlines the G&A cost estimate.
Table 21.18 G&A Cost Estimate
Arequipa Office Zafranal Site Total $/t G&A Cost Estimate ($) ($) ($) milled Office Expenses (all locations) 237,500 12,500 250,000 0.009 Office Supplies (all locations) 118,750 6,250 125,000 0.004 Professional Associations 62,500 - 62,500 0.002 Consultants 415,625 21,875 437,500 0.015 Insurance 5,021,250 128,750 5,150,000 0.176 Legal Services 300,000 - 300,000 0.010 Regulatory Compliance/Audit 200,000 - 200,000 0.007 Travel and Expenses 285,000 15,000 300,000 0.010 Communications: Tel/ Fax/ Internet 475,000 25,000 500,000 0.017 Computer and IT Services and Supplies 380,000 20,000 400,000 0.014 Community Public Relations and Donations 1,425,000 75,000 1,500,000 0.051 Recruitment 190,000 10,000 200,000 0.007 Training 170,000 30,000 200,000 0.007 Safety and Training Supplies 70,000 30,000 100,000 0.003 Safety Incentives 70,000 30,000 100,000 0.003 Medical Service/First Aid 120,000 80,000 200,000 0.007 Security Supplies 60,000 90,000 150,000 0.005 Environmental Management - 500,000 500,000 0.017 Janitorial 106,875 5,625 112,500 0.004 Power 285,000 15,000 300,000 0.010 Security 80,000 320,000 400,000 0.014 Wages and Salaries 2,060,000 515,000 2,575,000 0.088 Crew Transportation 200,000 800,000 1,000,000 0.034 Transportation of Supplies 450,000 300,000 750,000 0.026 Road Maintenance - 467,881 467,881 0.016 Camp Cost - 1,000,000 1,000,000 0.034 Miscellaneous 178,125 9,375 187,500 0.006 Annual G&A Expenses 12,960,625 4,507,256 17,467,881 0.600
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General Discussion
• The annual maintenance supplies requirement for the plant has been estimated to be $14.0 million, or $0.48/t of material treated. • The total cost of operating supplies for the main plant has been determined to be $89.7 million, or $3.070/t of material processed.
Assumptions Used in Power and Supplies Requirements
• The power value records the anticipated annual operating power usage and not the installed power for the equipment. The unit electrical cost of $0.070/kWh was provided by AQM. • The costs of the plant operating supplies costs include the reagent costs for the consumption value as given in the process design criteria. The grinding media costs are based on the calculated consumption values, and vendor estimates. • The maintenance supplies costs are estimated values and are reflected as an allowance as based on a percentage of the capital cost estimate. First Fills costs are estimates only based on the yearly consumption values used in the operating cost estimate. The reagent and steel allotment is an allowance for a two month supply to cater for possible delivery interruptions due to the location of the mill site.
OUTSOURCED CONTRACTS
Table 21.19 summarizes the costs for the analytical analysis requirements for the processing facility. This portion of the Project has been outsourced to a commercial laboratory. The overall costs include the manpower, equipment cost and reporting system. The manpower estimate is based on the analytical staff working 12-hour shifts.
Table 21.19 Assay Laboratory Costs
Total Cost Unit Cost Area Estimate Basis ($/a) ($/t treated) Assay Laboratory 18 Technical Personnel and Equipment 957,828 0.033 Total - 957,828 0.033
The contract estimate does not include transportation to site, food or lodging for the personnel as these costs are covered in the G&A cost summary.
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21.2.6 LEACH FACILITY OPERATING COST ESTIMATE
The leach facility operating cost estimate considered the consumption supplies resulting from the material balance; power consumption requirements of the installations, calculated on the basis of the installed capacities of the plant’s process equipment, and the associated dedicated manpower costs. Terral also considered the additional cost to transport the mineral to the leach pads as opposed to the mine waste dump, and third-party acid supply. Power outputs of each piece of equipment were used to calculation electrical consumption.
Under these conditions, the operating cost of the Project leach facility is $1.48/lb copper, equivalent to $5.58/t of plant feed.
SUPPLY COSTS
Table 21.20 details the basic supply costs that have been used for the calculation of operating costs.
Table 21.20 Leach Facility Supply Costs
Description Unit Value Sulphuric acid $/t 80.00 Fuel $/m3 1,211.04 Materials movement during work $/t 0.20 Electric power $/kWh 0.07 Extractant $/kg 15.00 Solvent $/m3 1,200.00 Anode $/unit 570.00 Permanent cathode $/unit 510.00 Colbat sulphate $/kg 12.00 Guar $/kg 5.00 Freight from plant to port $/total copper/km 0.10 Distance from plant to port km 150.00 Replacement water $/m3 1.11 Cost to the mine of plant feed in leach pad $/t 0.27
LABOUR
The leach facility labour costs include direct manpower, outsourced support services and internal administrative support.
Direct Manpower
A total staff of 64 is considered, distributed among administrative and operational shifts. A plant superintendent was also considered in the organization.
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Outsourcing
The services supporting the plant’s operation are outsourced or covered AQM personnel in other areas. The outsourced services considered were:
• accounting/payment of wages • plant protection • plant support for leach pad handling and cathode harvesting • personnel transport.
A total of 24 on-site contractor personnel were considered for the leach facility.
Table 21.21 summarizes the total unit operating costs for the leach facility.
Table 21.21 Leach Facility Operating Cost – 10,000 t/a Cathode
Daily Cost Cost $/t Cost Item Unit Consumption ($/d) ($/lb Cu) mineral Plant General Costs Plant general industrial water cost m3/d 37.56 22.30 0.0003 0.00 Plant general area power consumption kWh/d 269.99 18.90 0.0003 0.00 Contingencies estimate per area $/d - 2.06 0.0000 0.00 Work camp, administration and sales lb Cu/d 64,977.68 138.78 0.0021 0.01 Supervision/manpower cost $/mo 14,495,916 956.16 0.0147 0.06 Area outsourcing cost $/mo 13,950,000 920.15 0.0142 0.06 Total Plant General Costs - - 2,058.35 0.0317 0.13 Mine Cost Cost of MINERAL in leach pads t/d 16,374.27 4,486.55 0.0690 0.274* Total Mineral Cost - - 4,486.55 0.0690 0.27 Leaching Costs Cost of transport from agglomeration to leach pads t*km/d 0.00 0.00 0.0000 0.00 Cost of loading plant feed in leach pads t/d 16,374.27 4,912.28 0.0756 0.00 Power cost: agglomeration kWh/d 0.00 0.00 0.0000 0.00 Power cost: leaching kWh/d 46,709.39 3,269.66 0.0503 0.20 Curing water cost m3/d 0.00 0.00 0.0000 0.00 Leaching water replacement cost m3/d 1,996.68 1,185.33 0.0182 0.07 Cost of sulphuric acid onsumed t acid/d 391.69 31,335.56 0.4823 1.91 Contingencies estimate per area $/d - 2,035.14 0.0313 0.12 Work camp, administration and sales lb Cu/d 64,977.68 293.60 0.0045 0.02 Maintenance costs $/a 177,951.34 494.31 0.0076 0.03 Supervision/manpower cost $/mo 30,666.66 2,022.80 0.0311 0.12 Area outsourcing cost $/mo 3,060 201.84 0.0031 0.01 Site preparation: Year 2 onwards $/a 2,783 7,729.56 0.1190 0.47 Coatings and drains; Year 2 onwards $/a 5,860.39 16,278.85 0.2505 0.99 table continues…
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Daily Cost Cost $/t Cost Item Unit Consumption ($/d) ($/lb Cu) mineral Irrigation materials: adding/replacement $/a 910.36 2,528.78 0.0389 0.15 Total Cost of Placing Copper in Solution - - 72,287.71 1.1125 4.11 Cost of Extraction by Solvents SX Reagents: Extractant kg/d 35.98 539.77 0.0083 0.03 SX Reagents: Solvent m3/d 0.65 780.39 0.0120 0.05 Power cost in SX kWh/d 2,628.34 183.98 0.0028 0.01 Washing water cost in plant m3/d 37.56 22.30 0.0003 0.00 Contingencies estimate per area $/d - 76.32 0.0012 0.00 Work camp, administration and sales lb Cu/d 64,977.68 213.33 0.0033 0.01 Maintenance costs $/a 41,394.35 114.98 0.0018 0.01 Supervision/manpower cost $/mo 22,282,487 1,469.77 0.0226 0.09 Total Cost of Placing Copper in Electrolyte - - 3,324.17 0.0523 0.21 Electrowinning Cost Power cost in electrowinning kWh/d 62,341.48 4,456.11 0.0686 0.27 Electrolyte acid replacement t acid/d 9.26 740.80 0.0114 0.05 Treated water replacement m3/d 65.13 0.62 0.0000 0.00 Guar consumption kg/d 7.37 36.84 0.0006 0.00 Cobalt consumption kg/d 18.25 219.05 0.0034 0.01 Fuel consumption t/d 3.88 4,723.08 0.0727 0.29 Contingencies estimate per area $/d - 508.83 0.0078 0.03 Work camp, administration and sales lb Cu/d 64,977.68 254.34 0.0039 0.02 Maintenance costs $/a 229,146.88 636.52 0.0098 0.04 Supervision/manpower cost $/mo 26,566,481 1,752.35 0.0270 0.11 Cathode consumption units/d 0.00 29.47 0.0005 0.00 Anode consumption units/d 1.19 679.25 0.0105 0.04 Total Cost of Cathodes in Plant - - 14,037.26 0.2160 0.86 Total Unit Cost – Cathode Production - - 96,194.04 1.4816 5.58 Note: *Additional transport cost starts in Year 2.
21.2.7 WATER SUPPLY OPERATING COST ESTIMATE
The annual cost of operating the the water supply system is detailed in Table 21.22.
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Table 21.22 Water Supply Operating Cost Estimate
Rate Water Description Unit L/s Output Pumping Potable Total Production m3 586 18,480,096 18,480,096 150,500 18,480,096 Electrical Consumption kWh - 63,690,101 185,409,962 125,216 249,225,279 Ratio kWh/m3 - 3.45 10.03 0.83 13.49 Operating Cost Without Power $ $ 2,883,050 - 132,185 3,015,235 Electrical Consumption $ $ 4,458,307 12,978,697 8,765 17,445,770 Total Operating Cost Estimate $ $ 7,341,357 12,978,697 140,950 20,461,005 Unit Cost $/m3 $ 0.40 0.70 0.94 1.11 Concentrator Consumption (0.58 m3/t) L/s 542 - - - 18,939,910 Leach Consumption (0.133 m3/t) L/s 31 - - - 1,069,778 Dust Suppression (allowance) L/s 13 - - - 451,316 Total L/s 586 - - - 20,461,005
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22.0 ECONOMIC ANALYSIS
22.1 I NTRODUCTION
The PEA is preliminary in nature and includes Inferred Mineral Resources that are considered too speculative geologically to have the economic considerations applied to them that would enable them to be categorized as Mineral Reserves. There is no certainty that the financial results in the PEA will be realized.
Tetra Tech prepared an economic evaluation of the Project based on a pre-tax financial model. All costs in this section are expressed in US dollars.
The pre-tax financial results are:
• 26.7% IRR • 1.9-year payback from the start of mill operations on the $1,520 million of initial capital • $1,332 million NPV at an 8% discount rate.
AQM commissioned Ernst & Young, Perú to prepare a tax model for the a post-tax economic evaluation of the Project with the inclusion of applicable taxes and the Peruvian government’s mining royalty.
The following post-tax financial parameters were calculated:
• 17.4% IRR • 2.6-year payback from the start of mill operations on $1,520 million of initial capital • $588 million NPV at an 8% discount rate.
The base case metal prices used for this study are as follows:
• copper – $3.00/lb • gold – $1,274/oz.
Each of the metal prices used is calculated as the mean of the long-term (2018 to 2022 average) forecasts from 24 international financial institutions. Institutions include Bank of America, Australia and New Zealand Banking Group Limited (ANZ), Citigroup Inc., Credit Suisse Group AG, Morgan Stanley, Royal Bank of Canada (RBC), UBS, Barclays Bank PLC, and Deutsche Bank AG.
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Analyses were conducted to assess the sensitivity of the pre-tax Project merit measures (NPV, IRR and payback periods) to the main inputs.
22.2 P RE- TAX M ODEL
22.2.1 MINE/METAL PRODUCTION IN FINANCIAL MODEL
The life-of-project average material tonnages, grades and metal production are shown in Table 22.1 and Figure 22.1.
Table 22.1 Metal Production from the Zafranal Project
Description Value Mine Life (Years) 151 Material Milled/Leached Total Tonnes to Mill (‘000) 425,310 Average Annual Tonnes to Mill (‘000) 28,354 Total Tonnes to Leaching (‘000) 87,256 Average Annual Tonnes to Leaching (‘000) 6,712 Average Grade Copper (%) – Mill 0.378 Gold (g/t) – Mill 0.071 Copper (%) – Leaching 0.230 Gold (g/t) – Leaching 0.085 Total Production Copper (‘000 lb) – Mill 3,105,452 Gold (‘000 oz) – Mill 479 Copper (‘000 lb) – Leaching 265,863 Average Annual Production Copper (‘000 lb) – Mill 207,030 Gold (‘000 oz) – Mill 32 Copper (‘000 lb) – Leaching 20,451 Note: 1Leaching is active for 13 years.
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Figure 22.1 Total Metal Production
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22.2.2 BASIS OF FINANCIAL EVALUATIONS
The production schedule has been incorporated into the 100% equity pre-tax financial model to develop annual recovered metal production from the relationships of tonnage processed, head grades, and recoveries.
Metal payable values were calculated based on base case metal prices. Net invoice value was calculated each year by subtracting the applicable refining and smelting charges from the payable metal value. At-mine revenues are then estimated by subtracting transportation and insurance costs. Unit operating costs for mining, processing, power, fuel, and G&A were applied to annual mined/milled/leached tonnages to determine the overall operating cost which was deducted from the revenues to derive annual operating cash flow.
Initial capital costs as well as working capital have been incorporated on a year-by- year basis over the mine life. Salvage value and mine reclamation costs are applied to the capital expenditure in the last production year. Capital expenditures are then deducted from the operating cash flow to determine the net cash flow before taxes and mining royalty.
Initial capital expenditures include costs accumulated prior to first production of concentrate; sustaining capital includes expenditures for mining and processing additions, replacement of equipment, and tailings embankment construction.
Based on the mining schedule, first mill production will occur approximately three years following project approval.
Working capital is assumed to be three months of the annual operating cost and fluctuates from year to year based on the annual cost. The working capital is recovered at the end of the mine life.
The salvage value is assumed to be equal to the reclamation cost and both will occur at the end of mine life.
The undiscounted annual net cash flow (NCF) and cumulative net cash flow (CNCF) are illustrated in Figure 22.2.
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Figure 22.2 Pre-tax Undiscounted Annual and Cumulative Net Cash Flows
1,000 4,000
800 3,000 600 2,000 400 1,000 200
0 0 (3) (2) (1) 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 (200) (1,000) (400) Production year
Annual Net Cash Flow (US$M) (US$M) Flow Cash Net Annual (2,000)
(600) (US$M) Flow Cash Net Cumulative (3,000) (800)
(1,000) (4,000)
NCF CNCF
22.3 S UMMARY OF F INANCIAL R ESULTS
Tetra Tech evaluated the base case using long-term forecast copper and gold prices of $3.00/lb and $1,274/oz, respectively. The pre-tax financial model was established on a 100% equity basis, excluding debt financing, and loan interest charges. The financial results for the base case and for alternative cases are presented in Table 22.2.
Table 22.2 Summary of Pre-tax Financial Results
Alternate Case Base Description Case 1 2 Copper Price ($/lb) 3.00 2.70 3.30 Gold Price ($/oz) 1,274.00 1,274.00 1,274.00 Recovered Metal Value ($ million) 10,724 9,713 11,735 Off-site Costs & Deductions ($ million) 1,557 1,521 1,593 Operating Costs ($ million) 4,242 4,242 4,242 Operating Cash Flow ($ million) 4,936 3,961 5,912 Initial Capital Expenditure ($ million) 1,520 1,520 1,520 Total Capital Expenditure ($ million) 1,745 1,745 1,745 Net Cash Flow ($ million) 3,192 2,216 4,167 Discounted Cash Flow NPV ($ million) at 5% 1,852 1,206 2,498 Discounted Cash Flow NPV ($ million) at 8% 1,332 814 1,849 Discounted Cash Flow NPV ($ million) at 10% 1,062 611 1,512 Payback (years) 1.9 2.3 1.6 IRR (%) 26.7 20.4 32.5
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22.4 S ENSITIVITY A NALYSIS
Sensitivity of the project’s pre-tax NPV, IRR to the Project key variables was investigated. Using the base case as a reference, each of the key variables was changed between ±30% at 10% intervals while holding the other variables constant. The following are the key variables investigated:
• copper price • gold • capital costs • operating costs.
As shown in Figure 22.3, the Project NPV, calculated at an 8% discount, is most sensitive to the copper price and, in decreasing order, operating costs, capital costs, and gold price.
Figure 22.3 Pre-tax NPV Sensitivity Analysis
3,500 3,000 2,500 2,000 Copper price 1,500 Gold price Capital costs 1,000 Operating costs 500 0 NPV@8% Discount Rate (US$M) Rate Discount NPV@8% -30% -20% -10% 0% 10% 20% 30% -500 % Change from Base Case
As shown in Figure 22.4, the Project IRR is most sensitive to the copper price followed by the capital costs, operating costs and gold price.
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Figure 22.4 Pre-tax IRR Sensitivity Analysis
45% 40% 35% 30% Copper price 25% Gold price 20% Capital costs 15% Operating costs 10%
Internal Rate Rate of Internal Return (%) 5% 0% -30% -20% -10% 0% 10% 20% 30% % Change from Base Case
As shown in Figure 22.5, the payback period is most sensitive to the copper price and less sensitive to the rest of parameters.
Figure 22.5 Pre-tax Payback Sensitivity Analysis
6.0
5.0
4.0 Copper price 3.0 Gold price Capital costs 2.0 Operating costs
Payback Period (years) Period Payback 1.0
0.0 -30% -20% -10% 0% 10% 20% 30% % Change from Base Case
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22.5 P OST- TAX F INANCIAL A NALYSIS
AQM commissioned Ernst & Young, Perú to prepare a tax model for the a post-tax economic evaluation of the Project with the inclusion of applicable taxes and the Peruvian government’s mining royalty.
BACKGROUND
The Project is owned by CMZ, a Peruvian company formed by AQM Perú and Teck Resources, each with a 50% shareholding.
PERUVIAN TAX REGIME
CMZ is taxed with a 30% income tax rate determined over the net income which is the result of deducing from the gross income all the necessary expenses to produce taxable income or to maintain the source.
Every month, the taxpayer has to make advanced payments of the income tax of the corresponding fiscal year, applying for such effect a coefficient over the income obtained during the month. Those advanced payments are deemed credits that can be used against the payment of the income tax that will be determined at the end of the fiscal year (approximately March of the next fiscal year).
The Temporal Net Assets Tax (ITAN) is a Peruvian tax imposed on companies, agencies, branches and permanent establishments of foreign entities, over the value of its net assets. The tax base equals the value of the net assets of the taxpayer as of December 31 of the preceding year that exceeds S/.1 million (approximately US$390,000). From January 1, 2009, the tax rate is 0.4%, which is applied over the value of the net assets that exceeds S/.1 million.
The ITAN´s payments may be used to offset the advance payments required under the general Income Tax regime or may be claimed as a credit against the income tax payable for the taxable year. A refund may be requested for any balance of tax payment that is not used in the current year. It is important to consider that the ITAN is payable at the beginning of the year following the first year of productive activities.
Depreciation rates for buildings and construction are at 5%, and mining machinery or equipment is up to 20%. With the exception of buildings and construction, tax depreciation must be the same as financial reporting depreciation.
It is expected that CMZ will enter into a tax stabilization agreement with the Peruvian government that will entitle CMZ guaranteed stability concerning the tax regime, currency exchange regime, availability of foreign currency and entitlement to a depreciation rate of 20% of movable assets and 5% for buildings and construction. Notice that if the Tax Stabilization Agreement is signed by virtue of the General Mining Law, the corporate tax rate applicable would rise to 32%.
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In October 2011, a special mining tax was imposed, which is applied on operating profit derived from sales of metallic mineral resources (regardless of whether the mineral producer owns or leases the mining concession) based on a sliding scale with progressive marginal rates ranging from 2 to 8.4%. The tax liability arises and becomes payable on a quarterly basis.
Recently modified mining royalties will also be applied to AQM’s operating income. The royalty is payable on a quarterly basis with marginal rates ranging from 1 to 12%. This new system has been designed to provide both a minimum royalty of an additional amount based on the profitability of each project. AQM will pay at least the minimum royalty rate of 1% of sales, regardless of its profitability.
Tax losses can be utilized in one of the following ways to get some relief for losses: 1) be carried forward for four consecutive years, beginning with the first subsequent year in which the losses arise or, 2) be carried forward indefinitely, but with an annual limit equivalent to 50% of the taxpayer’s taxable income of each subsequent year.
A 4.1% withholding income tax rate is applied to the dividends agreed to be remitted only if the shareholders are individuals or non-resident entities, which is not the case of CMZ.
An 18% Value Added Tax (IGV) rate is applied to purchases of capital, goods and services used in operations subject to IGV. It is assumed that the IGV paid by CMZ is to be recovered by means of the recovery regimes established under Peruvian legislation.
Mining companies are obliged to pay a worker´s profit sharing of 8% on the net profits of the company, taking into account that CMZ will have more than 20 workers. The total sum received by the worker cannot exceed an amount up to 18 times the monthly salary, after which the balance will be paid to a special educational, social and recreational fund. The amount paid is allowed as a tax deduction for corporate tax purposes.
At the long-term metal prices that we have used for this study, total estimated unescalated taxes and royalties payable on Zafranal profits are $1,377.8 million over the 15-year mine life. The components of the various taxes that will be payable are shown in Table 22.3.
Table 22.3 Components of the Various Taxes
LOM Amount Tax Component ($ million) Income Taxes 869.8 Worker’s Profit Sharing Taxes 252.1 Royalty and Special Tax 255.9 Total Taxes 1,377.8
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Base case post-tax financial results for Zafranal are summarized in Table 22.4.
Table 22.4 Summary of Post-tax Financial Results
Description Value Copper Price ($/lb) 3.00 Gold Price ($/oz) 1,274 Net Cash Flow (US$ million) 1,814 Discounted Cash Flow NPV ($ million) at 5% 927 Discounted Cash Flow NPV ($ million) at 8% 588 Discounted Cash Flow NPV ($ million) at 10% 414 Payback (years) 2.6 IRR (%) 17.4
22.6 R OYALTIES
No royalties were applied in the pre-tax financial analysis. Royalties applied in the post-tax analysis are listed in Section 22.5.
22.7 S MELTER T ERMS
The following smelting terms for copper concentrate were applied in the financial analysis based on the marketing study by Ayrmin:
• Copper – pay 100% of content less 1.0 unit at the London Metal Exchange (LME) price for Grade A copper less a refining charge of $0.0635/accountable pound. The refining charge is not subject to price participation. • Treatment Charge – $63.50/dmt of concentrate delivered. • By-product Credit (gold) – pay 90% with a refining charge of $4.00/oz for gold content of 1 to 5 g/t. • Penalty Charge – none.
22.8 T RANSPORTATION L OGISTICS
Transportation costs for the copper concentrate are listed below based on the marketing study by Ayrmin.
• trucking – $1.00/wmt per each 10 km • port storage and handling – $17.00/wmt • ocean transport to Asian port – $65.00/wmt
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• moisture content – 9% • marketing – $12.00/dmt of concentrate.
Transportation costs for the copper cathode are listed below:
• trucking – $1.00/t cathode per each 10 km • port and insurance – $10.00/t cathode • ocean transport to Rotterdam port – $67.00/t cathode • marketing – $5.00/t cathode.
22.8.1 INSURANCE
Insurance is included in the port cost.
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23.0 ADJACENT PROPERTIES
There are no material properties adjacent to the Property.
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24.0 OTHER RELEVANT DATA AND INFORMATION
24.1 P ROJECT E XECUTION P LAN
24.1.1 INTRODUCTION
Reference to the term “the Company” or “AQM” in this section refers to the mine operator during the period of construction.
The Project Execution Plan (PEP) presents how the Company will successfully complete the Project. The PEP specifies the Project approach, tasks, and schedule. As well, it identifies and addresses any unique challenges facing the Project.
The Project will be designed and constructed to industry and regulatory standards, with emphasis on addressing all environmental and safety issues. Adherence to the PEP will ensure timely and cost effective completion while maintaining construction quality.
24.1.2 PROJECT APPROACH
To achieve successful project execution, AQM will hire and contract a project management team (PMT). The PMT will comprise personnel with appropriate skills, knowledge, and experience and will carry out their duties with the support of multi- discipline consultants and contractors. The PMT will ensure the implementation of procedural checks and balances, progress monitoring, regulatory guidance, and quality assurance and control. Strategies for the EPCM will be based on the following segregation of responsibilities:
• contracted specialized EPCM for water supply • contracted specialized EPCM for power supply • mining contracted engineer PMT in charge of CM, mine equipment procurement and pre-mining operations • process, TMF and related infrastructure: contracted engineering and procurement PMT in charge of CM.
The project management organization chart is shown in Figure 24.1.
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Figure 24.1 Project Management Organization Chart
Note: Start-up & Commissioning Manager role will transfer to site for construction.
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PROJECT MANAGEMENT SYSTEM
A proven and integrated PMS will be utilized by the Contractor to provide precise and accurate information to the Contractor and AQM, enabling them to make decisions and implement actions for the successful execution of the Project. The PMS and project cost control system will provide reports on the status and progress of the Project, ensure there is documentation of scope changes, track the schedule, and monitor actual costs to budgets and forecasts. The PMS will also compare actual performance with planned activities and report the effect of anticipated changes on the final date and cost.
PROJECT CONTROLS PERSONNEL
An integral part of the PMS is the project controls function. The personnel assigned to this function will plan and control the schedule and costs of the Project by use of an integrated project control system, which will encompass the functions of scheduling, cost control, estimating, change control, monitoring and reporting for the engineering, procurement, construction, and pre-operational testing of the Project.
24.1.3 PROJECT EXECUTION SUMMARY
A well-managed plan will be initiated from the date that project execution begins. An effective PMS will be implemented to assist in managing project costs and scheduling. The team will ensure that:
• the critical path schedule of construction is met or improved upon • engineering and procurement activities are completed to support construction requirements • scope changes are reflected in the PMS on a timely basis • costs are monitored, controlled, and reported to the Company on a regular basis.
Within six months from the Project go-ahead, the following will be completed:
• award of the project management contract • project control structure including budget, schedule, procedures, and work plans • bidders lists • completed flowsheets and material balances • project procedures manual (PPM) • process design • final site layout
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• all design criteria including, but not limited to, environmental, applicable codes, materials of construction, and control philosophy • process equipment list with request for proposal (RFP) packages • the assignment of package contract numbers • modularizing, pre-assembly, and purchasing strategies • finalized contracting strategy • approved training program • contracts for early construction activities tendered, received, and evaluated • health and safety management plan (HSMP) • QA/QC • EMP • construction plan • all project management systems in place • all geotechnical and site survey data completed.
24.1.4 ENGINEERING
The detailed design engineering program will include all disciplines from geotechnical to computerized controls. Each discipline will utilize both recent technological advances and proven techniques as are appropriate for the Project.
Once the Company has authorized the Project to proceed, the Contractors will establish the engineering organization and assemble the necessary resources required to meet the Project requirements.
24.1.5 PROCUREMENT
Procurement of goods and services will adhere to the highest ethical standards and will be performed in a transparent manner. The procurement group will develop and implement procurement policies that:
• comply with project technical requirements • comply with the Health, Safety, and Environmental (HSE) policy • comply with legal and regulatory requirements • deliver goods and services to satisfy project schedule requirements • source materials and equipment within South America, where quality, price, and availability are competitive.
The procurement group will prepare procurement procedures and a procurement plan for the execution of the Project, including procedures for purchasing, inspection,
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progress monitoring, material control, expediting, batch crating and packaging, transshipping, consolidating, and transportation.
24.1.6 CONSTRUCTION MANAGEMENT
The CM group will be responsible for the management of all field operations. Reporting to the Company, the construction manager will plan, organize, and manage construction quality, safety, budget, and schedule objectives. The key CM objectives are:
• Conduct HSE policy training and enforcement for all site and contractor staff. Site hazard management tools and programs will be employed to achieve the no harm/zero accident objective. • Apply contracting and construction infrastructure strategies to support the Project execution requirements. • Develop and implement a construction-sensitive and cost-effective master project schedule. • Establish a project cost control system to ensure effective cost reporting, monitoring, and forecasting as well as schedule reporting and control. A cost trending programme will be instigated whereby the Contractor will be responsible for evaluating costs on an ongoing basis for comparison to budget and forecasting for the cost report on a monthly basis. • Establish a field contract administration system to effectively manage, control, and coordinate the work performed by the contractors. • Apply an effective field constructability program, as a continuation of the constructability reviews performed in the design office. • Develop a detailed field logistics and material control plan to maintain the necessary flow and control of material and equipment to support construction operations. • Meet the schedule for handover of the construction plan to the commissioning team. • Develop a QA/QC plan to set guidelines in terms of plant operability, safety of operation and adherence to all regulatory requirements.
The CM organization chart (Figure 24.2) shows the CM team for the Project.
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Figure 24.2 CM Organization Chart
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24.1.7 CONSTRUCTION SCHEDULE
SCHEDULE OF DEVELOPMENT
The first construction schedule will outline the activities that comprise the sequence of project development steps.
The construction schedule will then be expanded to include sub-schedules addressing specific activities and contracts. A preliminary schedule is presented in Figure 24.3.
CONSTRUCTION CONTRACTING
The contracting strategy will be designed to maximize utilization of the local labour force, create a responsible and sustainable relationship with the nearby communities, and provide the senior management and specialists to support the safety, quality, schedule, and cost objectives of the Project. In addition, contracts will be designed to combine timing, scope, battery limits, and contract value into manageable packages.
AQM will provide contractors and CM staff with:
• on-site, project-wide first aid services • project-wide security • locations for construction offices and equipment/material laydown • local electrical panels and temporary generator sets • water sources • diesel fuel • storage for construction equipment • sources for all concrete and structural aggregates • permanent bulk materials and all capital equipment • QA/QC audits • vendor-representative assistance.
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Figure 24.3 Preliminary Project Development Schedule Summary
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TEMPORARY FACILITIES AND CONSTRUCTION SITE INFRASTRUCTURE
Construction Accommodation
Development of a full-service camp for construction contractors will begin immediately following the receipt of construction permits. The construction camp will be leased for the Project development period.
The camp of modular design, will accommodate up to 900 workers on-site. Accommodation types will include double and single occupancy rooms. The off-site EPCM Contractors will provide accommodation for their workforces as required.
Transportation, potable water, waste management and other support services will be scaled to support the various development stages. The CM team will ensure that the catering contractor meets all facilities, staffing, hygiene, food handling, storage, and meal regulations and expectations.
Communication
AQM systems manager will determine the appropriate telecommunications technologies for the Project. Requirements include voice and data link technologies adequate to support construction phase and plant operation requirements.
Construction Power
Permanent power will supply all mine equipment and construction power loads for the duration of the construction phase. Some initial temporary power for construction activities will be required during the construction and energizing of the site power distribution network.
First Aid and Site Security
AQM will provide a fully-equipped first aid facility and ambulance for project-wide use. The facility will normally be staffed 12 h/d, with on-call services ensuring continuous coverage. The first aid staff will live at the camp. Contractors will be expected to provide basic first aid stations at the site.
AQM will supply a 24-hour staffed site security program during the initial field mobilization. Access to the site will be controlled at the principal road entrances and will be limited to personnel who have attended induction training, as well as approved visitors.
Warehousing
Construction warehousing will evolve with the Project. All freight delivered will be received at a temporary warehouse and stored there or in designated laydown areas.
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Laydown Areas
Rapid development of the laydown areas at the site is absolutely essential to tie-in with the arriving loads. All laydown areas will be clearly marked with sign posts and will be laid out in a grid system.
Concrete Batch Plant
The concrete batch plant will be managed by the General Contractor as a service to the Project and will be operated by the General or Site Services Contractor.
Water Supply and Treatment Plant
Water for construction purposes will likely be sourced from the Majes River which has also been permitted to supply water for the site drilling activities.
The construction project will require fresh water for the following:
• potable drinking water (bottled water) • truck washing • concrete batching • road dust control • fire water • building cleaning • washroom and cleaning purposes.
Sewage Treatment Plant
A modular, portable temporary sewage treatment plant (STP) will be among the first items shipped to the Project.
24.1.8 PRE-OPERATIONAL TESTING AND START-UP
PROCESS PLANT
When construction is complete on any process unit, the Contractor will turn over responsibility to the Commissioning Manager for pre-operational testing and turnover of the facility to AQM prior to introducing mineralized material into the plant for commissioning and start-up.
AQM’s operating personnel will be involved in the pre-operational testing phase to the extent that they will progressively accept responsibility for sections of the plant as they are checked and handed over.
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Pre-operations testing of equipment will begin once the equipment items have been delivered to site, erected, and tested by the vendors’ engineers.
The pre-operational testing phase for the process facilities will include all aspects of dry mechanical and electrical testing of equipment and water testing of process equipment, including pressure testing of pipework and wet pre-operational testing to the extent practicable.
Visual inspection and pre-operational testing (yellow tag) will occur upon completion of installation of plant and equipment where the construction contractor will submit one copy of the appropriate pre-operational check forms, which will notify AQM that the plant and equipment are ready for inspection and the following have been put into effect and/or completed.
The following procedure and tagging system will be adopted in the execution of assignments as work is being completed.
Visual Inspection
Visual inspection is the non-operational examination of an installation to check that it is in accordance with the engineer’s and the manufacturer’s drawings, specifications, and manuals.
Pre-operational Test
A pre-operational test is the initial no-load test of a piece of equipment with test media such as water or air where required.
Checkout and Acceptance (Green Tag)
This procedure allows for the transfer of responsibility from the Contractor to AQM. This procedure establishes that the installation of the equipment and ancillaries has been completed in accordance with the Contractors’ drawings, specifications, and codes and the equipment has been energized to prove its readiness for the process commissioning and start-up. On acceptance, AQM assumes responsibility for operation and maintenance.
START-UP
Start-up (introduction of mineralized material) is performed under the direction of AQM’s Start-up Manager, and involves a select staff from pre-operations, process specialists, and AQM’s operating personnel. This will be the beginning of operations under load conditions and the systematic increase in capacity until process through- put and recovery requirements are met and sustained.
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25.0 INTERPRETATION AND CONCLUSIONS
25.1 R ISK A NALYSIS
INTRODUCTION
Since the inception of the Project planning process, AQM has developed a comprehensive risk management plan and has been actively working to address and mitigate key areas of potential risk identified for the Project.
RISK CONTROL PLAN
A number of factors have been identified by AQM as potential risk factors, and each is being dealt with through the application of a holistic project planning philosophy designed to mitigate, and where possible, successfully resolve potential risk factors as the Project moves towards development.
IDENTIFIED RISKS
Social/Political
Current social unrest presents a significant risk for any major mining project in Perú, especially when that development affects a water source previously utilized for agricultural purposes. In response to these concerns, AQM has undertaken an extensive community and stakeholder relations program at local and regional levels, in an effort to fully communicate with local communities and authorities, and to reduce and mitigate the potential political and social risks that could affect the Project. Based on the results of the ongoing programs, dialogue at this level has been very positive.
The inability to obtain property access or to purchase or lease the lands on which the project would be developed also poses a significant risk for the mine development. Right of ways are secured through negotiations with the land owners, which in the case of the Property is a single regional government agency, Autodema. Recently, AQM was able to renew and expand the surface access agreement with Autodema that was originally signed in 2009. This annual renewable agreement allows for continued exploration activities within the Property and may, under certain conditions, be modified to allow for future development of the Project.
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Environmental and Permitting
Water in the region is scarce and primarily used for agriculture and human consumption. For a PEA level of study, AQM has incorporated and costed the use of desalinated seawater as the water supply option for the Project; however, the Project team is actively talking to authorities and local communities to find alternative arrangements that would result in benefits for local communities in exchange for the use of local water sources by the Project.
There are potential environmental impacts associated with the Project, including:
• risk of an unplanned breach of an embankment causing release of tailings to the environment • seepage of groundwater • dust and noise • storm containment and unplanned TMF embankment failure • acid rock generation in waste dumps, abandoned leach pads and tailings disposal area.
These potential impacts should be studied in greater detail in the future and contingency plans put in place to reduce risk.
In Perú, the mining industry is expected to engage in thorough project due diligence prior to being granted permits and the necessary license-to-operate. AQM has undertaken the appropriate steps with respect to the environmental studies required to develop the plans necessary to reduce risk, and satisfy government requirements.
AQM intends to ensure that all environmental management aspects are looked after proactively and maintain that attitude throughout the operation of the mine.
Tetra Tech recommends that AQM carefully and proactively manage all environmental impacts associated with the Project, to minimize environmental risks and costs.
Permitting timelines that become longer than originally scheduled and more stringent environmental, safety and government regulations present high risk potential for the Project. CMZ has based its general manager in the regional capital, Arequipa, to ensure frequent and productive contact with regional authorities, as well as maintain a presence in Lima. This will enable the management team to detect difficulties or new requirements for obtaining permits early enough to be able to adjust the Project activities before incurring additional costs and delays.
Business
The risks associated with escalating capital and operating costs due to changing local and global economic conditions are potential high risk factors for the Project.
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Appropriate levels of engineering and project execution planning will allow the Project to adequately mitigate these cost risks.
Technical
The proposed Project does not pose any major challenging technical risks for its development. If constructed, the desalination plant and pipeline from the Peruvian coast will require specialized expertise for construction and operation, and AQM has planned for these tasks to be performed by an EPCM contractor knowledgeable in these types of projects, who could also be retained to operate the system during production.
Adequate metallurgical test work and engineering will be required to ensure the proper sizing of process equipment. The logistics of moving personnel and materials during the construction phase will require careful planning and early road construction to ensure the project stays on schedule. The movement of copper concentrate during production to the selected port site will be challenged by road congestion and further studies will be required in the next phase of project design to better mitigate this risk.
25.2 G EOLOGY
The Property contains several porphyry-type copper occurrences, of which four have been explored by surface sampling, geophysics and drilling: Zafranal Main, Victoria, Sicera Norte and Sicera Sur.
Geological work to date on the Property has confirmed that the occurrence of mineralized porphyries is structurally controlled by a northwest trending series of strike-slip faults belonging to the Incapuquio fault system, and regional east-west trending structures. The intersections of these two fault systems appears to be the main mineralization control on the Property.
Mineralization is hosted in several lithologies, particularly in the Zafranal Diorite and the Microdiorite, with lesser supergene mineralization hosted in volcanic rocks and younger cross-cutting intrusives. Additional exploration targets have been explored along both main structural trends, with geophysical and geochemical anomalies found on some of them. Drilling is required on these targets to confirm the presence of additional mineralization.
25.3 M INING
The summary of the mining plan is provided in Table 25.1.
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Table 25.1 Mining Plan
Flotation ROM Leach Cu Au Cu Au Mineralization Grade Grade Mineralization Grade Grade Types '000 t (%) (g/t) Types '000 t (%) (g/t) Supergene 135,762 0.626 0.081 Oxide 10,699 0.221 0.134 Hypogene 206,764 0.291 0.075 Leached 46,947 0.227 0.109 Transition 8,081 0.261 0.034 Mixed 6,336 0.271 0.025 Hypogene Low Grade 74,704 0.178 0.048 Supergene 20,663 0.238 0.032 Total 425,311 0.378 0.071 - 87,256 0.230 0.085
The Company (the mine operator) may elect to use its own workforce to carry out all mining activities with the exception of the MARC for all mining equipment and a contract blasting service. The mine operator proposes to use mining equipment to construct all external pit haul roads, to rough grade crusher and mine shop areas and to supply the fill required to construct the canyon crossing to access the crusher dump pocket.
The use of an earthmoving contractor should be considered for some of this work if this can be accomplished cost effectively as this would improve the probability of successfully meeting the pit stripping requirements.
The mine design takes full advantage of the local topography for the development of waste dumps and primary crusher location, and this is reflected in the mine operating costs. Over 84% of the mine haulage will be flat or downhill over the LOM.
The mining costs in Table 25.2 appear in line with costs of other mining operations in the area of the Project.
Table 25.2 Mining Pre-production and Production Costs
Pre-production $ million $/t mined Loading 18.6 0.16 Hauling 69.8 0.60 Drilling 9.2 0.08 Blasting (Contract Service) 17.3 0.15 Ancillary 20.9 0.18 Support 2.7 0.02 Indirects 13.2 0.11 Subtotal 151.7 1.30 Contingency 20% 30.3 0.26 Total Unit Costs 182.0 1.56
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Production $ million Unit 15-Year LOM Mining Operation 1,312.2 $/t mined $1.40 - $/t moved $1.29 Sustaining Capital 61.1 - -
25.4 C ONCENTRATOR
It was generally concluded that the proposed concentrator feed did not present any significant technical difficulties for beneficiation and that the concentrator would resemble similar concentrators processing copper porphyry mineralization of average characteristics.
The current grinding configuration considers wrap-around drives on a single 40 ft diameter SAG mill and two 25 diameter ball mills, which maximizes concentrator flexibility and operational performance. An alternative circuit with two smaller pinion- drive SAG mills with four smaller pinion-drive ball mills may be a cheaper investment option and should be studied in the next phase of engineering.
The current concentrator is an open-air structure with overhead cranes over the grinding circuit and primary crusher. An alternative arrangement with the overhead cranes being replaced with mobile cranes servicing the grinding and crushing areas may be a cheaper investment option and should be studied in the next phase of engineering.
25.5 L EACH F ACILITY
The leachable material must be removed as part of the pre-production stripping in preparation for the concentrator start-up. As a result, the cost of transporting the leachable material to the leach pad is the incremental cost above the cost of sending the material to the waste dump. This incremental cost has been determined to be $0.274/t sent to the leach pad or $24 million over the LOM. This incremental cost could be reduced or eliminated if a suitable alternative could be found with a favourable haul profile. Such a location is available but will require that the canyon crossover to access the crusher dump pocket be converted into a dam that would capture any water from a significant storm event.
This potential alternative leach facility location should be studied in the next phase of engineering. Additional leach metallurgical test work will also be required to determine the appropriate copper recoveries for the various mineral types.
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25.6 T AILINGS AND W ATER M ANAGEMENT
Knight Piésold completed the design and cost estimates for the tailings and water management components of the Zafranal conceptual level study.
The TMF design is based on the site characteristics, tailings properties, certain design standards and parameters, the geological and geotechnical conditions at the site, and the appropriate engineering analyses. The geological and geotechnical conditions at the TMF are based on general knowledge of the area and no specific geotechnical investigations have been completed.
The current disposal density for the tailings is estimated to be 65% solids (by weight) with no water reclaim. Dam construction will be with borrowed material as insufficient water will be available to cyclone tailings. Relocating the TMF closer to the plant would save capital and operating costs and open the possibility for further thickening of the tailings prior to disposal. These possibilities should be studied in the next level of engineering study.
25.7 O WNER’ S C OSTS
The Owner’s cost estimate was developed to cover the costs associated with the pre-production operating activities as well as the recruitment, training, transportation, accommodation, and catering for the employee and contractor workforce where appropriate. The production phase of the Project will require an estimated 509 employees and 328 contractors for a total of 837 personnel. It is envisaged that the bulk of the workforce will reside in the communities closest to the Project. Administrative staff and management will be based in Arequipa and reside on site during their weekly shift in a 233-man permanent camp. The study contemplates shift schedules of 7-days-on, 7-days-off for hourly employees and 4-days-on, 3-days- off for administrative staff and management. Peruvian law provides for alternative shift systems which significantly reduce manning requirements and these should be studied in the next phase of engineering.
25.8 E NVIRONMENTAL
In Perú, the mining industry is expected to engage in thorough project due diligence prior to being granted permits and the necessary license-to-operate. AQM has undertaken the appropriate steps with respect to the environmental studies required to develop the plans necessary to reduce risk and satisfy government requirements.
Tetra Tech recommends that AQM carefully and proactively manage all environmental impacts associated with the Project, to minimize environmental risks and costs.
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26.0 RECOMMENDATIONS
26.1 S UMMARY
Based on the results of the work presented in this report, Tetra Tech recommends that AQM proceed with the next phase of work to identify potential cost saving and additional opportunities in order to assess the viability of the Project.
As listed in Table 26.1, the recommendations described in this section have been estimated to require an expenditure of $9.5 million, if implemented.
Table 26.1 Recommended Future Work
PEA Recommendations US$ Expanded CRP 500,000 Surface Right of Ways 50,000 Social Studies Update 50,000 Environmental Management Plan 115,000 Closure Plan 115,000 Archeological Site Investigation 200,000 Acid Rock Drainage 200,000 Mine Plan Review 300,000 Exploration Drilling (7,000 m HQ) 2,000,000 Concentrator Design 360,000 Metallurgical Test Work (Flotation and Leach) 1,620,000 Overall Site Design 100,000 Concentrate Handling 100,000 Product Transportation 138,000 Leach Plant Design 300,000 TMFs 1,500,000 Desalination Water Supply Studies 300,000 Power Supply and Distribution 288,000 Subtotal 8,236,000 Contingency (15%) 1,235,400 Total 9,471,400
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26.2 R ISK A NALYSIS
26.2.1 SOCIAL/POLITICAL
Risk management activities should continue in the next phase of project development. The CRP should be expanded to include stakeholders potentially impacted by the construction and operation of the desalination plant and pipeline, and the construction of the main power line from Arequipa.
The estimated cost of this expanded CRP will be $500,000/a and a follow-up survey and report will cost $50,000.
As soon as marine test work can confirm the suitability of the proposed location for the desalination plant and pipeline, AQM will need to secure the surface rights required for these facilities, which is estimated to cost $50,000. This amount refers only to the cost of the due diligence and valuation of the land property.
26.2.2 ENVIRONMENTAL
Given the TMF’s location upstream of the community of Corire, AQM may wish to consider engaging an external expert during the detailed design stage to review the TMF design, as recommended by IFC Guidance Note 5.
An updated mine closure plan and reclamation will be required for the Project.
Long-term physical and geochemical stability, protection of the downstream environment and management of surface water are requirements for the TMF at closure.
There are various pieces of legislation that need to be adhered to for mining activities in Perú, such as environmental and tailings management, mine closure, and environmental guidelines. The preparation of the EIA must follow Articles 47, 48, 49 and 68 of the Ley del SEIA regulations.
Aside from reconnaissance and prospecting activities that do not involve environmental impacts, all other exploration and mining activities in Perú must comply with the Sistema de Concesión Minera or Mining Concessions System, governed by the General Mining Law. The owner holding title to a mining concession has to first obtain the necessary permits and approvals in order to perform any exploration or mining activities.
An Autorización para Comenzar la Explotación Minera or Authorization to Begin Mining Exploitation is required in order to start mining operations. The Plan de Minado or Blasting Plan must be reviewed prior to commencing mining exploitation activities as per the Reglamento Interno de Seguridad e Higiene Minera or Internal Regulations for Mining Health and Safety.
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The Plan de Manejo Ambiental or Environmental Management Plan for the mine should include a manager and environmental technicians to perform all monitoring activities at the mine, including the tailings facility. Quarterly groundwater seepage reporting is required as well as annual reporting.
The recommended environmental work includes:
• Environmental Management Plan: $115,000 • Closure Plan: $115,000 • Archeological Site Investigation: $200,000 • Acid Rock Drainage: $200,000.
26.3 M INING
Values used to determine the mining portion of the project investment were referenced from quotations obtained for recent mining projects developed by NCL. In the next stage of the engineering, site specific quotations for the proposed mining equipment and supplies are recommended for better definition of capital and operating costs.
A more refined strategy for equipment maintenance is recommended in the next phase of engineering to determine the equipment packages that will be maintained by contractors (the MARCs) or by the owner´s workforce. Formal quotations should be obtained for the MARCs. Furthermore, the truck manufactures should be required to provide detailed haulage simulations for the proposed profiles including truck speeds and fuel consumption.
The cost of executing these recommendations will be equivalent to the cost of an engineer required to manage the requests and report on the data collection, which is estimated to be $50,000.
Tetra Tech recommends that AQM also develop mining plan scenarios focusing on the higher-grade supergene mineralization present in the upper reaches of the Main Zone deposit, with a view to increasing the average feed grade to the concentrator and lowering the overall capital investment for the Project. The estimated cost of this mining plan exercise is $250,000.
26.4 G EOLOGY
It is recommended that exploration drilling (7,000 m) focus on areas to the north of the Main Zone, where recent exploration has shown anomalous surface values. Exploration drilling should also focus on those areas to the south of the Main Zone, where geophysical and geochemical surveys have detected anomalous areas.
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The QA/QC protocols that have been established for the Project should be monitored as results are received and any analyses of control samples that exceed acceptable bounds should be investigated and if necessary, samples associated with the out-of- bounds control samples should be re-analyzed.
The budget for the drilling is estimated at an all-in cost of US$2 million.
26.5 C ONCENTRATOR
Although the process design as described is relatively detailed in many respects, the following recommendations, estimated to cost $360,000, should be considered for the next phase of the Project.
26.5.1 WATER MANAGEMENT
A conceptual layout of the process plant has been completed and is presented in the report. However, a final layout arrangement was not confirmed given the early status of the Project. The final layout arrangement will ultimately affect the design of the water management system, including the fire water distribution which has not been defined in detail, nor the thickener overflow solutions and their destinations regarding whether pumpboxes and pumps are required, or whether gravity flow only can be utilized. Similarly, the details regarding the fresh water sytem, including the desalinated water system, have also not been finalized to date.
26.5.2 CLEANER FLOTATION CIRCUIT CONFIGURATION
The cleaner flotation circuit configuration and the regrinding of the rougher concentrate should be regarded as conceptual only at this stage of this PEA report. The cleaner flotation circuit has not been optimized and will certainly be changed for the next phase of the Project. It is anticipated that more laboratory test work results will then be available regarding the behaviour and characteristics of the cleaner scavenger circuit configuration, and, in particular, the regrind product particle size, and the number and/or types of cleaner stages and flotation cells used. If confirmed by test work, a coarser regrind size will affect the selection of the type and size of regrind mill which will be used in the design.
26.5.3 CYCLONE OVERFLOW/ROUGHER FLOTATION FEED
The two primary cyclone overflow products are shown as being delivered to the rougher flotation circuit and combined in a collection box for sampling purposes. The flotation feed is then split into two streams again with each stream feeding a rougher flotation line of cells by gravity flow. This configuration is also layout-related and depends on the available elevation. It will also require confirmation in the next stage of the Project regarding the gravity flow concept, and possibly whether the two
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streams should be kept separate from the cyclone overflow to its respective rougher flotation line.
26.5.4 WATER TREATMENT PLANT
A water treatment plant has been included on the flowsheet, but no design details have been shown. However, at this stage of the Project, the water treatment plant is conceptual only and no details regarding the flowrates and objectives of the water treatment plant are known.
26.5.5 PRIMARY GRINDING CIRCUIT
The primary grinding circuit has been conservatively designed for this phase of the Project, particularly with regard to the SAG mill size and power requirements. It is anticipated that the results of additional grinding test work being planned at present will be available for the next phase of the Project which will enable the grinding circuit to be defined to a geater degree of accuracy.
The grinding circuit also requires a review of the maximum possible treatment rate, and its impact on the rest of the plant design. This may afford an opportunity to treat in excess of 80,000 t/d with concomitant financial benefits.
26.5.6 COPPER RECOVERY
During the next phase of the Project, and subject to confirmation of the design copper recovery, it is possible that the overall copper recovery will be reduced by up to 1.5% to accommodate the effect of plant scale-up, as well as the effect of commissioning.
26.6 F UTURE M ETALLURGICAL T EST W ORK
The following minimum additional test work will be required to assist with the design of both the concentrator and leach facility for the next phase of the Project:
• a detailed grinding investigation • crushability index and abrasion index tests • variability test results • flotation reagent types to be defined • size distribution of rougher flotation concentrate products versus primary grind size to be established • optimal regrind particle size to be established • the effect of chalcocite on the copper concentrate grade to be quantified
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• the impurity elements and the respective concentrations present in the copper concentrate • settling data for copper concentrate to be defined • filtration data for copper concentrate to be established • settling data for tailings material to be quantified • bottle roll and column leaching of mineral types • variability testing for leach recoveries • solvent extraction testing • acid consumption.
The estimated cost of this program total $1,620,000.
26.7 C ONCENTRATE H ANDLING T RADE- O FF
At the next level of study, the use of a concentrate transfer pipeline will be explored to determine the viability of transporting slurry downhill from the mill to the load out area. This will be conducted in a trade-off study that will compare this option to transporting the concentrate from the mill via truck to a loading area. Possible optimization may include a reduction in road development costs. The study is estimated to cost $100,000.
26.8 P RODUCT T RANSPORTATION
Road congestion will impact product transportation to the port of Matarani and alternatives should be studied to mitigate this impact. One option that appears attractive is a combination truck and rail haulage, similar to the system employeed by Cerro Verde which ships via the same port. The transport study and rail loadout design is estimated to cost $138,000.
26.9 L EACH F ACILITY
As mentioned in the previous section, an effort should be made to relocate the leach facility to minimize additional haulage costs and improve the space available for facilities. An alternative has been identified to carry out this trade-off study and confirm the new location, at an estimated cost of $300,000.
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26.10 S ITE L AYOUT
Although the site layout is currently very strong from an engineering perspective, the design will likely receive further refinement and optimization at the next level of study. An allowance of $100,000 has been set aside for this work.
26.11 T AILINGS M ANAGEMENT F ACILITY
Recommendations for the next phase of engineering for the TMF area of the Project are summarized as follows:
• Limited meteorologic data was available for this study. It is recommended that site specific hydrometeorology, research and regional stream monitoring data be collected to better estimate the hydrological characteristics of the site. • Consolidation and strength testing on the tailings should be completed to provide essential inputs for large strain consolidation modelling. • Tailings rheology testing should be completed to determine the fluid flow characteristics of the tailings. This is required to more accurately determine the pumping and pipeline designs. • A tailings disposal options assessment should be completed during future studies to confirm the most suitable method for tailings disposal. • Environmental testing of the tailings products should be completed and assessed to confirm if any design modifications are required. • Site investigations should be conducted to confirm the characteristics of the basin and embankment foundations. The starter embankments will be constructed from materials removed from within the basin and detailed material properties are required to confirm the suitability and availability of the materials. • Detailed engineering analyses should be completed incorporating the new or updated information. • The cost estimates should be reviewed and confirmed because prices are fluctuating significantly at the current time. In addition, there may be opportunities to refine costs for certain items with additional knowledge, possibly by conducting trade-off studies as appropriate.
This work is estimated to cost $1.5 million.
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26.12 W ATER S UPPLY
The desalinated water supply option has significant capital and operating cost implications and AQM is encouraged to explore other potential water sources through discussions with local farming communities. Any arrangement would require a trade-off study which should be completed during the prefeasibility study stage to definitively determine the optimum water delivery methodology for the Project.
If the decision is to proceed with the desalinated water supply, then marine studies will need to be undertaken to confirm the suitability of the proposed location for the desalination plant, and detailed design of the pipeline route will also be required in the next phase of engineering. The estimated cost of this work is $300,000.
26.13 P OWER S UPPLY AND D ISTRIBUTION
Extending grid electrical power to the site is on the critical path for achieving the pre- production stripping schedule of 117 Mt. Knight Piésold recommends that additional engineering is undertaken during the next phase of the Project to ensure that the necessary environmental permitting is completed on schedule so as to permit an early start of the power line construction. The cost of this early engineering work is estimated to be $288,000.
26.14 E NVIRONMENTAL
There are various pieces of legislation that need to be followed for mining activities in Perú, such as environmental and tailings management, mine closure, environmental guidelines. The preparation of the EIA must follow articles 10, 47, 48 and 49 of the National Environmental Impact Assessment System Law Regulation.
Aside from reconnaissance and prospecting activities, all other mining activities in Perú must comply with the Sistema de Concesión Minera or Mining Concessions System, governed by the General Mining Law.
An Autorización para Comenzar la Explotación Minera or Authorization to Begin Mining Exploitation is required in order to start mining operations. The Plan de Minado or Blasting Plan must be reviewed prior to commencing mining exploitation activities as per the Reglamento Interno de Seguridad e Higiene Minera or Internal Regulations for Mining Health and Safety.
The Plan de Manejo Ambiental or Environmental Management Plan for the mine should include a manager and environmental technicians to perform all monitoring activities at the mine, including the tailings facility. Quarterly groundwater seepage reporting is required as well as annual reporting.
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Given the TMF’s location upstream of the community of Corire, AQM may wish to consider engaging an external expert during the detailed design stage to review the TMF design, as recommended by IFC Guidance Note 5.
An updated mine closure plan and reclamation will be required for the Project.
Long-term physical and geochemical stability, protection of the downstream environment and management of surface water are requirements for the TMF at closure.
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27.0 REFERENCES
GEOLOGY
Alván, Aldo; Acosta, Harmuth, (2008). The Proterozoic Basement of the Arequipa Massif, Southern Peru: Lithologic Domains and Tectonics 7th International Symposium on Andean Geodynamics (ISAG 2008, Nice), Extended Abstracts: 549-552.
AMEC Minproc (February 25, 2011). Zafranal Copper Project, Peru, Technical Report December 2010 Resource Estimate.
Cox, Dennis P. and Singer, Donald A., editors, (1985). Mineral Deposit Models United States Geological Survey Bulletin 1963, pages 76-79.
Panteleyev, A. (1995): Porphyry Cu-Au: Alkalic, in Selected British Columbia Mineral Deposit Profiles, Volume 1 - Metallics and Coal, Lefebure, D.V. and Ray, G.E., Editors, British Columbia Ministry of Energy of Employment and Investment, Open File 1995-20, pages 83-86.
Rivera, Fernando; León, Jorge; Cano, Oscar y Huamán, Moisés, (2010). Controles de Mineralizacion en el pórfido de Cu Zafranal, en el sur del Perú Minera AQM Copper Perú SAC.
TAILINGS MANAGEMENT FACILITY
Canadian Dam Association (CDA) Dam Safety Guidelines, 2007.
Earthquake Spectra (2008), “Special Issue on the Next Generation Attenuation Project”, Vol. 24, No. 1.
ICOLD - International Commission on Large Dams, (1995), “Tailings Dams and Seismicity: Review and Recommendations,” Bulletin 98.
International Commission on Large Dams (ICOLD), Tailings Dams and Seismicity: Review and Recommendations, Bulletin 98, 1995.
Machare, J., Fenton, C.H., Machette, M.N., Lavenu, A., Costa, C. and Dart, R.L., (2003), “Database and Map of Quaternary Faults and Folds in Peru and its Offshore Region”, USGS Open-File Report 03-451.
Saragoni, G.R., Astroza, M. And Ruiz, S., (2004), “Comparative Study of Subduction Earthquake Ground Motion of North, Central and South America”, Proceedings of
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the 13th World Conference on Earthquake Engineering, Vancouver, B.C., Canada.
Schlumberger Water Services, Zafranal Water Supply Technical Memorandum, July 19, 2011 (Ref. No. 50555)
Tanner, J.G. and Shedlock, K.M., (2004) “Seismic Hazard Maps of Mexico, the Caribbean, and Central and South America”, Tectonophysics 390, p. 159-175.
The Mining Association of Canada (MAC) Guide to the Management of Tailings Facilities, 1998.
Youngs, R.R., Chiou, S.-J., Silva, W.J. and Humphrey, J.R. (1997) “Strong Ground Motion Attenuation Relationships for Subduction Zone Earthquakes”, Seismological Society of America, Seismological Research Letters, Vol. 68, No.1, p.58-73.
ENVIRONMENTAL
Apoquindo Minerals. Proyecto Zafranal. Términos de Referencia – Línea Base Ambiental y Social. Marzo 2010
CESEL Ingenieros. Informe Técnico CSL-111500-11-IT-01. Estudio de la Línea Base Ambiental y Socioeconómica para el Proyecto Minero Zafranal, Minera AQM Copper Perú S.A.C. Plan de Trabajo y Selección de Puntos de Monitoreos para la elaboración de la Línea Base Ambiental y Socioeconómica para el Proyecto Minero Zafranal. Abril 2011
CESEL Ingenieros. Primer Informe Semestral. Línea Base Ambiental y Socioeconómica para el Proyecto Minero Zafranal, Minera AQM Copper Perú S.A.C. Noviembre 2011
CSA Consultores. Social Management Report – Enero-Abril 2012
Delapuente Abogados. AQM – EIA Proyecto Zafranal. Documentación Legal y Técnica Requerida. Octubre 2011
Knight Piésold Consulting. Conceptual Level Environmental Management Design – Tailings Management Facility – Zafranal Project. May 2012
Knight Piésold Consulting. Conceptual Level Closure Plan – Tailings Management Facility – Zafranal Project. May 2012
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MINERAL PROCESSING
Teck, 2005. Mineralized Material Characterization Study.
Amdel Mineral Laboratories, 2010. Phase I – Flotation Optimization, Flotation Variability Testing, Leaching, Mineralogical Identification, Sizing Analysis. May to August 2010.
JKTech Pty Ltd, 2010. SMC Test on 15 Samples. April 2010.
SGS, 2010. Abrasion Tests.
Philips Enterprises, 2010. Impact Tests. May 2010.
Amdel Mineral Laboratories, 2011. Phase II – Communition Tests, Locked Cycle and Variability Flotation Tests, Leach Amenability Tests. March to April 2011.
SGS, 2011. Bond, Abrasion and SAG Mill Comminution Tests. January 2011.
AMEC Minproc Limited, 2010. Communition Test Update. August 2010.
JKTech Pty Ltd, 2011. SAG Mill Comminution Test Report on 42 Samples from the Project. January 2011.
AMEC Minproc Limited, 2011. Technical Report December 2010 Resource Estimate. Document No. 60246-00000-23-002-001. February 2011.
AMEC Minproc Limited, 2012. Zafranal Copper Project – Stage 2 Flotation Testwork. Document No. 60246-00000-21-002-003. May 2012.
AMEC Minproc Limited, 2012. Zafranal Copper Project – Dewatered Tailings Testwork. Document No. 60246-00000-21-002-004. May 2012.
AMEC Minproc Limited and Amdel Mineral Laboratories, 2012. Zafranal Copper Project – Leaching Test Work. Document No. 60246-00000-21-002-005. May 2012.
G&T Metallurgical Testing Services Ltd. and Plenge Laboratories, 2011 and 2012. Phase III Test. Individual laboratory reports not issued at the time of this report.
Transmin Metallurgical Consultants, 2012. TM 621 Zafranal Metallurgical Testwork Report for AQM Copper Inc. June 12, 2012.
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28.0 CERTIFICATES OF QUALIFIED PERSON
MARINUS ANDRE DE RUIJTER, P.ENG.
I, Marinus Andre De Ruijter, P.Eng., of Delta, British Columbia, do hereby certify: • I am a Principal Metallurgical Engineer with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia, V6B 1M1. • This certificate applies to the technical report entitled “Technical Report and Preliminary Assessment of the Zafranal Project, Peru”, dated January 16, 2013 (the “Technical Report”). • I am a graduate of the University of Witwatersrand, Johannesburg, South Africa (M.Eng., 1979). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (#31031). My relevant experience with respect to this project includes copper and polymetallic base metal sulphide mineral flotation projects and their development, and plant audits. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”). • My most recent personal inspection of the Property was November 29, 2011 for one day. • I am responsible for Sections 1.6, 1.10.2, 13.0, 17.0, 21.2.1, 21.2.3, 21.2.5, 21.2.6, 21.2.7, 25.4, 26.5, 26.6, 26.7, 26.9, 27.0 (mineral processing section only) and 28.0 of the Technical Report. • I am independent of AQM Copper Inc. as defined by Section 1.5 of the Instrument. • I have no prior involvement with the Property that is the subject of the Technical Report. • I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument. • As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Signed and dated this 16th day of January, 2013 at Vancouver, British Columbia
Original document signed and sealed by Marinus Andre De Ruijter, P.Eng. Marinus Andre De Ruijter, P.Eng. Principal Metallurgical Engineer Tetra Tech WEI Inc.
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GREGORY Z. MOSHER, P.GEO.
I, Gregory Z. Mosher, P.Geo., of North Vancouver, British Columbia, do hereby certify: • I am a Senior Geologist with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia, V6B 1M1. • This certificate applies to the technical report entitled “Technical Report and Preliminary Assessment of the Zafranal Project, Peru”, dated January 16, 2013 (the “Technical Report”). • I am a graduate of Dalhousie University, (B.Sc. Hons., 1970) and McGill University (M.Sc. Applied, 1973). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (#19267). My relevant experience with respect to porphyry deposits includes over 20 years of exploration for and evaluation of such deposits in a wide variety of geological settings. Additionally I have conducted several resource estimates of porphyry (copper, copper-gold, molybdenum) deposits over the past six years. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”). • My most recent personal inspection of the Property was October 11 and 12, 2011 for two days. • I am responsible for Sections 1.2, 1.3, 1.4, 1.5, 4.0, 5.0, 6.0, 7.0, 8.0, 9.0, 10.0, 11.0, 12.0, 14.0, 23.0, 25.2, 26.4, 27.0 (geology section only), and 28.0 of the Technical Report. • I am independent of AQM Copper Inc. as defined by Section 1.5 of the Instrument. • I have no prior involvement with the Property that is the subject of the Technical Report. • I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument. • As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Signed and dated this 16th day of January, 2013 at Vancouver, British Columbia
Original document signed and sealed by Gregory Z. Mosher, P.Geo. Gregory Z. Mosher, P.Geo. Senior Geologist Tetra Tech WEI Inc.
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HASSAN GHAFFARI, P.ENG.
I, Hassan Ghaffari, P.Eng., of Vancouver, British Columbia, do hereby certify: • I am a Manager of Metallurgy with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia, V6B 1M1. • This certificate applies to the technical report entitled “Technical Report and Preliminary Assessment of the Zafranal Project, Peru”, dated January 16, 2013 (the “Technical Report”). • I am a graduate of the University of Tehran (M.A.Sc., Mining Engineering, 1998) and the University of British Columbia (M.A.Sc., Mineral Process Engineering, 2004). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (#30408). My relevant experience with respect to mineral process engineering includes 22 years of experience in mining and plant operation, project studies, management, and engineering. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”). • My most recent personal inspection of the Property was November 29, 2011 for one day. • I am responsible for Sections 1.1, 1.8, 1.10.1, 1.12, 1.13, 2.0, 3.0, 18.1, 18.2, 18.3, 18.5, 18.6, 18.7, 18.9, 21.1.1, 21.1.2, 21.1.3, 21.1.4, 21.1.5, 21.1.6, 21.1.7, 21.1.9, 21.1.11, 21.1.13, 24.0, 25.1, 25.5, 25.7, 26.1, 26.8, 26.10, 26.12, and 28.0 of the Technical Report. • I am independent of AQM Copper Inc. as defined by Section 1.5 of the Instrument. • I have no prior involvement with the Property that is the subject of the Technical Report. • I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument. • As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Signed and dated this 16th day of January, 2013 at Vancouver, British Columbia
Original document signed and sealed by Hassan Ghaffari, P.Eng. Hassan Ghaffari, P.Eng. Manager of Metallurgy Tetra Tech WEI Inc.
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MONICA DANON-SCHAFFER, PH.D., P.ENG.
I, Monica Danon-Schaffer, Ph.D., P.Eng., of Vancouver, British Columbia, do hereby certify: • I am a Manager, Environment and Sustainable Development with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia, V6B 1M1. • This certificate applies to the technical report entitled “Technical Report and Preliminary Assessment of the Zafranal Project, Peru”, dated January 16, 2013 (the “Technical Report”). • I am a graduate of the University of British Columbia (Ph.D, Chemical and Biological Engineering, 2010), University of British Columbia (M.Eng. Civil Engineering, 1993), and the Universidad Iberoamericana (B.Sc. Chemical Engineering, 1988). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (#22768), Ontario (#100045213), Saskatchewan (#20307), Yukon (#1655), Northwest Territories/Nunavut (#L2018). My relevant experience includes preparation of environmental risk assessments, evaluation of Equator Principles, environmental health and safety compliance audits and due diligence at mines, mine waste and management. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”). • I did not complete a personal inspection of the Property. • I am responsible for Sections 1.9, 20.0, 25.8, 26.2, 26.14, 27.0 (environmental section only), and 28.0 of the Technical Report. • I am independent of AQM Copper Inc. as defined by Section 1.5 of the Instrument. • I have no prior involvement with the Property that is the subject of the Technical Report. • I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument. • As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Signed and dated this 16th day of January, 2013 at Vancouver, British Columbia
Original document signed and sealed by Monica Danon-Schaffer, Ph.D., P.Eng. Monica Danon-Schaffer, Ph.D., P.Eng. Manager, Environment and Sustainable Development Tetra Tech WEI Inc.
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SABRY ABDEL HAFEZ, PH.D., P.ENG.
I, Sabry Abdel Hafez, Ph.D., P.Eng., of Vancouver, British Columbia, do hereby certify: • I am a Senior Mining Engineer with Tetra Tech WEI Inc. with a business address at 800-555 West Hastings Street, Vancouver, British Columbia, V6B 1M1. • This certificate applies to the technical report entitled “Technical Report and Preliminary Assessment of the Zafranal Project, Peru”, dated January 16, 2013 (the “Technical Report”). • I am a graduate of Assiut University, (B.Sc Mining Engineering, 1991; M.Sc. in Mining Engineering, 1996; Ph.D. in Mineral Economics, 2000). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (License #34975). My relevant experience is in mine evaluation. I have more than 19 years of experience in the evaluation of mining projects, advanced financial analysis, and mine planning and optimization. My capabilities range from the conventional mine planning and evaluation to the advanced simulation-based techniques that incorporate both market and geological uncertainties. I have been involved in the technical studies of several base metals, gold, coal, and aggregate mining projects in Canada and abroad. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”). • I did not complete a personal inspection of the Property. • I am responsible for Sections 1.11, 19.0, 22.0, and 28.0 of the Technical Report. • I am independent of AQM Copper Inc. as defined by Section 1.5 of the Instrument. • I have no prior involvement with the Property that is the subject of the Technical Report. • I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument. • As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Signed and dated this 16th day of January, 2013 at Vancouver, British Columbia
Original document signed and sealed by Sabry Abdel Hafez, Ph.D., P.Eng. Sabry Abdel Hafez, Ph.D., P.Eng. Senior Mining Engineer Tetra Tech WEI Inc.
AQM Copper Inc. 28-5 1295840100-REP-R0001-05 Technical Report and Preliminary Assessment of the Zafranal Project, Perú
CARLOS GUZMÁN
I, Carlos Guzmán, of Santiago, Chile, do hereby certify: • I am a Mining Engineer with NCL Ingeniería y Construcción Ltd with a business address at General del Canto 235, Providencia, Santiago, 750- 0588, Chile. • This certificate applies to the technical report entitled “Technical Report and Preliminary Assessment of the Zafranal Project, Peru”, dated January 16, 2013 (the “Technical Report”). • I am a graduate of Universidad de Chile, (Mining Engineer, 1995). I am a Registered Member in good standing of Comisión Calificadora de Competencias en Recursos y Reservas Mineras (Chilean Mining Commission). I am a member of the Australasian Institute of Mining and Metallurgy (AusIMM 229036). I have practiced my profession continuously since 1995. Currently I am Principal and Project Director with the firm NCL Ingeniería y Construcción Ltd, Santiago, Chile. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”). • I conducted a personal inspection of the Zafranal Property in November 29, 2011 for a period of one day. • I am responsible for Sections 1.7, 15.0, 16.0, 21.1.8, 21.2.2, 25.3, 26.3, and 28.0 of the Technical Report. • I am independent of AQM Copper Inc. as defined by Section 1.5 of the Instrument. • I have no prior involvement with the Property that is the subject of the Technical Report. • I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument. • As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Signed and dated this 16th day of January, 2013 at Santiago, Chile
Original document signed and sealed by Carlos Guzmán Carlos Guzmán Mining Engineer NCL Ingeniería y Construcción Ltd
AQM Copper Inc. 28-6 1295840100-REP-R0001-05 Technical Report and Preliminary Assessment of the Zafranal Project, Perú
WILSON MUIR, P.ENG.
I, Wilson Muir, P.Eng., of North Bay, Ontario, do hereby certify: • I am a Senior Engineer with Knight Piésold Ltd. with a business address at 1650 Main Street West, North Bay, Ontario, P1B 8G5. • This certificate applies to the technical report entitled “Technical Report and Preliminary Assessment of the Zafranal Project, Peru”, dated January 16, 2013 (the “Technical Report”). • I am a graduate of the University of British Columbia, (Bachelor of Applied Science, Geological Engineering, 1994). I am a member in good standing of the Association of Professional Engineers and Geoscientists of British Columbia (#24345) and Professional Engineers Ontario (#100060272). My relevant experience includes tailings and water management planning and design. I am a “Qualified Person” for the purposes of National Instrument 43-101 (the “Instrument”). • My most recent personal inspection of the Property was July 6 and 7, 2011 for two days. • I am responsible for Sections 18.4, 18.8, 21.1.10, 21.1.12, 21.2.4, 25.6, 26.11, 26.13, 27.0 (tailings management facility section only), and 28.0 of the Technical Report. • I am independent of AQM Copper Inc. as defined by Section 1.5 of the Instrument. • I have no prior involvement with the Property that is the subject of the Technical Report. • I have read the Instrument and the sections of the Technical Report that I am responsible for has been prepared in compliance with the Instrument. • As of the date of this certificate, to the best of my knowledge, information and belief, the sections of the Technical Report that I am responsible for contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. Signed and dated this 16th day of January, 2013 at North Bay, Ontario
Original document signed and sealed by Wilson Muir, P.Eng. Wilson Muir, P.Eng. Senior Engineer Knight Piésold Ltd.
AQM Copper Inc. 28-7 1295840100-REP-R0001-05 Technical Report and Preliminary Assessment of the Zafranal Project, Perú