Lumina Copper Corp Taca Taca Copper/ Gold Molybdenum Project

2317-RPT-003 Revision Number 0

Preliminary Economic Assessment Report Date of Report: May 24, 2013

Effective Date: April 9, 2013

QUALIFIED PERSONS:

Scott C. Elfen, P. E. Robert Sim, P. Geo. Bruce M. Davis, FAusIMM William L. Rose, P. E. Kevin Scott, P. Eng. (Principal)

Prepared by: Prepared for:

855 Homer Street Vancouver, BC V6B 2W2 Canada

Table of Contents 1 Summary 1 1.1 Introduction 1 1.2 Project Location 2 1.3 Property Ownership 2 1.4 Property Description 4 1.5 Geology and Mineralization 5 1.6 Drilling 5 1.7 Mineral Resource 5 1.8 Mining 6 1.9 Mineral Reserve Estimate 8 1.10 Metallurgical Testing 8 1.11 Process Design and Recovery 8 1.12 Execution Plan and Schedule 9 1.13 Capital Cost 9 1.14 Operating Cost 10 1.15 Marketing Studies 11 1.16 Economic Evaluation 12 1.17 Interpretations and Conclusions 15 1.18 Recommendations 16 1.19 Cautionary Note Regarding Forward-Looking Information and Statements 17 2 Introduction 21 2.1 Purpose of the Technical Report 21 2.2 Sources of Information 21 2.3 Personal Inspection of the Taca Taca Project 22 2.4 Currency Assumptions 22 3 Reliance on Other Experts 23 4 Property Description and Location 24 4.1 Property Location 24 4.2 Property Ownership and Agreements 24 4.3 Environmental Liabilities and Permitting 29 4.4 Other Significant Factors and Risks Affecting Access or Title 29 5 Accessibility, Climate, Local Resources, Infrastructure, & Physiography 30 5.1 Location and Access 30 5.2 Physiography and Vegetation 30 5.3 Climate and Topography 30 5.4 Local Resources 30 5.5 Regional Infrastructure 31 6 History 32 7 Geological Setting and Mineralization 34 7.1 Regional Geology 34 7.2 Local and Property Geology 36 7.3 Mineralization 38 7.4 Structure 44 8 Deposit Types 46 9 Exploration 47

9.1 Historic Exploration Programs (Non-Drilling) 47 9.2 Exploration by Lumina Copper (2010 - 2012) 47 10 Drilling 49 10.1 Drilling 49 10.2 Historic Drilling 50 10.3 Drilling by Lumina Copper (2010-2012) 51 11 Sample Preparation, Analyses, and Security 53 11.1 Historic Drilling and Sample Preparation 53 11.2 Lumina Copper Drilling (2010-2012) 56 12 Data Verification 65 12.1 Historic Drilling 65 12.2 Lumina Copper Drilling 65 12.3 Conclusions 66 13 Mineral Processing and Metallurgical Testing 67 13.1 General 67 13.2 Metallurgical Testing 67 13.3 Comminution Test Work 69 13.4 Flotation Test Work 70 13.5 Sedimentation and Filtration Test Work 82 14 Mineral Resource Estimates 84 14.1 Introduction 84 14.2 Geologic Model, Domains, and Coding 84 14.3 Available Data 86 14.4 Compositing 90 14.5 Exploratory Data Analysis 91 14.6 Bulk Density Data 94 14.7 Evaluation of Outlier Grades 95 14.8 Variography 96 14.9 Model Setup and Limits 97 14.10 Interpolation Parameters 98 14.11 Validation 99 14.12 Resource Classification 105 14.13 Mineral Resources 106 14.14 Comparison with Previous Resource Estimate 111 14.15 Discussion of Factors Materially Affecting Mineral Resources 114 15 Mineral Reserve Estimates 115 16 Mining Methods 116 16.1 Economic Pit Limit Evaluations 116 16.2 Mining Phase/Pit Designs 118 16.3 Mine Production Schedule 120 16.4 Waste Rock Storage and Stockpiling 124 16.5 Mine Equipment 125 16.6 Mine Workforce 128 17 Recovery Methods 129 17.1 Introduction 129 17.2 Process Design Basis and Design Criteria Summary 129 17.3 Flow Sheet Development and Equipment Sizing 130 17.4 Comminution Circuit Design 133

17.5 Bulk Copper-Molybdenum Flotation Circuit Design 134 17.6 Copper-Molybdenum Separation and Molybdenum Cleaning 135 17.7 Copper Concentrate Thickening and Filtration 136 17.8 Molybdenum Concentrate Thickening and Filtration 136 17.9 Concentrate Rail Load-Out Facility 136 17.10 Tailings Processing 136 17.11 Brine, Desalinated, and Process Water 137 17.12 Reagents 137 17.13 Plant Layout & Description 137 18 Project Infrastructure 139 18.2 Roads 139 18.3 Power Supply 144 18.4 Site Electrical Power Distribution 146 18.5 Tailings Storage Facility 146 18.6 Reclaim Water System 149 18.7 Hydrogeology and Raw Water Supply Sources 149 18.8 Process Water Supply 151 18.9 Potable Water 154 18.10 Logistics and Transportation 154 18.11 Concentrate Storage, Handling, and Transport 159 18.12 Ancillary Site Buildings and Facilities 165 18.13 Permanent Camp 165 19 Market Studies and Contracts 167 19.1 Introduction 167 19.2 Supply and Demand 167 19.3 Smelter Capacities and Utilization 167 19.4 Ocean Freight 168 19.5 Future Metals Pricing 168 20 Environmental Studies, Permitting, and Social or Community Impact 169 20.1 Environmental Baseline 169 20.2 Permitting 170 20.3 Conceptual Closure 170 20.4 Socioeconomic Conditions 173 21 Capital and Operating Costs 177 21.1 Capital Cost Estimate 177 21.2 Estimate Exclusions 178 21.3 Basis of Estimate – Mining 178 21.4 Basis of Estimate – Process and Infrastructure Direct Costs 178 21.5 Basis of Estimate – Indirect Costs 179 21.6 Owner’s Costs 179 21.7 Other Costs – Freight, Duties & Taxes 180 21.8 Contingency 180 21.9 Accuracy 181 21.10 Initial Capital 181 21.11 Operating Costs 182 21.12 Basis of Estimate 183 21.13 Mining Operating Costs 183 21.14 Process Operating Costs 184 21.15 Infrastructure Maintenance 185 21.16 Railroad Operation 186

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21.17 General and Administrative 186 21.18 Mine Reclamation 187 21.19 C-1 Cash Costs (net of credits) 187 22 Economic Analysis 189 22.1 General Criteria 189 22.2 Production Summary 190 22.3 Gross Income from Mining 191 22.4 Net Smelter Revenue (NSR) Calculation 192 22.5 Royalty Calculation 193 22.6 Operating Margin 193 22.7 Retention Taxes 193 22.8 Depreciation and Income Tax 193 22.9 Initial Capital Costs 195 22.10 Sustaining Capital Costs 195 22.11 Working Capital 197 22.12 Base Case Analysis 198 22.13 Base Case Sensitivity Analysis 198 22.14 Economic Model 201 23 Adjacent Properties 204 24 Other Relevant Data and Information 205 25 Interpretations and Conclusions 206 25.1 Interpretations and Conclusions 206 25.2 Risks and Opportunities 207 26 Recommendations 210 27 References 212 28 Date and Signature Page 215

List of Figures

Figure 1-1: Property Location Map (Afrainlle, June 2012) ...... 2 Figure 1-2: Taca Taca Property ...... 4 Figure 1-3: IRR Sensitivity Analysis – Metal Prices ...... 14 Figure 1-4: NPV Sensitivity Analysis – Metal Prices ...... 14 Figure 4-1: Property Location Map (Afrainlle, June 2012) ...... 24 Figure 4-2: Claim Map ...... 27 Figure 7-1: Regional Geology (Almandoz, 2008) ...... 35 Figure 7-2: Surface Geology of the Taca Taca Project (Almandoz, 2008) ...... 36 Figure 7-3: Paragenesis of Mineralization at Taca Taca (Almandoz, 2008) ...... 39 Figure 7-4: Extent of Hypogene Sulphide Mineralization (Wells, 2012) ...... 40 Figure 7-5: Vertical East-West Oriented Cross Section 728 3950 N (Wells, 2012) ...... 41 Figure 7-6: Extent of Supergene Enrichment Zone (Wells, 2012) ...... 42 Figure 7-7: Extent of Gold in Leach Cap (Wells, 2012) ...... 43 Figure 7-8: Hematite Copper-Gold Veins (Wells, 2012) ...... 44 Figure 9-1: Titan 24 IP and MT Anomalies (Sim, 2011) ...... 48 Figure 10-1: Drill Collar Plan Map – Taca Taca Project (Wells, 2012) ...... 50 Figure 11-1: ALS Chemex OREAS 50C SRM Copper Control Chart ...... 59 Figure 11-2: ALS Chemex OREAS 50C SRM Gold Control Chart ...... 59 Figure 11-3: ALS Chemex OREAS 50C SRM Molybdenum Control Chart ...... 60 Figure 11-4: ALS Chemex Coarse Blank Copper Control Chart ...... 60 Figure 11-5: ALS Chemex Copper Coarse Duplicate Comparisons ...... 61 Figure 11-6: Alex Stewart OREAS 503 SRM Copper Control Chart ...... 62 Figure 11-7: Alex Stewart OREAS 503 SRM Gold Control Chart ...... 62 Figure 11-8: Alex Stewart OREAS 503 SRM Molybdenum Control Chart ...... 63 Figure 11-9: Alex Stewart Coarse Blank Copper Control Chart ...... 63 Figure 11-10: Alex Stewart Copper Coarse Duplicate Comparisons ...... 64 Figure 13-1: Copper Recovery vs Copper Feed Grade, Roughers – Supergene (Plenge, 2012) ...... 73 Figure 13-2: Molybdenum Recovery vs Molybdenum Feed Grade, Roughers – Supergene (Plenge, 2012) ...... 74 Figure 13-3: Copper Recovery vs Copper Feed Grade, Roughers – Primary (Plenge, 2012)...... 75 Figure 13-4: Molybdenum Recovery vs Molybdenum Feed Grade, Roughers – Primary (Plenge, 2012) ...... 76 Figure 13-5: Copper Recovery vs Copper Feed Grade, Cleaners – Supergene (Plenge, 2012) ...... 77

Figure 13-6: Molybdenum Recovery vs Molybdenum Feed Grade, Cleaners – Supergene (Plenge, 2012) ...... 78 Figure 13-7: Copper Recovery vs Copper Feed Grade, Cleaners – Primary (Plenge, 2012) ...... 79 Figure 13-8: Molybdenum Recovery vs Molybdenum Feed Grade, Cleaners – Primary (Plenge, 2012) ...... 80 Figure 14-1: Isometric Views of MinZone Domains (Sim, 2013) ...... 86 Figure 14-2: Distribution of Copper Grades in Drill Holes (Sim, 2013) ...... 88 Figure 14-3: Distribution of Additional Sample Data since the April 2012 Resource Estimate (Sim, 2013) ...... 89 Figure 14-4: Box Plot Copper by MinZone Domain ...... 92 Figure 14-5: Contact Profiles Copper between LX and SS MinZones ...... 93 Figure 14-6: Contact Profiles Copper between SS and PR MinZones ...... 93 Figure 14-7: Herco Copper in SS Zone ...... 100 Figure 14-8: Herco Copper in PR Zone ...... 101 Figure 14-9: GT Comparison of OK, IDW, and NN Copper Models ...... 102 Figure 14-10: GT Comparison of OK, IDW, and NN Molybdenum Models ...... 102 Figure 14-11: GT Comparison of OK, IDW, and NN Gold Models ...... 103 Figure 14-12: Copper Model Swath Plot by Easting ...... 104 Figure 14-13: Copper Model Swath Plot by Northing ...... 104 Figure 14-14: Copper Model Swath Plot by Elevation ...... 105 Figure 14-15: Plan Map Showing Limit of Porphyry Type Mineralization (Sim, 2013) ...... 106 Figure 14-16: Extents of Base Case Indicated and Inferred Copper Resources (Sim, 2013) ...... 109 Figure 14-17: Extents of Base Case Resources Including LX Zone Gold Resource (Sim, 2013) ...... 109 Figure 14-18: Changes to Base Case Indicated Resources Nov 2012 vs. April 2012 (Sim, 2013) ...... 113 Figure 14-19: Changes to Base Case Indicated & Inferred Resources Nov 2012 vs. April 2012 (Sim, 2013) ...... 113 Figure 16-1: Ultimate Pit Design (WLRC, 2013) ...... 120 Figure 16-2: Waste Rock Storage and Oxide Stockpile Facilities (WLRC, 2013) ...... 125 Figure 17-1: Overall Process Flow Diagram ...... 131 Figure 17-2: Pebble Crushing – Mill Building Plan and Elevations ...... 138 Figure 18-1: General Mine Access Road (Ausenco, 2013) ...... 140 Figure 18-2: Mine Access Road (Ausenco, 2013) ...... 141 Figure 18-3: Out of Pit Haul Roads (Ausenco, 2013) ...... 143 Figure 18-4: Transmission Line Routing from Olacapato to Taca Taca (Hugo Gil, 2013) ...... 145 Figure 18-5: Tailings Discharge System – Jacking Headers ...... 148 Figure 18-6: Exit from the Project along Ruta 27 and Access to Ruta 51 to Paso Sico (TUSA, 2011) 157

Figure 18-7: Travel from Paso Sico to Antofagasta through Lacos, Socaine, and Peine (TUSA, 2011) ...... 158 Figure 18-8: Travel from Paso Sico, to San Pedro de Atacama and Calama, arriving to Antofagasta (TUSA, 2011) ...... 158 Figure 18-9: Map of Railway Routes between Taca Taca and Mejillones (TFP, 2013) ...... 160 Figure 18-10: Site Layout ...... 161 Figure 21-1: Percentage of Processing Costs by Component ...... 185 Figure 21-2: C-1 Cash Costs Net of Credits ...... 188 Figure 22-1: IRR Sensitivity Analysis – Metal Prices ...... 199 Figure 22-2: NPV Sensitivity Analysis – Metal Prices ...... 199 Figure 22-3: IRR Sensitivity Analysis – Capital v Operating Costs ...... 200 Figure 22-4: NPV Sensitivity Analysis – Capital v Operating Costs ...... 200 Figure 22-5: IRR Sensitivity Analysis – Metallurgical Recovery ...... 200 Figure 22-6: NPV Sensitivity Analysis – Metallurgical Recovery ...... 201 Figure 22-7: Cash Flow Model Forecast ...... 202

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List of Tables

Table 1-1: Mineral Resource Summary ...... 6 Table 1-2: Capital Cost Summary ...... 9 Table 1-3: Operating Cost Summary ...... 10 Table 1-4: Metal Price Forecast ...... 12 Table 1-5: Project Metals Pricing ...... 12 Table 1-6: Economic Model Inputs ...... 12 Table 1-7: Net Present Value with Mid-Period Adjustment ...... 14 Table 1-8: Estimated Cost for Recommended Future Work ...... 16 Table 4-1: Mining Concessions ...... 25 Table 6-1: Exploration and Ownership History of the Taca Taca Project ...... 32 Table 10-1: Drilling History ...... 49 Table 13-1: Summary of Comminution Test Results ...... 69 Table 13-2: Summary of SAGDesign® Results ...... 69 Table 13-3: JKTech – SMC Test Results ...... 70 Table 13-4: Summary of Locked Cycle Tests Using Tap Water ...... 70 Table 13-5: Locked Cycle Tests on Composites Comparing Use of Brine Water and Tap Water in Roughers ...... 71 Table 13-6: Variability Rougher Test – Supergene ...... 72 Table 13-7: Variability Rougher Tests – Primary ...... 74 Table 13-8: Variability Cleaner Tests – Supergene ...... 76 Table 13-9: Variability Cleaner Tests – Primary ...... 78 Table 13-10: Batch Cleaner Test Results at Different Water Quality Blends ...... 80 Table 13-11: Summary of Brine Water Analyses ...... 81 Table 13-12: Copper Concentrate Sedimentation Results ...... 82 Table 13-13: Tailings Sedimentation Results ...... 82 Table 13-14: Filtration Results for Locked Cycle Test Copper Concentrates ...... 82 Table 14-1: MinZone Domains and Coding ...... 85 Table 14-2: Basic Summary of All Sample Data ...... 90 Table 14-3: Basic Summary of Sample Data Proximal to the Resource Model ...... 90 Table 14-4: Summary of Interpolation Domains ...... 94 Table 14-5: Summary of Bulk Density by MinZone Domain ...... 95 Table 14-6: Summary of Outlier Limits ...... 95 Table 14-7: Variogram Parameters - Copper ...... 96

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Table 14-8: Variogram Parameters - Molybdenum ...... 97 Table 14-9: Variogram Parameters - Gold ...... 97 Table 14-10: Block Model Limits ...... 98 Table 14-11: Interpolation Parameters ...... 99 Table 14-12: Supergene and Primary Sulphide Zone Mineral Resource Estimate ...... 107 Table 14-13: Leach Zone Oxide Mineral Resource Estimate ...... 108 Table 14-14: Copper Mineral Resource Estimate by Type ...... 110 Table 14-15: Comparison of Copper Mineral Resources – April 2012 vs. Nov 2012 ...... 111 Table 14-16: Comparison of Leach Zone Mineral Resources – April 2012 vs. Nov 2012 ...... 112 Table 16-1: Floating Cone Recovery and Economic Parameters ...... 116 Table 16-2: Basic Pit Design Parameters ...... 119 Table 16-3: Pit Design Recommendations ...... 119 Table 16-4: Production Scheduling Parameters ...... 121 Table 16-5: Geologic Zones and Bulk Densities ...... 122 Table 16-6: Taca Taca Mine Production Schedule ...... 123 Table 16-7: Resource Classification of Concentrator Feed – PEA Schedule ...... 124 Table 16-8: Mine Equipment Fleet for Selected Years ...... 127 Table 16-9: Projected Mine Workforce for Selected Years ...... 128 Table 17-1: Summary of Key Process Design Criteria ...... 129 Table 17-2: Mill Design Criteria ...... 133 Table 17-3: Summary of Reagent Consumption ...... 137 Table 18-1: Components of 600 l/s of Untreated Brine Water Make-Up ...... 152 Table 18-2: Components of 200 l/s of Fresh Water Make-Up...... 152 Table 18-3: Estimated Permeate and Concentrate Chemistry ...... 153 Table 18-4: Camp Arrangement ...... 166 Table 19-1: Metal Price Forecast ...... 168 Table 19-2: Project Metals Pricing ...... 168 Table 21-1: Mine Capital Cost Summary ...... 177 Table 21-2: Labour Hourly Rates ...... 179 Table 21-3: Brownfield Productivity Factors ...... 179 Table 21-4: Freight Applied to Material/Equipment Cost ...... 180 Table 21-5: Summary of Initial Capital Costs ...... 181 Table 21-6: LOM Operating Cost Summary ...... 182 Table 21-7: Main Operating Cost Assumptions ...... 183

Table 21-8: LOM Operating Cost Summary - Mining ...... 184 Table 21-9: LOM Operating Cost Summary - Process ...... 184 Table 21-10: Derivation of Plant Operating Estimate ...... 185 Table 21-11: LOM Operating Cost Summary – Infrastructure Maintenance ...... 186 Table 21-12: LOM Operating Cost Summary – Railroad Operation ...... 186 Table 21-13: LOM Operating Cost Summary – G&A ...... 187 Table 21-14: Average LOM C-1 Cash Costs ...... 188 Table 22-1: Economic Model Inputs ...... 189 Table 22-2: Process Production Summary ...... 190 Table 22-3: Calculation of Gross Revenue ...... 192 Table 22-4: TCs, RCs, and Freight ...... 192 Table 22-5: Deduction of Capitalized Costs to Date ...... 194 Table 22-6: MIL Section 13 Depreciation ...... 194 Table 22-7: Sustaining Capital Cost Summary ...... 196 Table 22-8: Sustaining Capital for Concentrator Expansion ...... 197 Table 22-9: Net Present Value with Mid-Period Adjustment ...... 198 Table 22-10: Sensitivity Analysis of IRR and NPV ...... 198 Table 26-1: Estimated Cost for Recommended Future Work ...... 210

Nomenclature and Abbreviations

Abbrev. Description Abbrev. Description A Ampere MCC Motor control centre Bt Billions tonnes MIBC Methyl isobutyl carbinol C Celsius, as in degrees, °C min Minute cm Centimetre mg Milligrams CSS Closed side setting mg/l Milligrams per litre Cu Copper mm Millimetre d Day Mt Million tonnes dmt Dry metric tonne Mt/y Million tonnes per year Engineering, procurement and EPCM MPa Megapascal construction management Fe Iron MVA Megavolt-ampere FEL Front-end loader MW Megawatt g Gram NQ Drill core size (about 47.5 mm) g/l Grams per litre PLC Programmable logic controller g/mol Mole mass, in grams ppb Parts per billion g/t Grams per tonne ppm Parts per million (equivalent to g/t) h Hour PAX Potassium amyl xanthate ha Hectare P80 Size at which 80% (mass) is finer HP Horsepower ROM Run-of-mine HQ Drill core size (about 63.5 mm) s Second k Kilo or thousand SAG Semi-autogenous grinding Supervisory, Control and Data kg Kilogram SCADA Acquisition kg/m3 Kilogram per cubic metre SG Specific gravity km Kilometre t Tonne km2 Square kilometre t/d Tonnes per day kPa Kilopascal t/h Tonnes per hour kV Kilovolt t/y Tonnes per year kVA Kilovolt-ampere TSF Tailings storage facility kW Kilowatt (power) V Volt kWh Kilowatt-hour WRSF Waste rock storage facility l Litre wmt Wet metric tonne m Metre y Year M Mega or million µm Micrometre or micron m2 Square metre  Approximately ° Degrees

1 Summary

1.1 Introduction

The Taca Taca Project (the Project) is a porphyry copper-gold-molybdenum deposit located in northwestern Argentina in the Puna (Altiplano) region of Salta Province. The Project is beneficially owned by Lumina Copper Corp. (Lumina Copper), based in Vancouver, British Columbia, Canada. It is approximately 230 km west of the city of Salta and 55 km east of the Chilean border (Figure 1- 1).

The Project envisages a mine operating over a 28-year life delivering 120,000 t/d of throughput (for an initial period of seven years) to a concentrator comprising two milling and flotation lines. An expansion of the concentrator through the addition of a third line is contemplated to increase the total throughput to 180,000 t/d from Year 8 onwards.

While the Project is remote, there is good regional infrastructure. A network of paved and gravel roads from Salta to the towns of San Antonio de los Cobres and Tolar Grande provide access to the Project. The road continues west beyond the Project to the Socompa Pass on the Chilean border and eventually to the port city of Antofagasta, Chile.

The Project is located within 10 km of the railway line that connects Salta with Antofagasta. This rail infrastructure will be used to transport concentrates and select consumables to and from the Port of Mejillones, 65 km north of Antofagasta and other consumables from within Argentina.

Electrical power connection to the national power grid is available in the region at Olacapato to the north of the Project and a 345 kV transmission line over a length of 144 km is proposed to provide electrical power to the Project.

Although the region is arid, subsurface water is available in the local area. A water supply study and water balance analysis were completed for the Project and based on these studies, the processing water flow sheet shows that there are adequate sources of water available for the Project. Additional prospective areas have recently been identified near the Project that may be sources of additional fresh water, which could further optimize the process water flow sheet.

Figure 1-1: Property Location Map (Afrainlle, June 2012)

1.2 Project Location

The Project is located in the Puna (Altiplano) region of Salta Province, northwest Argentina at an elevation of 3,585 meters above sea level (masl). It is approximately 230 km west of the city of Salta. The centre of the property is at latitude 24.7oS and longitude 68.0oW. The UTM coordinates are 7283500 N and 2628000 E (geographic projection: Gauss-Kruger POSGAR 94/Argentina WGS84, Zone 2).

1.3 Property Ownership

The Project consists of the Grupo Minero Taca Taca concession (Grupo Minero) that covers 2,559.96 ha, and 27 additional mining (mina) and exploration (cateo) concessions and one land use application, owned wholly or in part, that covers 58,001.32 ha. In the region, one land use application, one additional mining concession application, and two additional exploration concession applications have been registered, but not yet granted; these total 3,291.95 ha. The mining properties and rights comprising the Project, including the applicable file numbers, area,

status, and contractual royalty encumbrances are listed in and shown in Table 4-1 and are shown in Figure 4-2 as part of a detailed discussion in Item 4.0.

The Grupo Minero concessions are maintained through annual fees (canon) which at current official exchange rates are equivalent to $4,410 per year. The remaining concessions have an annual canon fee of $23,360. One-half of the total annual canon fees ($13,885) are paid semi-annually in June and December.

As of May 2, 2013, all of the mining and exploration concessions were in good standing and canon fees have been paid through June 30, 2013. The Grupo Minero and other concessions are valid for an unlimited period of time as long as the semi-annual canon payments are made. All exploration concessions will have to be converted to mine properties in 2013 and 2014. The Grupo Minero and other mining concessions include the right to exploit, subject to being granted by an environmental permit for exploitation. The exploration concessions include the right to explore for all metals or minerals.

The surface lands covering the Project are owned by the Province of Salta and the necessary access permits were granted for the current drilling work. All known mineralized zones are located hundreds of metres within the limits of the Project boundaries.

The Grupo Minero and other mining and exploration concessions are registered under the name Corriente Argentina S.A. (CASA) except for Mina Don Francisco which is held in trust for CASA by S. Arbeleche and Mina Francisco 1 and Mina Francisco 2 which are jointly owned through a company 50% owned by CASA and 50% owned by Salta Exploraciones S.A. (SESA). Lumina Copper is the beneficial owner of all of the issued and outstanding shares of CASA. A member of Lumina Copper management holds approximately 1.0% of the issued and outstanding shares in the capital of CASA in trust for the benefit of Lumina Copper in order to address certain requirements applicable to Argentinian companies under Argentine corporate law.

Lumina Copper first acquired an interest in the Taca Taca property when shareholders of Global Copper Corp. (Global Copper) approved a corporate reorganization effective August 1, 2008 by way of a statutory plan of arrangement (Global Arrangement); pursuant to the Global Arrangement, Teck Resources acquired all Global Copper’s shares. Global Copper’s assets, excluding the Relincho Project in Chile, were transferred to Lumina Copper; this included ownership of Minera Corriente Chile S.A. (Minera Corriente) which at the time indirectly held a 100% interest in the Taca Taca property as it was then structured. Effective August 19, 2012, Minera Corriente was wound up into Lumina Copper, leaving Lumina Copper the beneficial owner of all of the issued and outstanding shares of CASA.

Since completion of the Global Arrangement, CASA has subsequently acquired additional mineral concessions through a combination of purchases from third party owners, administrative lotteries, and staking. The present property position is shown in Figure 4-2 and concessions are listed in Table 4-1.

Some of the mining concessions that form the Grupo Minero and one ancillary mining concession are subject to a contractual royalty of 1.5% of net smelter returns (NSR) (Taca Taca Royalty). Franco Nevada Corp., through a wholly-owned subsidiary, holds the right to receive a 72% interest in the Taca Taca Royalty, and the remaining 28% interest is held by two individuals. A number of additional mining concessions are subject to contractual royalties of up to 1.5% of NSR. Table 4-1 includes a summary of the contractual royalties that apply with respect to the mining properties and rights comprising the Taca Taca property.

A royalty of up to 3% net of smelting/refining, transportation, administrative, and plant processing costs (also known as the “mine mouth” value) is payable to the Province of Salta.

1.4 Property Description

The Project is located in a remote area of Argentina, as can be seen in Figure 1-1 and Figure 1-2. A network of paved and gravel roads from Salta to the towns of San Antonio de los Cobres and Tolar Grande provide access to the Project. The property is at an approximate elevation of 3,585 masl.

The Project is located within 10 km of the railway line that connects Salta with Antofagasta, Chile. The rail infrastructure is contemplated to be used to transport concentrates and select consumables to and from the Port of Mejillones, 65 km north of Antofagasta and other consumables from within Argentina.

Figure 1-2: Taca Taca Property

1.5 Geology and Mineralization

The Taca Taca porphyry copper-gold-molybdenum deposit is hosted in the southern half of a 50 km long Ordovician batholith, which forms the Sierra de Taca Taca mountain range. The batholith consists of coarse-grained granite that is cut by several aplite dykes. This Early Paleozoic intrusion is intruded by Late Permian granites and aplites and overlain by Late Permian sediments and volcaniclastics. Narrow, north-south striking, steeply dipping rhyolitic dykes of Permo-Triassic age outcrop throughout the region. Oligocene rhyodacitic intrusions of the Santa Inés Formation are responsible for the porphyry copper mineralization and alteration at Taca Taca.

Late Tertiary red-bed sedimentary rocks are widely distributed in the region, but are most abundant east of Salar de Arizaro. These rocks possibly constitute the basal section of the sedimentary sequence that fills the salar basin. Lavas from recent (Pliocene to Pleistocene) volcanoes are exposed to the west and north of Taca Taca. Large evaporite deposits of alternating salts and sand were deposited in regional intermontane basins to form the present-day salars (Almandoz, 2008).

There are three main styles of mineralization associated with the Taca Taca copper-gold- molybdenum porphyry: supergene/hypogene porphyry copper mineralization, remnant oxide copper-gold mineralization in the leach cap, and hematite-quartz copper-gold veins. Each style of mineralization is discussed in more detail within Item 7.0.

1.6 Drilling

From 1975 through 2012, six different companies have completed seven drilling campaigns at the Project (Table 10-1). The drill collar locations are illustrated in Figure 10-1.

A total of 440 holes (163,537 m) have been drilled on the Taca Taca property. From 2010 to 2012, Lumina Copper drilled 273 holes (134,033 m). The mineral resource model is based on 310 drill holes (147,449 m) which tested the extent of the known porphyry mineralization.

1.7 Mineral Resource

The most recent mineral resource estimate for the Project (effective date of October 30, 2012) completed by SIM Geological Inc. (SIM Geological) and BD Resource Consulting, Inc. (BDRC), and the corresponding block model, has been used to develop the mine plan and production schedule for the PEA.

The majority of drill holes in the main deposit area are vertically oriented with holes spaced on a nominal 150 m grid pattern. At the northern end of the deposit, the final 500 m has been tested with holes that are consistently inclined -70° east.

The resource estimate was generated using drill hole sample assay results and an interpretation of the geologic model which relates to the spatial distribution of copper, gold, and molybdenum. Interpolation characteristics were defined based on the geology, drill hole spacing, and geostatistical analysis of the data. The resources were classified by their proximity to the sample locations and are reported, as required by NI 43-101, according to the CIM standards on Mineral Resources and Mineral Reserves (November 2010).

The Project's current mineral resource estimate (at a 0.3% copper equivalent cut-off grade) is shown in the table below:

Table 1-1: Mineral Resource Summary

Size(1) Grade Contained Metal

Cu Eq(3) Cu Au Mo Cu Au Mo Tonnes (Million) (%) (%) (g/t) (%) (B lb) (M oz) (M lb)

Indicated Resources

2,165 0.57 0.44 0.08 0.013 21.15 5.56 615.8

Inferred Resources (2)

921 0.47 0.37 0.05 0.012 7.55 1.57 235.4 Notes: (1) Mineral resources have been estimated as at October 30, 2012. Totals may not add up due to rounding. (2) Inferred mineral resources have a great amount of uncertainty as to their existence and as to whether they can be mined legally or economically. It cannot be assumed that all or any part of inferred mineral resources will ever be upgraded to a higher category. (3) Copper equivalent (CuEq) calculated using $2.00/lb Cu, $800/oz Au, and $12.00/lb Mo and is not adjusted for mining and metallurgical recoveries as these remain uncertain. The formula used is as follows: CuEq = Cu% + (Au g/t x 0.583) + (Mo% x 6).

The Project's quality assurance/quality control (QA/QC) program was monitored by independent consultant Dr. Bruce M. Davis, FAusIMM of BDRC. Logging and sampling was completed at Lumina Copper’s secure facility located at the Project. Drill core was mechanically split on site before being sent to either ALS Chemex's or Alex Stewart Argentina's preparation facilities in Mendoza, Argentina.

Lumina Copper established a QA/QC protocol that comprised the use of reject duplicates, standards, and blanks inserted into the sample batches at regular intervals. Fine material duplicates were inserted during sample preparation (independent from the assay laboratory) by splitting of the pulps. A range of copper-gold-molybdenum standard reference materials (SRMs) of suitable matrix composition, together with blanks, was inserted by Lumina Copper during the core sampling procedure. The structure of this QA/QC program follows accepted industry standards. Irregular or suspect results were addressed in a timely manner in order to ensure the integrity of the database. The results of this QA/QC program indicate that the Project's database of sampling, analytical, and test data is of sufficient accuracy and precision to be used for the generation of mineral resource estimates.

The oxide gold mineral resource estimate defined within the Project's leached cap was treated as waste in the production schedule. Metallurgical test work completed to date on the oxide gold resource has been limited. Preliminary test results indicate, however, that future studies should be considered and performed in sufficient detail to support an economic analysis.

1.8 Mining

The Taca Taca deposit is amenable to conventional, large-scale, open pit mining methods. Floating cone (FC) evaluations were conducted to determine potentially economic pit limits and the mining phase (pushback) development sequence. Six mining phases were designed and used to estimate contained mineral resources, from which a mine production schedule was developed. This schedule was based on two grinding lines in the concentrator processing a total of 120,000 t/d for the first seven years of operation with the addition of a third grinding line in Year 8 increasing the daily mineralized material processing rate to 180,000 t/d.

Wyllie & Norrish Rock Engineers Inc. (W&N), with sub consultants Fisher & Strickler Rock Engineering, LLC, (FSR) performed geotechnical analyses in support of scoping-level pit slope recommendations for the Taca Taca Project. These analyses were based primarily upon geotechnical data collected by W&N and Lumina Copper personnel during the course of exploration core drilling and dedicated geotechnical drilling in the years 2010 to 2012.

Six slope sectors were defined to control the floating cone projections and subsequent pushback designs. Overall slope angles for floating cones ranged from about 38° in the north walls, 42° in the east and southeast walls, to 44-46° in the southern walls, and about 42° in the northwestern quadrant of the pit.

Preproduction stripping is estimated at 251 Mt during a three-year development period. Sulfide milling operations are projected to last about 28 years.

Indicated sulphide resources of 1,545 Mt and grading 0.46% Cu and inferred sulphide resources of 106 Mt and grading 0.43% Cu, at an average stripping ratio of 1.57, were estimated in this production schedule using a declining cut-off grade strategy that increases the potential present value.

This 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 results of this PEA will be realized. Mineral resources that are not mineral reserves have no demonstrated economic viability.

Workforce and equipment requirements were estimated on the basis of working two 12-hour shifts per day, 7 d/wk, 360 d/y (allotting five days for weather and holiday-related interruptions). A four- rotating-crew work schedule will be required for continuous operator and maintenance coverage. A peak mine workforce of 874 is projected in Year 7.

Large-scale equipment was selected for maximum economies of scale. Taca Taca mining would be conducted from 15 m high benches using conventional shovel-truck practices. Electric drills and shovels were chosen because of their lower operating costs and better reliability. The rest of the mining fleet would consist of diesel-powered machinery that provides flexibility in pit operations.

The selected primary mining fleet includes the following equipment:

 Rotary blasthole drills capable of 270 to 311 mm-diameter holes  60 m3 rope shovels  42 m3 hydraulic shovel  40 m3 front-end loader (FEL)  363 t off-highway haul trucks  635 kW and 435 kW crawler dozers  600 kW rubber-tired dozer  370 kW rubber-tired dozers  400 kW and 220 kW motor graders  120,000 l (134 t) water trucks

1.9 Mineral Reserve Estimate

This item is not applicable to this PEA.

1.10 Metallurgical Testing

Laboratorio Plenge (Plenge) of Lima, Peru conducted a series of test programs on Taca Taca samples over a period of three years starting in April 2010. The final test program was completed in October 2012. Programs consisted of:

 Scoping level comminution and flotation tests;  Variability and optimization flotation tests, including copper/molybdenum separation; and,  Settling and filtration tests.

Plenge’s final test program was completed on representative core samples selected from across the deposit on the basis of spatial, physical, mineralogical, and chemical characteristics. For the majority of the test work that forms the basis of the PEA, core was divided into four composite samples, two for each major alteration type, supergene and primary.

The two composites for each alteration type represented two different periods in the anticipated mine production schedule, Years 1-5 and Years 6-10.

Observations are summarized below:

 Ausenco reviewed the test work and concluded it is reasonably extensive and suitable for this level of study. The comminution data are considered adequate for a conceptual milling circuit design. The Taca Taca mineralized material is of moderate competency and hardness, and amenable to grinding in a conventional SAG-ball milling circuit with pebble crushing (SABC) and re-grinding.  The location of the samples and drill holes are reasonably representative of the deposit for a PEA given the nature and continuity of rock types and mineralization of the deposit.  Recoveries over the life of mine (LOM) are expected to range from 88% to 92% for copper, 56% to 57% for molybdenum, and 61% to 65% for gold, which will be contained in the copper concentrate. LOM metal recoveries are expected to average approximately 90% for copper, 64% for gold, and 57% for molybdenum.

1.11 Process Design and Recovery

The Taca Taca concentrator is planned to process 120,000 t/d of run of mine (ROM) material. Copper and molybdenum concentrates and tailings will be produced. The proposed process includes crushing and grinding of the ROM material, bulk copper-molybdenum rougher and cleaner flotation, regrinding, copper-molybdenum separation, molybdenum flotation, and dewatering of copper and molybdenum concentrates. The flotation tailings will be thickened before placement in a tailings storage facility (TSF). Prior to Year 8 an expansion is planned to be undertaken to increase the capacity of the concentrator to 180,000 t/d with the addition of a third milling and bulk rougher flotation line.

The copper and molybdenum concentrates are planned to be transported by truck from the concentrator to the rail load-out facility where they will be transported by rail across the Argentinian- Chilean border to the Port of Mejillones, Chile for shipment to international customers.

1.12 Execution Plan and Schedule

An execution schedule has been developed to advance the Project if Feasibility Study results are positive and a development decision is made. The overall preproduction period of 48 months is driven by the need to construct the site infrastructure primarily required to support preproduction stripping. Two years have been estimated for construction of the processing facilities and related infrastructure. The preproduction period ends with the commissioning of the first grinding line (one SAG mill and two ball mills) and its associated flotation circuit. Commissioning of the second grinding line would follow three months later, the staggered start allowing for normal production ramp-up activities. A third line would be constructed in the sixth and seventh years of operation and commissioned in the eighth year, bringing total production to 180,000 t/d.

1.13 Capital Cost

Capital cost estimates for the Technical Report have been completed. Approximately 251 Mt of preproduction stripping, including stockpiling, will be required prior to mining operations commencing. It has been estimated that it will take approximately 2.5 years to complete preproduction stripping at a capitalized cost of $416.1 million. The cost of preproduction stripping, including removal of the leach cap, has been included in the initial capital cost estimate. The total initial capital cost including the mine, concentrator, and infrastructure is estimated to be approximately $3.0 billion including a total of 15% contingency based on total direct and indirect costs. The following table summarizes the capital cost estimate:

Table 1-2: Capital Cost Summary

Description Estimate ($US M) Mine $581.5 Preproduction stripping $416.1 Plant and Processing $685.5 Infrastructure $215.3 Tailings Storage and Waste Rock Storage $139.5 Other $581.0 Contingency $386.5 Total Initial Capital Costs $3,005.5 Start-up Working Capital $54.9 Capital costs for 3rd mill and flotation line and $431.2 associated ancillary costs LOM Sustaining Capital Costs $1,375.5

Note: Figures above may not add up due to rounding.

The capital cost estimates have been compiled with an accuracy level of -25% to +35%.

The initial capital cost estimate for this Technical Report compares favourably to the initial capital costs cited for comparable copper projects under development. The capital intensity ratio on initial capital (initial capital divided by average annual copper production) is a measure of the amount of investment in initial capital infrastructure required to produce a tonne of copper at a project. The capital intensity ratio on initial capital is estimated to be $11,090/t of copper production (for production prior to the expansion in Year 8). The term "capital intensity ratio" does not have a standard meaning and may not be directly comparable to capital intensity ratios presented by other issuers.

The Qualified Person for this section has reviewed and approved the capital cost estimates for inclusion in this Technical Report.

Capital costs are discussed in more detail in Item 21.0.

This 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. These is no certainty that the results of this PEA will be realized. Mineral resources that are not mineral reserves have no demonstrated economic viability.

1.14 Operating Cost

The PEA estimates that the C-1 cash costs (net of by-product credits) over the LOM will average $1.11/lb copper sold. C-1 cash costs include at-mine cash operating costs, treatment and refining charges, royalties, mine reclamation and closure costs, and copper and molybdenum concentrate transportation, and freight costs.

The following LOM operating costs have been forecast for the Project:

Table 1-3: Operating Cost Summary

Site Operating Costs US $/t mill feed Mining $4.67 Processing $4.26 Infrastructure Maintenance $0.06 Railroad Operations $0.19 General & Administration $0.57 Mine Closure Costs $0.02 Total Site Operating Costs $9.77 Other Key Costs $70/dry metric tonnes (dmt) Cu Copper Concentrate Treatment Charges concentrate Copper Refining Charges $0.07/lb Cu $79.67/wet metric tonnes (wmt) Ocean Freight, Port Handling & Other Costs Cu concentrate

Methodologies and further details are presented in Item 21.0.

1.15 Marketing Studies

H&H Metals Corp. (H&H) of New York completed a marketing study for the Project. Using contacts with smelters and recent data from other projects, H&H completed the following marketing aspects in its report: supply and demand for both copper and molybdenum concentrates, including anticipated production from new projects in the construction or planning stages; smelter capacities, utilization, and the impacts on treatment charges (TCs), refining charges (RCs), and payfors; ocean freight costs; and future metals pricing. Details are provided below.

Supply and Demand

While supply and demand dictates pricing for many items, it is only one consideration for metals such as copper and molybdenum. Speculation by investors complicates market analysis. Less than transparent data by China regarding production and inventories further complicates analysis of supply and demand, as does the impact of worldwide economic conditions.

The trend to more affluent populations on the whole, as well as sustained population growth in the populous countries of China and India, support steady growth in demand for copper, in particular, for housing and other infrastructure.

In the short term, existing inventories are expected to be drawn down as new projects and expansions of existing operations are planned for completion. Both of these factors are expected to place downward pressure on copper pricing until later in the decade. At that point, the steadily increasing demand mentioned above is expected to overtake and surpass the increases in supply.

Uncertainty in the supply and demand scenario summarized above is introduced by: new or expanded production facilities coming online later than originally planned; cancellation of new or expanded production facilities due to market, general economic, or company financial conditions; or management of the market sector by the government in a large consumer/producer such as China.

Smelter Capacities and Utilization

For the most part, smelter capacity is fixed. The relationship between capacity and utilization dictates a smelter’s profitability, hence its setting of TCs, RCs, and payfors. There is significant variability of utilization among smelters across the industry, hence considerable variability of TCs, RCs, and payfors. In order to obtain more concentrates, some smelters are agreeing to lower charges.

In the long term, TCs and RCs are expected to increase for smelters to better cover their costs. However, it is worth noting that Lumina Copper may expect to receive premium terms for its clean concentrate. According to H&H, a number of smelters need clean concentrates to blend with their substantial supplies and inventories of dirty concentrates.

H&H’s forecasts for TCs, RCs, and payfors were utilized in the Project cash flow, as detailed in Table 22-4.

Ocean Freight

During the minerals boom period, shipping companies aggressively built vessels to transport the mineral products from producers to markets. Currently, the availability of vessels significantly exceeds the need. Many vessels remain idle with financial building costs exceeding their current value.

Opportunities for reasonable freight costs are available, particularly with negotiation of long-term freight contracts. H&H’s forecast for freight cost of $53 wmt for shipment to Asia has been incorporated in the Project cash flow.

Future Metals Pricing

After analysis and discussion, H&H forecast the following metal prices for the Project shown below in Table 1-4.

Table 1-4: Metal Price Forecast

Product Range Recommend Price

Copper $2.90 - $4.10 $3.10

Molybdenum $11.00 - $15.00 $13.00

Gold $1,350 - $1,900 $1,650

Silver $26.00 - $38.00 $35.00

In addition to the H&H forecast, consideration was given to a $2.89/lb consensus long-term price forecast made by 21 financial institutions (Thomson One Analytics, March 2013). The PEA economic analysis utilizes the following more conservative pricing forecast (Table 1-5), which has been applied throughout the projected 28 year mine life of the Project.

Table 1-5: Project Metals Pricing

Product Price

Copper $2.75/lb

Molybdenum $12.00/lb

Gold $1,200/Toz

1.16 Economic Evaluation

Inputs for the economic model are listed below for easy reference, again as Table 22-1, and discussed in detail in Item 22.0.

Table 1-6: Economic Model Inputs

Description Values

Construction Period 30 months Preproduction Period 4 years LOM after Preproduction 28 years LOM Sulfide Mill Feed (Kt) 1,650,792 LOM Copper Concentrate (DMT) 21,615,739 LOM Molybdenum Concentrate (DMT) 227,884

Description Values

Market Price Copper Price (LOM $/lb Cu) 2.75 Gold Price (LOM $/oz Au) 1,200.00 Molybdenum Price (LOM $/lb Mo) 12.00 Cost and Tax Criteria Estimate Basis 1st Qtr 2013 USD Inflation/Currency Fluctuation None Leverage 100% Equity Income Tax - Argentina 35% Argentine Retention Tax 10% Less: Puna Investment Credit -2.5% Net Retention Tax 7.5% (on Gross Revenue less: TCs, RCs, port handling, ocean freight, and

75% of rail use fees and railroad operation costs to be incurred in Chile)

Depreciation - Argentina Section 13 Infrastructure Construction & Equipment 60/20/20 Machinery, Equipment, and Vehicles 1/3 for 3 years IVA - Value Added Tax (VAT) Payment/Recovery Excluded Royalties Salta Provincial Minemouth Royalty 3.0% (on Gross Revenue less: TCs, RCs, freight, processing, infrastructure maintenance, railroad operation, and general & administration costs) Third Party Royalty on Modified NSR 1.5% (on Gross Revenue less: TCs, RCs, freight, railroad operation, Provincial Minemouth Royalty, and Net Retention Tax paid) Concentrate (Con) Transportation Charges Cu Con: Rail Use Fees, Port Handling, and Ocean Shipping ($/WMT) 79.67 Mo Con: Rail Use Fees and Port Handling ($/WMT) 15.67 Assuming containerization of Mo con product and considered delivered

at port. Payment Terms Cu Con: Cash Against Documents (CAD) one week after shipping 90% Balance received eight weeks after shipping 10% Mo Con: Cash Against Documents (CAD) week of arrival at port 90% Balance received four weeks after delivery to port 10%

The cash flow forecast by the PEA estimates payback of initial capital investment to occur approximately 3.8 years after start of production. The Project is estimated to have an after tax internal rate of return (IRR) of 17.2%.

The following table shows the estimated after tax net present value (NPV) for the Project’s cash flows at various discount rates.

Table 1-7: Net Present Value with Mid-Period Adjustment

Discount Rate 4% 6% 8% 10% 12%

NPV ($M) $4,610.1 $3,121.5 $2,087.3 $1,352.6 $820.5

The following figures show the sensitivity of the estimated after tax IRR and NPV (8% discount rate) to changes in metal prices.

Figure 1-3: IRR Sensitivity Analysis – Metal Prices

Figure 1-4: NPV Sensitivity Analysis – Metal Prices

1.17 Interpretations and Conclusions

Interpretations and conclusions of the Qualified Persons of the Taca Taca PEA are listed below:

 The results of the PEA indicate that the Project is a robust project at this stage of development demonstrating favourable economic potential that warrants further work toward the next stage of development, prefeasibility. The exploration program continues to demonstrate the potential for future growth of the resource.  The sample preparation, security, and procedures followed by Lumina Copper are adequate to support a mineral resource estimate.  Assay data provided by Lumina Copper were represented accurately and suitable for use in resource estimation.  There are no environmental issues existing or anticipated that could materially impact the ability to develop the Taca Taca mine.  There are no known factors related to metallurgical, environmental, permitting, legal, title, taxation, socio-economic, marketing, or political issues which could materially affect the mineral resource estimates.  Safety factors and probabilities of failure (pf) estimated for the final pit slopes at Taca Taca are well within acceptable limits as defined by the state-of-practice rock mass strength considerations. The recommended overall pit slope templates are more conservative by rock mass strength because the designs are based on bench configurations controlled by structural fabric. Consequently, there is a future opportunity to optimize overall slopes by incorporating controlled blasting and/or single benching to steepen bench face angles as well as an opportunity to lessen the degree of pit slope depressurization.  The metallurgical test work undertaken is reasonably extensive and suitable for this level of study. The comminution data are considered adequate for a conceptual milling circuit design. The design of the processing circuits is based on this test work data in conjunction with assumptions based on typical industry values.  The Taca Taca mineralized material is of moderate competency and hardness, and amenable to grinding in a conventional SABC. The mineralogy is fine grained and test work indicates a requirement to re-grind to a fine particle size to achieve adequate liberation for flotation as is common within the industry.  Recoveries over the LOM are expected to range from 88% to 92% for copper, 56% to 57% for molybdenum, and 61% to 65% for gold, which will be contained in the copper concentrate. LOM metal recoveries are expected to average approximately 90% for copper, 64% for gold, and 57% for molybdenum.  The cost to construct a 144 km transmission line has been included in the initial capital cost estimate. This power line will connect the Project to a 1,000 MW capacity line that was previously used to export power from the TermoAndes S.A. power plant located near the city of Salta. The Salta power plant is comprised of two 365 MW gas fired generating units that use gas sourced from the gas fields located in northern Salta Province. The plant currently has approximately 200 MW of excess generating capacity that is considered to be sufficient for the Project’s power requirements as defined in this PEA.

 The railway line connecting Antofagasta and the city of Salta is located approximately 10 km from the Project. The railway line has the capacity to handle the transportation of all of the concentrates produced at the Project. The construction of a load-out facility and rail spur near the concentrator has been included as part of the capital cost estimate. It has been assumed that the Project will operate its own fleet of locomotives and rolling stock, which have also been included in the Project costs, to transport the concentrates and consumables between the Project and the Port of Mejillones, 65 km to the north of Antofagasta.  A water supply and water balance analysis were completed for the Project. This analysis, supported by metallurgical test work using water with varying salinity levels, derived a processing make-up water flow sheet comprising a combination of high salinity water from the neighbouring salar, fresh water derived from wells, and desalinated brackish well water. Conceptual design, capital and operating costs for the construction and operation of a water treatment plant were completed for the PEA. Based on the current studies, the processing water flow sheet shows that there are adequate sources of water available for the Project. Additional prospective areas have recently been identified near the Project that may be sources of additional fresh water, which could further optimize the process water flow sheet.  The Project has been designed to meet World Bank Guidelines for social and environmental management practices. Baseline studies completed to date have included surface and groundwater quality, acid rock drainage, meteorological, and social. Provisions have been made within the mine plan and operating costs to account for the environmental protection and rehabilitation of the site once mining has been completed to meet the World Bank Guidelines.  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 characterized as mineral reserves. There is no certainty that these inferred resources will ever be upgraded or that the results stated in this PEA will be realized. Mineral resources that are not mineral reserves have no demonstrated economic viability.

1.18 Recommendations

Based on the results of the PEA, the Qualified Persons recommend that Lumina Copper complete a Prefeasibility Study (PFS) to further define the Project in order to assess its technical and economic viability and to support permitting activities.

General tasks and estimated costs to complete the PFS are summarized below in Table 1-8.

Table 1-8: Estimated Cost for Recommended Future Work

Estimated Cost Task (US$000)

1. Explore potential extensions of pit mineralization, including 10,000 m of DD drilling at approximately 10 locations, and necessary geologic, $5,000 logistical, assaying, and administrative support ($500/m allowance; 21% IVA included).

2. Update resource model/estimate with additional exploration drilling $50 results, including QA/QC.

Subtotal Exploration and Resource Estimate $5,050

Estimated Cost Task (US$000)

3. Complete Prefeasibility Study a) Complete condemnation drilling for facility siting, including 5,000 $2,500 m of drilling ($500/m allowance; 21% IVA included). b) Additional mine geotechnical data collection and engineering analysis, including 4,200 m of DD drilling at 5-7 locations ($500/m $2,700 allowance; 21% IVA included). c) Revise mine design and production schedule based on updated resource model; upgrade accuracy of mining capital, operating, $150 and sustaining capital cost estimates to PFS level. d) Complete additional hydrogeological field investigations, engineering analysis, and computer modeling in support of mine $1,500 dewatering and project water supply. e) Complete further evaluation of new water feed to the water treatment plant and upgrade capital and operating costs to PFS $15 level. f) Conduct additional metallurgical test work, primarily consisting of spatial, grade, and mineralogic variability testing of representative $750 samples from across the deposit, in numbers accepted by the industry as being statistically representative. g) Complete process and infrastructure design and associated costs $1,200 to a PFS level. h) Complete field geotechnical and laboratory investigations; complete PFS level design and associated costs for geotechnical $725 infrastructure (TSF, WRSF, and roads). i) Perform environmental baseline studies and permitting activities. $250

j) Complete an updated marketing study. $50

k) Complete an updated power supply study. $50 l) Update and upgrade logistics and transportation study to PFS $100 level. m) Perform study management and coordination, execution planning and scheduling, Owner’s cost estimating, and economic $450 evaluation. Subtotal PFS Estimate $10,440 Total Estimated Cost for Recommended Future Work $15,490

1.19 Cautionary Note Regarding Forward-Looking Information and Statements

Information and statements contained in this Technical Report that are not historical facts are “forward-looking information” or “forward-looking statements” within the meaning of applicable Canadian securities legislation and the United States Private Securities Litigation Reform Act of 1995, respectively, and involve risks and uncertainties. Examples of forward-looking information and statements contained in this Technical Report include information and statements with respect to:

 Lumina Copper’s plans and expectations for the Project;  the results of the economic analysis of Lumina Copper’s Taca Taca, including, but not limited to, base case parameters and assumptions, base case analysis, forecasts of net present value, internal rate of return, initial and sustaining capital costs, operating costs and cash flows and sensitivity analyses, taxes and royalties;  Lumina Copper’s plans related to mine and concentrator development and design, operations, equipment, and infrastructure;  Lumina Copper’s proposed throughput, production schedule, LOM estimates, and preproduction stripping requirements;  Capital intensity on initial capital estimates and underlying capital cost and production forecasts;  metal price projections;  Lumina Copper’s plans related to mineral processing and recovery methods;  mineral resource estimates and assumptions;  the potential of Lumina Copper to extend areas of known mineralization;  estimates of long term power, transportation, and labor costs, water requirements and waste to mineralized material stripping ratios;  estimates of mine reclamation and closure costs;  the results and further testing of Lumina Copper’s metallurgical testing programs including, but not limited to estimates of metal recovery rates;  Lumina Copper’s plans relating to exploration and development of the project, including permitting and regulatory requirements related to any such plans;  Lumina Copper’s plans and projected costs to complete additional drilling and a PFS;  Lumina Copper’s plans to meet World Bank’s Group Policy and Performance Standards and Equator Principles regarding social and environmental sustainability;  Lumina Copper’s plans to address environmental compliance, reclamation, and liabilities;  potential to optimize the process water flow sheet;  risks related to performance of the Project and opportunities to improve Project performance; and,  potential opportunities identified in Section 25.22.

In certain cases, forward looking information can be identified by the use of words such as “plans”, “expects”, “is expected”, “budgets”, “forecasts”, “anticipates”, “estimates”, “intends”, “targets”, “scheduled”, “believes”, “appears”, “likely”, “typically”, “potential”, “continue”, “strategy”, or “proposed”, or variations (including negative variations) of such words and phrases or may be identified by statements to the effect that certain actions, events or results, “may”, “could”, “should”, “would”, “will be” or “shall” be taken, occur or be achieved.

Various assumptions or factors are typically applied in drawing conclusions or making the forecasts or projections set out in forward-looking information and statements. In some instances, material assumptions and factors are presented or discussed elsewhere in this Technical Report in connection with the statements or disclosure containing the forward-looking information and

statements. You are cautioned that the following list of material factors and assumptions is not exhaustive. The factors and assumptions include, but are not limited to, assumptions concerning metals prices; cut-off grades; short and long term power prices; transportation, water, and labour requirements; processing recovery rates; mine plans and production scheduling; process and infrastructure design and implementation; accuracy of the estimation of operating and capital costs; applicable tax and royalty rates; open-pit design; accuracy of mineral resource estimates and resource modeling; reliability of sampling and assay data; representativeness of mineralization; accuracy of metallurgical test work; and timely receipt of regulatory approvals.

Forward-looking statements are subject to a variety of known and unknown risks, uncertainties and other factors which could cause actual events or results to differ materially from those expressed or implied by the forward-looking statements, including, without limitation:

 risks relating to copper and other mineral price fluctuations;

 risks relating to estimates of mineral resources, production, purchases, capital and operating costs, decommissioning or reclamation expenses, proving to be inaccurate;

 the inherent operational risks associated with mining and mineral exploration activities, many of which are beyond Lumina Copper’s control, including competition, accidents, and labour disputes;

 risks relating to Lumina Copper’s ability to enforce Lumina Copper’s legal rights under permits or licenses or risk that Lumina Copper will become subject to litigation or arbitration that has an adverse outcome;

 risks relating to Lumina Copper’s project being in Argentina, including political, economic and regulatory instability, expropriation, and financing risk;

 risks related to restrictions on import of mining and plant equipment, supplies, and reagents;

 risks relating to potential challenges to Lumina Copper’s right to explore and/or develop the Project, including title disputes or claims;

 risks relating to mineral resource estimates being based on interpretations and assumptions which may result in less mineral production under actual circumstances;

 risks relating to Lumina Copper’s operations being subject to environmental compliance and remediation requirements, which may increase the cost of doing business and restrict Lumina Copper’s operations;

 risks relating to being adversely affected by environmental, safety and regulatory risks, including increased regulatory burdens or delays and changes of law;

 risks relating to inadequate insurance or inability to obtain insurance;

 risks relating to the fact that Lumina Copper’s properties are not yet in commercial production;

 risks related to the failure of plant equipment or processes to operate as anticipated;

 risks relating to the uncertainty as to whether Lumina Copper will acquire permitting required to further explore and develop the Project and risks related to the permitting timelines;

 risks related to changes in Project parameters as plans continue to be refined;

 risks relating to fluctuations in foreign currency exchange rates, interest rates and tax rates;

 risks relating to Lumina Copper’s ability to raise funding to continue its exploration, development and mining activities; and,

 risks identified in Section 25.2.1 of this Technical Report.

This list is not exhaustive of the factors that may affect the forward-looking information and statements contained in this Technical Report. Should one or more of these risks and uncertainties materialize, or should underlying assumptions prove incorrect, actual results may vary materially from those described in the forward-looking information and statements. The forward-looking information and statements contained in this Technical Report are based on beliefs, expectations and opinions as of the effective date of this Technical Report. For the reasons set forth above, readers are cautioned not to place undue reliance on forward-looking information. Lumina Copper and the Qualified Persons of this Technical Report do not undertake to update any forward-looking information and statements included herein, except in accordance with applicable securities laws.

2 Introduction

2.1 Purpose of the Technical Report

The Taca Taca porphyry copper-gold-molybdenum deposit located in northwestern Argentina in the Puna (Altiplano) region of Salta Province is beneficially owned by Lumina Copper, based in Vancouver, British Columbia, Canada. This Technical Report has been prepared for Lumina Copper by or under the supervision of Qualified Persons within the meaning of NI 43-101 in support of Lumina Copper’s disclosure of scientific and technical information for the Project.

This Technical Report supersedes and provides an update of the Technical Report prepared for Lumina Copper entitled “Taca Taca Property Porphyry Copper – Gold – Molybdenum Project, Argentina, NI 43-101 Technical Report”, dated January 4, 2013, with an effective date of October 30, 2012.

The PEA defines the current overall scope of the Project and provides the information required by Lumina Copper to make decisions regarding further evaluation and development of mining, processing, and infrastructure facilities and provides the basis for the estimates, assumptions, parameters, designs, and criteria included in this Technical Report.

MTB Project Management Professionals, Inc. (MTB), a project management consulting firm, assisted Ausenco in completing this Technical Report under the supervision of the Qualified Persons listed below, and under the overall supervision of Kevin Scott, P.Eng.

2.2 Sources of Information

This Technical Report is based on data supplied by Lumina Copper. The information presented, opinions, and conclusions stated, and estimates made are based on the following information:

 Source documents used for this Technical Report are summarized in Item 27.0 of this Technical Report;  Assumptions, conditions, and qualifications as set forth in the Technical Report. This report is based on drilling and sampling data available as of October 30, 2012. The resource model, including subsequent data validation and review, was completed in November 2012 and released on November 21, 2012 in a press release by Lumina Copper;  Data, reports, and opinions from prior owners and third-party entities; and  Personal inspection and review.

The below-listed Qualified Persons are responsible for the information provided in the indicated items.

William L. Rose, P.E., of WLR Consulting, Inc. (WLRC), is responsible for the information provided in Items 15 and 16 and portions of 1, 2, 21, 22, 25, 26 and 27.

Robert Sim, P. Geo., of SIM Geological, is responsible for the information provided in Item 14, and portions of 1, 7, 8, 9, 10, 25, 26, and 27.

Bruce Davis, Ph.D., FAusIMM, of BDRC, is responsible for the information provided in Items 11, 12, and portions of 1, 7, 8, 9, 10, 14, 25, 26, and 27.

Kevin Scott, P. Eng., of Ausenco, is responsible for the information provided in Items 1, 2, 3, 4, 5, 6, 13, 17, 18, 19, 22, 23, 24, and portions of 21, 25, 26, and 27.

Scott C. Elfen, P. E., of Ausenco, is responsible for the information provided in Item 20 and portions of Items 1, 4, 17, 18, 21, 25, and 26.

2.3 Personal Inspection of the Taca Taca Project

The below-listed Qualified Persons conducted personal inspections of the Project as indicated:

Kevin Scott conducted a site visit on October 23-24, 2012 for the purposes of understanding the local conditions as they relate to process plant and infrastructure development activities and to verify other information required for the Technical Report.

William Rose completed a site visit January 21-22, 2011. The purpose of the site visit was to review progress and area geology, view core samples, observe drill sites, and inspect topography.

Robert Sim visited the site from July 1-3, 2008 and again from June 21-23, 2012; he reviewed drilling activities, inspected core from numerous drill holes, reviewed sampling procedures, and visited a series of drill sites on the property. On June 23, 2012, Robert Sim also visited Lumina Copper’s core facility and offices in Salta where he examined core from recent drilling and reviewed Lumina Copper’s sampling procedures.

Bruce Davis visited the site from January 20-22, 2011 and again from June 21-23, 2012; he reviewed Lumina Copper’s sampling procedures, inspected core from select drill holes, and visited select drill sites. On June 23, 2012, Bruce Davis also visited Lumina Copper’s core facility and offices in Salta where he examined core from recent drilling and reviewed Lumina Copper’s sampling procedures.

Scott Elfen completed a site visit January 21-22, 2011. The purpose of the site visit was to gain an understanding of the local conditions as they relate to Project infrastructure development activities and Project environmental and social concerns.

2.4 Currency Assumptions

All measurement units used in this report are metric, and currency is expressed in US dollars ($) unless stated otherwise. The currency used in Argentina is the peso (ARS $): the official exchange rate at the effective date of this Technical Report was approximately 5.04 ARS $ per $1.

The effective date for this Technical Report and the mineral resource estimate is April 9, 2013.

3 Reliance on Other Experts

The Qualified Persons have not independently conducted any title or related searches, but have relied upon a legal title opinion to Lumina Copper from Zaballa-Carchio Abogados based in Buenos Aires, Argentina, dated May 2, 2013 concerning the ownership and title to the Project (described in Items 1.3 and 4.2). The Qualified Persons believe this data and information are essentially complete and correct to the best of their knowledge and that no information was intentionally withheld that would affect the conclusions made herein. The authors have not researched the property title or mineral rights for the Project and express no legal opinion as to the ownership status of the property.

4 Property Description and Location

4.1 Property Location

The Taca Taca property is located in the Puna (Altiplano) region of Salta Province, northwest Argentina at an approximate elevation of 3,585 masl. It is approximately 230 km west of the city of Salta and 55 km east of the Chilean border (Figure 4-1). The center of the property is at latitude 24.7oS and longitude 68.0oW. The UTM coordinates are 7283500 N and 2628000 E (geographic projection: Gauss-Kruger POSGAR 94/Argentina WGS84, Zone 2).

Figure 4-1: Property Location Map (Afrainlle, June 2012)

4.2 Property Ownership and Agreements

numbers, area, status, and contractual royalty encumbrances are listed in Table 4-1 and the present property position is shown in Figure 4-2.

Table 4-1: Mining Concessions

Contractual Concession File Number Area (ha) Status Royalty Mina Fruso Corriente Claim 18.646/2007 4,502.57** Granted property Pending Mina Fruso Corriente Sur Claim 21.956 1,002.57** application Mina Corriente IV Claim 19.716/2009 3,493.84 Granted property Mina Fruso Corriente II Claim 18.685/2007 2,504.00 Granted property Mina Corriente III Claim 19.715/2009 2,426.63 Granted property Mina Amira Norte Claim 18.832/2007 1,500.15 Granted property 1.5% NSR Granted property 0.75% NSR Mina Francisco 1 Claim 18.048/2005 1,313.49 (50%) Granted property 0.75% NSR Mina Francisco 2 Claim 18.049 1,000.22 (50%) Mina Corriente V Claim 20.821/2011 522.96 Granted property Mina Amira Claim 18.794/2007 434.16 Granted property 1.5% NSR Mina Taca Taca 9 Claim 15.949/1997 376.46 Granted property 1.5% NSR Mina La Sarita Claim 1434/1942 167.99 Granted property Mina Corriente I Claim 19.694/2009 134.44 Granted property Mina Amira Este Claim 19.249/2008 81.05 Granted property 1.5% NSR Mina Corriente II Claim 19.693/2009 71.88 Granted property Mina Federico Claim 9078/1974 39.98 Granted property Mina Don Ramón Claim 18.851/2007 26.51 Granted property Grupo Minero Taca Taca Claim 18.690 2,559.96 Granted property 1.5% NSR* • Mina Carla Claim 14.460 Granted property •Mina Paula Claim 14.461 Granted property •Mina Punilla V Claim 15.478 Granted property •Mina Tacalto 6 Claim 15.727 Granted property •Mina Tacalto 8 Claim 15.834 Granted property •Mina Taca Taca 1 Claim 7.578 Granted property •Mina Taca Taca 2 Claim 7.579 Granted property •Mina Taca Taca 3 Claim 7.580 Granted property •Mina Taca Taca 4 Claim 7.581 Granted property •Mina Taca Taca 5 Claim 7.582 Granted property •Mina Taca Taca 6 Claim 7.583 Granted property •Mina Taca Taca 7 Claim 7.584 Granted property •Mina Taca Taca 8 Claim 15.948 Granted property Cateo Claim 21.227/2011 8,924.29 Granted property Cateo Claim 21.226/2011 7,752.04 Granted property

Contractual Concession File Number Area (ha) Status Royalty Cateo Claim 21.225/2011 6,262.51 Granted property Cateo "Aracar" Claim 21.386 927.75 Granted property Mina “ La Escondida” Claim 17.642 37.00 Granted property Mina “La Escondidita” Claim 17.879 6.00 Granted property Mina La Gloria Claim 21.307 196.88 Granted property 1.5% NSR Granted –held in Mina Don Francisco Claim 18.034 trust for CASA by 340.04 S. Arbeleche Pending Land Use application Claim 21.679 90.46 application

Cateo Claim 21.705 4,760.00 Granted property Cateo Claim 21.709 5,599.00 Granted property Pending Cateo "Chuculaqui" Claim 21.387 3,154.12 application Cateo " EX3" Claim 21.390 4,599.48 Granted property Pending Johncito Claim 21.498 47.37 application Granted properties total (ha) 60,561.28 Pending applications total (ha) 3,291.95

Grand total (ha) 63,853.23

* A 1.5% NSR is payable in respect of each of the concessions formerly comprising the area covered by the Grupo Minero, other than Mina Punilla V, Mina Tacalto 6, and Mina Tacalto 8. **Mina Fruso Corriente will be reduced to 3,500 ha as soon as the pending application for Mina Fruso Corriente Sur is granted. The 1,002.57 hectares to be transferred from Mina Fruso Corriente to Mina Fruso Corriente Sur are currently granted under Mina Fruso Corriente concession; only the transfer is pending.

A royalty of up to 3% net of smelting/refining, transportation, administrative, and plant processing costs (also known as the “mine mouth” value) is payable to the Province of Salta.

Figure 4-2: Claim Map

As of May 2, 2013, all of the mining and exploration concessions were in good standing and canon fees have been paid through June 30, 2013. The Grupo Minero and other concessions are valid for an unlimited period of time as long as the semi-annual canon payments are made. All exploration concessions will have to be converted to mine properties in 2013 and 2014. The Grupo Minero and mining concessions include the right to exploit, subject to being granted by an environmental permit for exploitation. The exploration concessions include the right to explore for all metals or minerals.

The surface lands covering the Taca Taca property are owned by the Province of Salta and the necessary access permits were granted for the current drilling work. All known mineralized zones are located hundreds of metres within the limits of the Taca Taca property.

Lumina Copper first acquired an interest in the Taca Taca property when shareholders of Global Copper Corp. (Global Copper) approved a corporate reorganization effective August 1, 2008 by way of a statutory plan of arrangement (Global Arrangement); pursuant to the Global Arrangement, Teck Resources acquired all Global Copper’s shares. Global Copper’s assets, excluding the Relincho Project in Chile, were transferred to Lumina Copper; this included ownership of Minera Corriente which at the time indirectly held a 100% interest in the Taca Taca property as it was then structured. Effective August 19, 2012, Minera Corriente Chile S.A. (Minera Corriente) was wound up into Lumina Copper, leaving Lumina Copper the beneficial owner of all of the issued and outstanding shares of CASA.

Since completion of the Global Arrangement, CASA has subsequently acquired additional mineral concessions through a combination of purchases from third party owners, lotteries, and staking. The present property position is shown in Figure 4-2 and concessions are listed in Table 4-1.

Some of the mining concessions that form the Grupo Minero and one ancillary mining concession are subject to a contractual royalty of 1.5% NSR (Taca Taca Royalty). Franco Nevada Corp., through a wholly-owned subsidiary, holds the right to receive a 72% interest in the Taca Taca Royalty, and the remaining 28% interest is held by two individuals. A number of additional mining concessions are subject to contractual royalties of up to 1.5% of NSR. Table 4-1 includes a summary of the contractual royalties that apply with respect to the mining properties and rights comprising the Taca Taca property.

A royalty of up to 3% net of smelting/refining, transportation, administrative, and plant processing costs (also known as the “mine mouth” value) is payable to the Province of Salta.

Existing surface rights are adequate for the facilities planned for the Project at this time.

4.3 Environmental Liabilities and Permitting

The 1995 Environmental Protection Mining Code of Argentina requires that each Provincial government monitor and enforce the laws pertaining to sustainable development and protection of the environment. A party that wants to modify or begin any mining-related activity as defined by the Mining Code (prospecting, exploration, exploitation, development, preparation, extraction, storage of mineral substances, property abandonment, or mine closure activity) must submit an application to the Provincial Environmental Management Unit (PEMU) and obtain an approved Informe de Impacto Ambiental or Environmental Impact Assessment (EIA) prior to the start of work (Bastida, 2002). Each EIA must describe the nature of the proposed work, its potential risk to the environment, and the measures that will be taken to mitigate that risk. The PEMU has a 60-day period to review and either approve or reject the EIA; however, if the PEMU has not responded within 60 days, that does not constitute an approval (Bastida, 2002). If the PEMU deems that the EIA does not have sufficient content or scope, the party submitting the EIA is granted 30 days to resubmit their document.

If accepted by the PEMU, the EIA is used as the basis to create a Declaración de Impacto Ambiental or Declaration of Environmental Impact (DEI) to which the party must swear to uphold during the mining-related activity in question. The DEI must be updated at least once every two years. Sanctions and penalties for DEI non-compliance are outlined in the Environmental Protection Mining Code, and may include warnings; fines; a suspension of the Environmental Quality Certification; restoration of the environment; temporary or permanent closure of activities; and/or, removal of authorization to conduct mining-related activities.

CASA filed updated EIA documents and received approval to proceed with the 2010, 2011, and 2012 drilling programs.

The permitting requirements anticipated to develop and operate the Project were developed by Vector Argentina with input from Argentine legal counsel. Together, they produced the “Taca Taca Permitting Handbook”, which is intended as a guide to permitting. The overall permitting timeframe is expected to take 12-18 months.

There are no known environmental liabilities currently existing on the Taca Taca property. In the unlikely case of abandonment of the mineral rights, closure activities may require restorative action to the original surface: filling trenches and removing the drill platforms.

Environmental and permitting activities are discussed in more detail in Item 20.0.

4.4 Other Significant Factors and Risks Affecting Access or Title

The Qualified Persons of this Technical Report are unaware of any other significant factors and risks that may affect access, title, or the right or ability to perform work on the property.

5 Accessibility, Climate, Local Resources, Infrastructure, & Physiography

5.1 Location and Access

The Project is located in the remote Puna (Altiplano) region of Salta Province in northwest Argentina at an approximate elevation of 3,585 masl. The center of the Project is at latitude 24.7oS and longitude 68.0oW. The UTM coordinates are 7283500 N and 2628000 E (geographic projection: Gauss-Kruger POSGAR 94/Argentina WGS84, Zone 2).

Access by road and rail to the Project is discussed in detail in Item 5.5.

5.2 Physiography and Vegetation

The Project is located in the Puna (Altiplano) region of western Salta Province. It is located on the east side of the Sierra de Taca Taca and the western edge of the Salar de Arizaro. The topographic relief is low to moderate and has two prominent 3,700 m hills: Cerro de Cobre and Cerro Agua del Desierto. The Salar de Arizaro is at an elevation of 3,470 m. The property has many flat areas to accommodate a variety of site layouts.

Vegetation is sparse to nonexistent in the Project area.

5.3 Climate and Topography

The Project is located next to the Salar de Arizaro. Most of the surrounding terrain is rugged and steep, typical of the central Andean Altiplano, known as the Puna in Argentina. The elevations of the property range from approximately 3,480 to 3,770 masl, and approximately 30 km to the north is the volcano Cerro Aracar.

The climatic data for the area is limited and new meteorological stations at Quevar and Rio Grande have only been in operation for the last three to four years and are privately owned. Lumina Copper installed a weather station at the Project in 2010.

Site data has been compared to historical data for the region to develop the Project climatology. The climate for the Project is typical for the region; very arid with a low average annual precipitation of approximately 110 mm/y and high annual potential evaporation rate of 2,500 mm/y (calculated). The average relative humidity is approximately 34% and temperatures range from -110C to 200C, with January being the warmest month and July being the coldest month. Wind speeds range from 3.8 to 23.2 m/s, predominantly from the northwest. Although westerly winds are generally strong, particularly during the winter months, exploration activities can be carried out year round and are not significantly hindered by local climatic conditions.

5.4 Local Resources

The village of Tolar Grande (population under 150) is located 35 km east of the Project. Tolar Grande may provide some manual labour, housing and cooking facilities. The city of Salta (population 535,000) is the nearest major centre in Argentina and can provide basic goods and services for all stages of exploration and mining. Salta’s airport receives daily flights from Buenos Aires and numerous other destinations in Argentina. The city of Antofagasta, Chile, which is approximately 470 km west of the Project, has a deep-water port that is used by many mines in northern Chile.

5.5 Regional Infrastructure

Although the Project is located in a remote area of northwestern Argentina, there is reasonably good access to regional infrastructure. A network of paved and gravel roads from Salta to the towns of San Antonio de los Cobres and Tolar Grande provide road access to the Project. The road continues 55 km west beyond the Project to the Socompa Pass on the Chilean border and eventually to the port city of Antofagasta, Chile.

The Project is located within 10 km of the railway line that connects Salta with Antofagasta. Rehabilitation of approximately 40 km of track between the Project and Socompa will be required, but it is envisioned that this railway will be used for concentrate transport to Chile and that consumables may also be transported by empty railcars returning to the Project.

Electrical connection to the national power grid is available at the Olacapato substation along the existing Guemes-Chile 345 kV transmission line, north of the Project. A new 144 km long 345 kV electrical transmission line from Olacapato to the Project is proposed.

It is expected that the Project’s water requirements can be met through accessing this subsurface water from a series of well fields. It has been assumed that water with varying amounts of salinity and total dissolved solids (TDS) will be pumped from multiple wells at three sites in HDPE pipe to process water tanks at the concentrator. A water treatment plant using reverse osmosis, as well as settling ponds, chlorination systems, and multi-media filtration is contemplated for the Project.

Project infrastructure is described in detail in Item 18.0.

6 History

Previous exploration on the Taca Taca property is summarized in Table 6-1 and discussed in more detail in Amended Taca Taca Technical Report (SIM, 2010).

Table 6-1: Exploration and Ownership History of the Taca Taca Project

Year Company Description

Fabricaciones Late 1960s Discovery of porphyry copper mineralization at Taca Taca. Militares 1975 Falconbridge Drilled three holes into leach cap and abandoned property. Taca Taca S.A. (TACSA) acquired tenements over Taca Taca prospect, entered into an exploration agreement with Recursos Americanos Argentinos (RAA) and explored the property with GAMSA, 1990-1995 GAMSA (Gencor) a subsidiary of Gencor. Drilled 18 RC holes. Tested porphyry copper mineralization and an area of copper-gold veins north of the porphyry zone. GAMSA returned the property to RAA, who in turn returned the property to TACSA in 1995. Corriente Resources Inc. (Corriente) signed an exploration agreement 1995 Corriente with TACSA. Corriente formed a joint venture with BHP Minerals (BHP) in 1996. Mapping, geophysics (36.8 km of TEM surveying), geochemistry, and 1996-1997 Corriente/BHP drilling. Discovered supergene mineralization at base of leach cap. Target did not meet BHP’s corporate criteria and the property was returned to Corriente in 1997. Corriente acquired all shares of TACSA, which merged into Corriente Argentina S.A. (CASA). Mapping, trenching (130 backhoe trenches), 1998-1999 CASA geochemistry, and drilling. Tested for shallow supergene and exotic copper mineralization. Rio Tinto options property from CASA. Mapping, geophysics [ground magnetics (136 km), radiometrics (K/Th)], and drilling, Tested for 1999 Rio Tinto remnant oxide copper, supergene and exotic copper. Targets did not meet Rio Tinto size criteria, option with CASA terminated in 1999. Acquires property after acquiring 100% interest in CASA, sampling of 2003 Lumina Copper (1) surface oxide copper zones. Acquires property after corporate reorganization of Lumina Copper 2005 Global Copper Corp.(1) Rio Tinto options property from Global Copper. Mapping, radiometric dating, spectral analysis, and drilling. Tested for deep hypogene 2007 Rio Tinto copper-molybdenum mineralization. Results of drilling unfavourable and property returned to Global Copper. Acquires property when Global Copper transferred its assets 2008 Lumina Copper (2) (excluding the Relincho Project) to Lumina Copper prior to being acquired by Teck Resources Limited. Lumina On 19 August 2012, Lumina Copper became the beneficial owner of all 2012 Copper/CASA issued and outstanding shares of CASA.

(1) Lumina Copper Corp. formed in 2003, reorganized in 2005, and changed name to Regalito Copper Corp. (2) Lumina Copper Corp. formed in August 2008.

Additional details concerning exploration and drilling are presented in Items 9.0 and 10.0 of this Technical Report.

7 Geological Setting and Mineralization

7.1 Regional Geology

The Taca Taca porphyry copper-gold-molybdenum deposit is hosted in the southern half of a 50 km long Ordovician batholith, which forms the Sierra de Taca Taca mountain range (Figure 7-1). The batholith consists of coarse-grained granite that is cut by several aplite dykes. This Early Paleozoic intrusion is intruded by Late Permian granites and aplites and overlain by Late Permian sediments and volcaniclastics. Narrow, north-south striking, steeply dipping rhyolitic dykes of Permo-Triassic age outcrop throughout the region. Oligocene rhyodacitic intrusions of the Santa Inés Formation are responsible for the porphyry copper mineralization and alteration at Taca Taca.

The Sierra de Taca Taca is interpreted to be an uplifted block of Paleozoic intrusive rocks. Oligocene volcanics that are exposed to the west of the property dip to the west. This suggests that the Sierra de Taca Taca was uplifted with an eastern convergence along a major, high angle reverse fault located near the western border of the Salar de Arizaro. Regional evidence suggests uplift occurred during the Oligocene (Almandoz, 2008).

Figure 7-1: Regional Geology (Almandoz, 2008)

7.2 Local and Property Geology

7.2.1 Lithology The surface geology of the Taca Taca property is presented in Figure 7-2.

Figure 7-2: Surface Geology of the Taca Taca Project (Almandoz, 2008)

The Taca Taca deposit is hosted by pink, coarse grained, porphyritic granite (Chavez, 2008) or granodiorite (Cornejo, 2008) of Ordovician age (441.5 +/- 3.4 Ma U-Pb date from zircons). It exhibits an equigranular texture and is composed of phenocrysts of plagioclase, quartz (2-4 mm “eyes”), K feldspar, and rare rutile after amphibole and biotite. This intrusion is cut by several aplitic and aplo- granitic dykes that formed at the same time as the Ordovician granites. Minor foliated dolerite dykes are interpreted as the final stage in the formation of the Ordovician batholith (Sillitoe, 2008).

Narrow, north-south striking rhyolitic dykes occur mainly in the eastern part of the Project and minor thin acidic dykes occur in the west part of the Project. These are related to the Choiyoi volcanic event of Permo-Triassic age (262.4 +/- 2.3 Ma U-Pb date from zircons).

These older lithologies are cut by a number of northeast-southwest striking, steeply dipping porphyritic rhyodacitic dykes which coalesce at shallow depths and are interpreted as the source pluton for the porphyry mineralization. These dykes have an Oligocene age (29.30+/-0.57 Ma-U/Pb date from zircons) which is contemporaneous with the porphyry copper mineralization. This lithology is characterized by large plagioclase, K-feldspar, and quartz phenocrysts.

Two different Oligocene intrusive events were recognized:

1. Early-stage rhyodacite that is characterized by crowded phenocrysts (up to 1 cm) of feldspar and quartz hosted in a biotite-rich groundmass. This unit has a strongly developed stockwork of early white to grey quartz veins.

2. Late-stage intermineral rhyodacite is characterized by fewer phenocrysts of feldspar and quartz, an aplitic groundmass and quartz veinlet xenoliths. This unit has a weakly developed quartz stockwork.

Both rhyodacitic phases are strongly altered, but have low-grade copper mineralization.

7.2.2 Alteration

Hypogene Alteration

Hydrothermal alteration associated with the Taca Taca copper-gold-molybdenum porphyry is typical of the Andean porphyry systems. Alteration types include potassic, propylitic, and phyllic, beginning with the earliest phase and progressing to assemblages that overlap or occur later in the development of the hydrothermal system, respectively. Descriptions of the following alterations are taken from Almandoz (2008).

Potassic: This is characterized by abundant, flaky, secondary biotite replacement of mafic minerals and rare secondary K feldspar that occurs as selvages of early veins. Potassic alteration occurs as remnant rafts in the central part of the mineralized zone due to a strong phyllic (sericite-quartz) alteration overprint.

Propylitic: It is characterized by illite-chlorite alteration of feldspars and mafic minerals with minor epidote alteration of plagioclase. Pyrite is common and varies 3-5%, although locally can be up to 10%. Propylitic alteration occurs on the peripheral edges of the hydrothermal system.

Phyllic: Phyllic (sericite-quartz) alteration is the most widely distributed and pervasive alteration phase associated with the Taca Taca porphyry copper-gold-molybdenum mineralization. It is exposed over an area measuring 3.5 km by 2 km. Two stages of phyllic alteration are present:

 An early phase is characterized by the presence of pale green sericite and quartz. The pale green sericite is related to an intermediate sulphidation mineral assemblage, which is characterized by chalcopyrite, minor pink bornite, and virtually no pyrite. The highest hypogene copper grades are directly associated with this alteration type.  A late phase of phyllic alteration overprints potassic, propylitic, and green sericite phyllic alteration phases. It is characterized by coarse white sericite that completely replaces feldspar and mafic minerals. Pyrite commonly occurs as disseminations and in veinlets. The white sericite indicates a change in the sulphidation state of the mineralizing fluid from intermediate to high sulphidation. This change of sulphidation may be explained by cooler temperatures which produce more acidic, hydrothermal fluids.

Supergene Argillic Alteration

A well-developed, thick (150-300 m) leach cap overlies the porphyry copper-gold-molybdenum mineralization. It is characterized by abundant secondary kaolinite and hematite-jarosite fractures that replaced pre-existing sulphide veins. Copper oxides are rare although brochantite is common at the base of the leach cap and within a restricted area about the summit of Cerro de Cobre. The base of the leach cap is sub-horizontal and well-defined. A number of sub-vertical structures with supergene alteration were seen at depths up to 800 m below the surface. Secondary kaolinite, silica (chalcedony), alunite, and chalcocite are present in these structures.

7.3 Mineralization

There are three main styles of mineralization associated with the Taca Taca copper-gold- molybdenum porphyry: supergene/hypogene porphyry copper mineralization, remnant oxide copper-gold mineralization in the leach cap and hematite-quartz copper-gold veins. Each style of mineralization is discussed in more detail within this section.

Re-Os dating of the molybdenite has shown that the porphyry mineralization is Oligocene in age. The paragenesis of the mineralization and rock units is summarized in Figure 7-3.

Figure 7-3: Paragenesis of Mineralization at Taca Taca (Almandoz, 2008)

Supergene/Hypogene Porphyry Copper Mineralization

Hypogene sulphide mineralization consists of chalcopyrite and pyrite with lesser bornite, chalcocite, digenite, and molybdenite occurring as disseminations and in quartz vein stockworks. Most of the copper mineralization is hosted by Ordovician granite and associated aplite and aplogranite dykes. Minor dolerite dikes have high copper contents due to the abundance of mafic minerals containing ferrous iron, which facilitates the precipitation of copper from the hydrothermal solutions. Molybdenite is more common in the aplite dykes than the granite and occurs primarily in early quartz veins. In the central part of the system, total sulphide content is relatively low ranging from 3% to 5%. A pyritic shell with up to 10% sulphide content is peripheral to the copper-molybdenum hypogene core (Almandoz, 2008). Based on the present drilling, the hypogene porphyry mineralization has a northeasterly trend with dimensions of 3.0 km north-south by 2.7 km east-west (Figure 7-4). The sulphide mineralization remains open at depth in some areas and along the southern boundary of the deposit’s northeastern limb.

The relationship of the hypogene sulphide zone to the overlying supergene enriched zone and leach cap is shown in Figure 7-5.

Figure 7-4: Extent of Hypogene Sulphide Mineralization (Wells, 2012)

Figure 7-5: Vertical East-West Oriented Cross Section 728 3950 N (Wells, 2012)

In the potassic alteration phase, minor chalcopyrite with subordinate bornite is associated with secondary biotite. The strong “A” type quartz vein stockwork is essentially barren of sulphides. Milky quartz “B” veins commonly contain molybdenite with subordinate chalcopyrite.

The two phases of phyllic (quartz-sericite) alteration are associated with the highest hypogene copper and gold grades. The sulphides are commonly disseminated in sericitic vein selvages, microfractures, and intergrown with the quartz veins. The early green sericite is associated with chalcopyrite-bornite and has the highest copper grades and above-average gold grades. The green sericite is a high temperature mineral associated with low sulphidation mineral assemblages. The late quartz and white sericite phase is associated with pyrite-bornite and pyrite-chalcocite-covellite sulphide assemblages. Copper grades decrease slightly and the gold grades are approximately half that seen in the early green phyllic alteration phase. The white sericite is formed at lower temperatures and is associated with the high sulphidation mineral assemblages.

In the central part of the porphyry system, the leach cap is strongly developed, but the supergene enrichment blanket is thin (< 5 m) or virtually absent. Two thicker zones (> 100 m) of supergene copper enrichment are present (Figure 7-6). One zone is associated with the West Fault and the second is located in the northeastern part of the deposit. The northeastern zone is up to 300 m thick. Copper mineralization in the Supergene Zones is dominantly fine-grained, black chalcocite with minor covellite. In addition to the stratiform supergene enrichment zones, several deep (> 500 m), steeply dipping supergene enriched structures were identified.

Figure 7-6: Extent of Supergene Enrichment Zone (Wells, 2012)

Remnant Oxide Copper-Gold Mineralization in the Leach Cap

The leached portion of the porphyry deposit, which ranges from 150-300 m thick, is almost completely depleted of copper mineralization and is dominated by limonite assemblages consisting of hematite, jarosite, and goethite. Remnant zones of copper oxide mineralization consisting of malachite, chrysocolla, atacamite, and brochantite are present, but are limited to small sub- horizontal lenses up to several tens of metres in size. Concentrations of molybdenite and gold, being relatively immobile in supergene weathering environments, are approximately the same as that present in the hypogene zone. Reverse circulation (RC) drilling has helped define the extent of

the gold mineralization in the Leach Zone. In general, the highest gold concentrations occur in the thickest portions of the leach cap and above the best hypogene copper-molybdenum mineralization.

The extent of gold mineralization with grades > 0.2 g/t in the leach cap is shown in Figure 7-7.

Figure 7-7: Extent of Gold in Leach Cap (Wells, 2012)

Hematite-Quartz Copper-Gold Vein Mineralization

Numerous parallel, north-striking and steeply dipping quartz-pyrite veins that were oxidized to quartz-jarosite and quartz-hematite veins occur in the Planicie Norte and Oeste areas (Figure 7-8). The veins are 0.5 m to 2 m thick and consist of quartz with massive to semi-massive pyrite or hematite-jarosite with minor alunite. Chalcocite coatings on the sulphides are common. In the surface exposures, chrysocolla and brochantite occur as weathering products of the chalcocite coatings. Argillic alteration envelopes, comprised of sericite and kaolinite, commonly extend several metres from the veins. Copper found in this zone is secondary in nature, apparently having

migrated northward from the porphyry and re-precipitated on sulphide grains associated with the vein’s alteration selvages as well as within distal pyrite-bearing alteration phases (largely propylitic). A subsidiary of Corriente Resources explored these veins by trenching and shallow drilling. Significant supergene copper mineralization was found in this area by deeper drilling; it remains partially open. Note that there are no resources derived from these gold-copper, quartz-hematite veins included in the mineral resources listed in Item 14.0.

Figure 7-8: Hematite Copper-Gold Veins (Wells, 2012)

7.4 Structure

The structural fabric of the Ordovician granitic host rock is characterized by the presence of discrete but widespread north-northeast and northwest trending, steeply dipping proto-mylonite to mylonite

zones. The emplacement of the Oligocene rhyodacitic dykes, quartz veining related to the porphyry system, fractures, and small-scale faults were controlled by these pre-existing zones of structural weakness.

A vertical, north-northwest striking, normal fault (West Fault) is located in the western part of the Project. The rocks to the west of the fault have a thinner leach zone and are uplifted relative to those in the east. This normal fault was probably active during Miocene times and may have controlled development of a zone of supergene copper enrichment associated with the structure.

8 Deposit Types

Mineralization at the Project is a typical Andean porphyry copper-gold-molybdenum deposit (Lowell and Guilbert, 1970 and Pantaleyev, 1995). Common features of a porphyry deposit include the following:

 Large zones (> 10 km2) of hydrothermally altered rocks that commonly grade from a central potassic core to peripheral phyllic, argillic, and propylitic altered zones.  Generally low-grade mineralization consisting of disseminated, fracture, veinlet, and quartz stockwork-controlled sulphide mineralization. Deposit boundaries are determined by economic factors that outline the mineralized zones.  Mineralization commonly zoned with a chalcopyrite-bornite-molybdenite core and peripheral chalcopyrite-pyrite and pyrite. Enrichment of primary copper mineralization by late-stage hypogene high sulphidation events can sometimes occur.  Important geological controls on porphyry mineralization that include igneous contacts, cupolas, and the uppermost, bifurcating parts of stocks and dyke swarms. Intrusive and hydrothermal breccias and zones of intensely developed fracturing, due to coincident or intersecting multiple mineralized fracture sets that commonly coincide with the highest metal concentrations.  Modification by surface oxidation in weathered environments (for example, Escondida). Low pH meteoric waters generated by the oxidation of iron sulphides, leach copper from hypogene copper sulphides, and oxidized copper minerals such as malachite, chrysocolla, and brochantite and re-deposit copper as secondary chalcocite and covellite immediately below the water table in flat tabular zones of supergene enrichment. The process results in a copper-poor leach cap lying above a relatively thin but high-grade zone of supergene enrichment that caps a thicker zone of moderate grade primary hypogene mineralization.  Presence of precious-metal-rich epithermal, and other quartz vein systems, skarns, and exotic secondary copper deposits formed by the lateral migration of metal in low pH fluids away from the main body of porphyry mineralization.

9 Exploration

Copper-gold-molybdenum porphyry-style mineralization was discovered at the Project in the late 1960s. Since that time, six companies have explored the property completing seven drilling campaigns. These campaigns are summarized in Table 10-1 and Figure 10-1 shows the location of the drill hole collars.

A more detailed review of the historic work is provided in Amended Taca Taca Technical Report NI 43-101 (Sim, 2010) and discussed in Item 6.0. A summary of significant results is included in the following sections.

9.1 Historic Exploration Programs (Non-Drilling)

BHP Minerals (1996-1997)

In 1995, Corriente Resources acquired the property and formed a joint venture with BHP. BHP carried out mapping, geophysical surveys (TEM and IP), and drilling. They outlined a large zone of supergene chalcocite and covellite enrichment on the northwest side of the core of the porphyry mineralization. The enrichment zone lies beneath 200 m to 300 m of leach cap and ranges from 20 m to almost 200 m thick. An estimate of the resource in the supergene blanket was made, but the tonnage was considered too small to meet corporate objectives. The property was returned to Corriente Resources.

Corriente Resources (1998-1999)(CASA)

In 1998 and 1999, Corriente Resources’ exploration activities focused on the gold-copper, quartz- hematite (pyrite at depth) veins located north of the porphyry leach cap. Exploration work included ground magnetic and radiometric surveys, excavator trenching, geochemical sampling, and drilling. Best results were obtained in the Planicie Norte area (for example, hole TK-53: 1.31% Cu and 3.32 g/t Au over 24 m core length). Similar but lower grade mineralization was encountered in the Planicie Oeste area (for example, TK-59: 0.88% Cu and 0.24 g/t Au over 26 m core length).

In 1999, Corriente Resources explored for an exotic copper deposit associated with the Taca Taca porphyry system. They conducted salar sand geochemical and gravity surveys to identify drill targets beneath the Salar de Arizaro.

Rio Tinto (1999 and 2008)

In 1999, Rio Tinto explored for zones of unleached copper oxides and/or unleached “perched” supergene enrichment zones within the near surface portion of the porphyry system at the Project. Evidence of this type of mineralization consisted of mineralized outcrops, trenches, and road cuts. Based on the results of a small seven-hole drill program, Río Tinto concluded that the extent of remnant oxide/supergene mineralization within the leach cap was not large enough to meet their corporate target.

9.2 Exploration by Lumina Copper (2010 - 2012)

In 2010, Quantec Geoscience, on behalf of Lumina Copper, completed a Titan 24 survey over the Taca Taca porphyry mineralized zone to look for deep areas of sulphide enrichment. The survey

was done on four north-east-oriented lines (5 km long) spaced at 400 m intervals. Several deep IP and MT targets were defined and provided some early targets for Lumina Copper’s drilling campaign (Figure 9-1).

Figure 9-1: Titan 24 IP and MT Anomalies (Sim, 2011)

10 Drilling

10.1 Drilling

Six different companies have completed seven drilling campaigns at the Project. These campaigns are summarized in Table 10-1 and Figure 10-1 shows the location of the drill hole collars.

Table 10-1: Drilling History

NUMBER HOLE TOTAL TYPE OF COMPANY YEAR METRES SERIES HOLES

Falconbridge TT01-03 BQ Core 3 529 1975 Argentina S.A.

Gatro Argentina TL01-08 RC 18 1,603 1995 Minera S.A.

TK001-033 RC, NQ core 35 11,483 BHP Minerals 1997

Corriente TK034-047 NQ core 14 3,246 1998 Resources

Corriente TK048-126 RC 80 4,428 1999 Resources

CCR001-007 RC 7 2,732 Rio Tinto 1999

ARI001-002 RC 2 606 Rio Tinto 1999

TTBJ0001-0008 HQ core 8 4,877 Rio Tinto 2008

TTBJ10-01 - 12-136 HQ core 137 89,058 Lumina Copper 2010-2012

TTRC11-01 - 12-97 RC 97 32,055 Lumina Copper 2011-2012

TTEX-01 - 07 HQ core 4 1,696 Lumina Copper 2011

TTGT-01 - 04 HQ core 4 2,404 Lumina Copper 2011

T7 – 21, GW3 HQ core 15 2,206 Lumina Copper 2011-2012

TTTV1 - 11, TW6 HQ core 12 6,094 Lumina Copper 2012

AV-SP4D - 5S HQ core 4 520 Lumina Copper 2012

TOTAL 440 163,537

Figure 10-1: Drill Collar Plan Map – Taca Taca Project (Wells, 2012)

10.2 Historic Drilling

In 1975, Falconbridge drilled three short holes into the leach cap, but did not intersect any significant mineralization. Core from this phase of drilling was discarded. The collars of these holes have been found in the field and the locations have been confirmed through surveying.

Gatro Argentina Minera S.A. (GAMSA)

In 1994, GAMSA drilled 18 short RC holes that targeted epithermal gold mineralization in the Planicie Norte area and porphyry-style mineralization associated with the leach cap. No significant mineralization was encountered and they dropped the property in 1995. Cuttings from the RC holes were discarded; however, the location of some of these holes was confirmed.

BHP Minerals

In 1997, BHP tested the potential for a zone of supergene enrichment beneath the leached cap of the Taca Taca porphyry system. Thirty-five combined RC and NQ diamond drill (DD) holes (TK 001- 033) with a total length of 11,483 m were completed on a 400 m x 400 m grid pattern; this includes two holes, TK015A and 30A, which were abandoned and redrilled. Most of these holes were vertical and their lengths ranged between 91-520 m. Drill core from this program is stored at CASA’s secure core storage facility in Salta. Cuttings from the RC section of the holes were discarded. Drill hole locations were confirmed.

Corriente Resources

In 1998, Corriente Resources drilled 14 NQ DD holes (TK034-047) to test for the presence of shallow supergene enrichment zones in the Planicie Norte area. Most of these holes were drilled to the east at dips of -60°, but a few holes were vertical. The holes ranged between 120-406 m deep. Drill core from this program no longer exists, but the drill hole locations were confirmed.

In 1999, Corriente Resources drilled 80 RC holes (TK048-126) totalling 4,428 m in four different areas: Planicie Norte, Planicie Oeste, the Graben area, and beneath the Salar de Arizaro. The target in all four areas was exotic copper mineralization. The holes were drilled at dips ranging between vertical and -60°, and to depths ranging between 15 m and 117 m. Cuttings from these RC holes are no longer available. Drill hole locations were confirmed.

Rio Tinto

In 1999, Río Tinto completed nine RC drill holes in two different areas. Seven of the holes (CCR001-007), with lengths between 360 m and 408 m, and totalling 2,732 m, were targeted on remnant oxide mineralization within the leach cap in the central portion of the Taca Taca porphyry system. The other two holes (ARI001 and ARI002) were drilled to the east of Taca Taca to explore for exotic copper mineralization underlying the Salar de Arizaro. The RC cuttings are no longer available. Drill hole locations were confirmed.

In 2008, Rio Tinto completed eight HQ drill holes totalling 4,877 m, which targeted deep hypogene mineralization. Major Perforaciones Argentina provided the drilling services using an AVD 600 machine. Drill core from this program is stored at CASA’s secure core storage facility in Salta. Drill hole locations were confirmed.

10.3 Drilling by Lumina Copper (2010-2012)

In 2010, Lumina Copper began a 99,500 m DD program to test the extent of porphyry copper mineralization at the Project and to assess the significance of the Titan 24 anomalies.

During the 2010 field season, Major Perforaciones of Mendoza (Major) was contracted to provide drilling services. They provided one skid-mounted ED50 diamond drill rig. The drilling began on August 3, 2010 and five holes were completed in 2010.

The early drilling discovered extensive porphyry-style supergene and deep hypogene copper mineralization over an area of 2.5 km by 1.5 km. The exploration drilling program was redesigned and expanded to meet the following objectives:

 Delineate the shape and thickness of the supergene enrichment blankets.

 Assess the extent and potential for high-grade hypogene mineralization in the deeper parts of the porphyry system.  Evaluate the gold potential of the leach cap.

Drilling continued until September 2012, with up to seven diamond drills operating on the property. Core drills were supplied by Major, Boart Longyear, and Alta Drilling. Two Schramm T685 WS-C RC rotary rigs from Major were used to evaluate the gold potential of the leach cap and provide pre- collars for the diamond drilling. Since 2010, Lumina Copper has completed 273 holes totaling 134,033 m.

REFLEX or Peewee survey tools were used to provide downhole orientation data for the DD holes. All collar locations were initially located using a handheld GPS (Garmin 60CSx), but, after the hole was completed, the collar location is surveyed using a differential Trimble GPS, accurate to +/- 10 cm.

Drill core and RC rejects from this program are stored at CASA’s secure core storage facilities in Salta. For a further discussion of sampling, see Item 11.0.

11 Sample Preparation, Analyses, and Security

11.1 Historic Drilling and Sample Preparation

Falconbridge and GAMSA

No written record of the sample preparation or analytical methods was available for the Falconbridge (1975 core hole) or GAMSA (1994 RC hole) drilling programs. There is also no record of any QA/QC protocols or results for these two drilling programs.

BHP Minerals

BHP cut the drill core in half using a diamond saw. One-half of the core was put into plastic sample bags and shipped for analysis, and the other half was retained as a permanent record. Most drill core was sampled using 2 m intervals. No written descriptions of any sampling methods for the RC drill campaigns were available for review. Sample cutting piles were observed at several collar locations during the site visit; this indicates the drill cuttings were probably split and sampled on-site before shipment to the assay lab. Samples were analyzed at Bondar Clegg in La Serena, American in Mendoza, and SGS in Salta and Santiago. The analytical methods and sample preparation protocols were not discussed in any of the BHP reports, but the results were presented in a spreadsheet format without original assay certificates. BHP’s quality control protocols involved submitting one-quarter core duplicate sample for every 20 samples and submitting them to the original laboratory and submitting 300 coarse reject duplicates to Bondar Clegg as check assays. The results of the duplicate sampling or coarse reject duplicates were not available for review.

Corriente Resources

Corriente Resources cut the drill core in half using a diamond saw and submitted half of the core for analysis at ALS Chemex Labs (ALS Chemex) in Mendoza. Corriente Resource’s sample preparation protocols were not available for review. Copper analyses were done using atomic absorption spectroscopy and gold was analyzed by fire assay with an atomic absorption spectroscopy finish. Corriente Resources did not collect quality control data, other than a routine submission of pulp duplicates to another lab for check analyses. The check lab data was not available for review.

Rio Tinto

1999

Samples from Río Tinto’s 1999 RC drilling program were prepared by Bondar Clegg’s laboratory in Mendoza using a “Large Pulp Preparation” procedure. The entire sample was crushed to -80 mesh and a 1 kg split of this material was pulverized. Bondar Clegg of Vancouver, Canada analyzed the samples for gold and 34 other elements using fire assay with an atomic absorption finish and total digestion ICP analysis, respectively.

Río Tinto carried out systematic QA/QC procedures. One field duplicate was inserted for approximately every 12 samples to check for splitting and lab errors. The samples were randomized and pulp duplicates, standards, and blanks were added. This QA/QC program adhered to accepted industry standards. Río Tinto’s QA/QC results were not analyzed in detail by Robert Sim, SIM Geological, but they were reviewed in graphical form. SIM Geological concluded that they appeared

to support Río Tinto’s conclusion that “the analysis of the results obtained from the quality control indicates the lab’s performance was satisfactory” (Río Tinto, 1999).

2008

In 2008, Rio Tinto used the following procedures for the eight HQ drill holes:

Core Sampling

Core was placed in boxes by the drill crew and logged by the geology staff at the core shed. The sampling staff photographed the core and marked it with a line drawn down the centre. The core was cut along this line using a diamond rock saw. Samples were taken at 2 m intervals. Half the core was placed in a plastic bag for analysis, and the remaining half was placed back in the core box for reference. A lab-generated sample ticket was inserted in the plastic bag with the sample, and a second ticket was stapled into the throat of the bag. Nylon cable ties were used to seal the bags. The bags were taken from the sawing area to the core shed where the sample number was written on the bag. Each bag was weighed and up to five sample bags were sealed in a larger mesh sack. The sacks were sealed with a large, numbered cable tie and labelled “secured.” Samples were shipped only after all samples from one hole were complete.

Chain of Custody

The sample bags were checked onto the truck, an inventory was sent with the shipment, and a copy kept on-site. Samples were transported from site in a covered pick-up truck, driven by a Rio Tinto staff member. Any tampering with individual bags or bag ties would have been immediately evident when they arrived at the lab: the lab was notified of the sample numbers, and, on arrival, the lab sent confirmation of the samples being received and their condition. No irregularities in any sample shipment were detected during the course of the program.

During the 2008 visit, SIM Geological concluded that all procedures were carefully observed and met or they exceeded industry standards for the collection, handling, and transport of drill core samples.

Analyses

The drill core samples were analyzed by Alex Stewart (Assayers) Argentina S.A., Rodriguez Pena 1140, Luzuriaga, Maipu, M5516 BBX, Mendoza, Argentina. This lab is 9001:2000 certified.

The samples were analyzed for gold using a fire assay/atomic absorption finish on a 30 g charge. An additional 39 elements were analyzed by four acid digestion and an ICP finish; the elements included in the ICP package were: Ag, Al, As, Ba, Bi, Ca, Cd, Co, Cr, Cu, Fe, Ga, Hg, K, La, Li, Mg, Mn, Mo, Na, Nb, Ni, P, Pb, S, Sb, Sc, Se, Sn, Sr, Ta, Te, Ti, Tl, V, W, Y, Zn, and Zr. The copper over-limit values were analyzed by atomic absorption (AAS).

A suite of 133 inter-laboratory check samples was also assayed at ALS Chemex in Lima, Peru; this lab is also ISO 9001:2000 certified.

QA/QC Procedures

During the 2008 Rio Tinto drill program, the sampling staff inserted standards, blanks, and duplicates as specified by Rio Tinto site geologists. A standard was systematically inserted for every 25 core samples. A few (seven or eight) field duplicate samples, which consist of the other

half of the drill core, were also taken. Blank material was inserted at the rate of one in every 20 samples. The blank material consisted of quartzite from a quarry in San Luis Province, Argentina. It was included in the form of rock chips (BLQZ) and pulverized material (BLKL) to test sample preparation and analysis, respectively. In addition, the lab analyzes a split of the coarse and pulp reject at a rate of one per 25 samples.

A suite of 133 pulp duplicate samples from hole TTBJ0003, including four samples of the Altar 3 Gold Standard Reference Material, was sent to ALS Chemex in Lima and analyzed using a comparable four-acid digestion with atomic absorption finish for copper and molybdenum and 30 gram fire assay finish for gold.

Results of QA/QC Work

A detailed review including charts and graphs of the QA/QC data from Rio Tinto’s eight drill holes were presented in Amended Taca Taca Technical Report NI 43-101 (Sim, 2010).

The results are summarized as follows:

 Field duplicate samples were taken to check the geological variability of the sample size. Duplicate samples of coarse reject material were assayed to check the sample preparation protocol. If the protocol was adequate, 90% of the duplicate pairs of assays should fall within ± 30% of each other (pass). A comparison of field duplicate data and coarse reject duplicate data can be used to estimate geological variability. For copper and molybdenum, the sample preparation protocol appeared to be good. For molybdenum, the percentage of field duplicates falling within the ± 30% control limit was 86% which, given the small number of samples (seven), is not of concern and may indicate a higher geological variability than copper. Field duplicate results for gold, however, showed a poor reproducibility which improves markedly once the samples are crushed as indicated by the acceptable coarse reject duplicate results. This indicated that the high variability seen in field duplicate results may also be due to geological variability (also known as the nugget effect).  For the pulp duplicates, over 90% of the pairs falling within ± 10% of each other is (control limit) considered adequate. Results obtained for the program indicated that the sample assay protocol was adequate for copper, but required review for gold and molybdenum.  Rio Tinto used SRM for copper and gold. Most samples plotted within one standard deviation of the accepted values. Consequently, there were no significant issues with the copper or gold analyses. No reference material for molybdenum was specifically used, but the molybdenum results for the copper and gold standards indicated a consistency of results for both standards and no problems with analysis for this element was suspected.  Blank results for copper, molybdenum, and gold were considered acceptable. Some contamination in the sample preparation stage was detected for copper and the five high values (> 100 ppm Cu) came from sample batches containing highly mineralized material ( 2.5% Cu). These spikes most likely resulted from sample to sample contamination at the jaw crusher stage. It was not thought to have affected many samples to a significant extent.

 Sample results for the inter-laboratory check samples were not available, but a visual inspection of graphs provided by Rio Tinto showed an excellent correlation for copper, a good correlation for molybdenum, and an acceptable correlation for gold. For copper, 97% of all checks returned a mean percentage difference of less than 10%. For molybdenum, 87% of all checks returned a mean percentage difference of less than 5%. For gold 72% of all checks returned a mean percentage difference that was less than 20% (Almandoz, 2008). Results from the four SRM samples submitted with this batch were not available, but Rio Tinto reported that results produced for gold were outside acceptable limits (Almandoz, 2008).

Conclusions

Results from SRM indicated that the copper, gold, and molybdenum assay procedures were producing reliable assay data. Blank results indicated that there may be a minor contamination of blank samples by copper, which probably occurred when high grade copper samples contaminated a few samples that followed them through the preparation stage. This was not thought to be significant for resource estimation purposes. Blank material indicated that there was no detectable contamination of the sample preparation or assay procedure for molybdenum and gold.

Duplicate data indicated that the sample preparation stage worked well for copper, molybdenum, and gold. The high variation in field duplicate data for gold indicated a high geological variability (nugget effect). Molybdenum field duplicate variation was not quite as high as for gold.

Inter-lab duplicates showed excellent, good, and acceptable reproducibility for copper, molybdenum, and gold, respectively.

Robert Sim and Bruce Davis concluded that Rio Tinto’s drill core sampling protocols at the Project are similar to industry standard procedures. No recovery information was available for review, but the recovery seemed good for the examined diamond drill core and there is no evidence that diamond drill or reverse circulation recovery could materially impact the drill results.

None of the preceding conclusions are sufficiently important to make the Rio Tinto results unreliable for publication, particularly considering that gold and molybdenum make only minor contributions to the value of the contained metal at present day metal prices. The influence of Rio Tinto’s drilling on the current mineral resource estimate is relatively minor due to the number and distribution of more recent drill holes completed by Lumina Copper.

11.2 Lumina Copper Drilling (2010-2012)

Lumina Copper used the following procedures during their 2010-2012 DD core and RC drill programs:

Core Sampling

The drill contractor places the HQ drill core into wooden boxes (1 m long by 3 rows) at the drill rig. Wooden tags marked with the downhole depth are placed in the box. Lids are placed on the box and it is transported by truck to the core shack. Upon receipt, Lumina Copper field assistants check the depth and mark out the samples at 2 m intervals. Photos are taken of both dry and wet core. Two boxes are included in each photo. Lumina Copper geologists and technicians examine the core and prepare geological and geotechnical logs for each hole. The geotechnical log includes: rock quality data (RQD), core recovery, fracture and vein quantity, vein angles, and density

measurements. Samples are taken at 10 m intervals for point load tests and density measurements. This information is entered directly into an Excel® spreadsheet for each hole.

The core is cut in half using a diamond saw. For each 2 m sample, one-half is put into a plastic bag and the other half is returned to the wooden box for reference. Bar coded sample tags are included in each sample bag and the sample number is written in permanent marker on the sample box. Sample bags are secured and put into a larger mesh sack with a tamper-proof nylon tie. Duplicate and standard samples are included, as required. When a hole is complete, the samples are sent by truck to either ALS Chemex or Alex Stewart in Mendoza, Argentina.

The remaining core is initially stored on pallets at the exploration camp and then moved to CASA’s secure core storage facility in Salta.

RC Sampling

Lumina Copper geologists and assistants are present during the drilling of RC holes and they are responsible for taking the samples for assay. Samples are taken every 2 m. Cuttings are collected from the cyclone and placed directly into the Gilson adjustable sample splitter. The sample is split three times until there are two samples representing approximately 25% of the initial weight. Each sample weighs approximately 6-10 kg. One sample is sent to the lab for analysis and the other is stored for reference or future sampling. A small sample (100 g) is taken from the reject bag and placed on a chip strip where it is visually inspected and logged by a Lumina Copper geologist.

The Gilson adjustable sample splitter and all the tools used in the sampling process were cleaned with compressed air after every sample to reduce contamination.

Major water intersections encountered during drilling are noted by the geologist on site. Wet samples were split using a rotary wet splitter; one-half of the material was collected in a big bucket and left for a reasonable time to decant. After the water is removed, the sample is divided in two; one is sent to the assay lab and the other is put into storage.

Samples, duplicates, standards, and blanks from each hole are sorted at the exploration camp and transported directly to the ALS Chemex facility in Mendoza, Argentina.

Chain of Custody

Lumina Copper uses the same sample procedures as Rio Tinto did in 2008. The sample bags are checked onto the truck, and an inventory of samples is sent with the shipment and a copy is kept on-site. Any tampering with individual bags or bag ties would be immediately obvious when the samples arrived at the lab. The lab is notified of the sample numbers that were sent and on arrival the lab sends confirmation of the samples received and their condition. No irregularities in any sample shipment were detected during the course of the program.

During Bruce Davis’s site visit in 2012, he concluded that all procedures were being carefully observed and met or exceeded industry standards for the collection, handling, and transport of drill core samples.

Analyses

Samples from the Lumina Copper drill programs were sent to: ALS Geochemistry - Mendoza, Altos Hornos Zapla 1605, Mendoza Godoy Cruz, Argentina, (ALS) or Alex Stewart Argentina S.A., Carril Rodriguez Pena, M5516 Maipu, Mendoza, Argentina (Alex Stewart).

Both analytical labs are ISO 9001:2008 certified. The samples were analyzed for gold using a fire assay/atomic absorption finish on a 30 g charge and for another 35 elements by four acid digestion and an ICP finish. The elements included in the ICP package are: Ag, Al, As, Ba, Be, Bi, Ca, Cd, Co, Cr, Cu, Fe, Ga, K, La, Mg, Mn, Mo, Na, Ni, P, Pb, S, Sb, Sc, Se, Sn, Sr, Th, Ti, Tl, U, V, W, and Zn. Samples with copper values >10,000 ppm are analyzed by atomic absorption spectrometry.

QA/QC Procedures

The performance of ALS Chemex and Alex Stewart is monitored through the implementation of a QA/QC program. The results of this program are tracked by Lumina Copper and reviewed by Bruce Davis on an ongoing basis. Irregular or suspect results were addressed in a timely manner to ensure the integrity of the database.

Lumina Copper established a QA/QC protocol that uses reject duplicates, standards, and blanks inserted into the sample batches at regular intervals. Duplicates are inserted during sample preparation (independent from the assay laboratory) by splitting the pulps. A range of copper-gold- molybdenum SRMs, of suitable matrix composition, including blanks, is inserted by Lumina Copper during the core sampling procedure. The structure of this QA/QC program follows accepted industry standards.

There is at least one of the eleven different SRMs, plus duplicates and blanks (QA/QC samples), included in every batch. A QA/QC sample is inserted into the sample stream once every eight samples. The SRMs cover a broad range of copper-gold-molybdenum concentrations encountered at the Project and one is inserted every 24 samples sequentially independent of the blanks and duplicates. A duplicate is selected from samples with sufficient material and inserted once in every 24 samples. A blank is inserted once in every 24 samples. When smaller batches of samples are sent to the lab, containing insufficient samples to maintain this frequency, at least one of the eleven SRMs, a duplicate, or a blank are inserted.

Assay results for the copper, gold, and molybdenum are compared with the accepted values for standards and blanks. Duplicates are compared with original values. Example ALS Chemex control charts for the OREAS 50C SRM are shown in Figure 11-1 to Figure 11-3. The red lines indicate the upper and lower control limits (UCL and LCL) which are defined as ± 10% of the accepted value. The coarse blank for copper control chart appears in Figure 11-4. The UCL for blank material is three times the average detection value of all samples analyzed. The ALS Chemex coarse duplicate performance for copper is shown in Figure 11-5. UCL and LCL for coarse material are defined as ± 30% of the relative difference between both assay values.

Taca Taca QC Results Through December 2012

OREAS 50C Copper

9000

UCL = 8162 8000

AVG = 7523 Cu ppm

7000 Accepted value = 7420 LCL = 6678

6000 0 100 200 300

Sequence Number

Figure 11-1: ALS Chemex OREAS 50C SRM Copper Control Chart

Figure 11-2: ALS Chemex OREAS 50C SRM Gold Control Chart

Figure 11-3: ALS Chemex OREAS 50C SRM Molybdenum Control Chart

Figure 11-4: ALS Chemex Coarse Blank Copper Control Chart

Figure 11-5: ALS Chemex Copper Coarse Duplicate Comparisons

Examples of the Alex Stewart control charts for the OREAS 503 SRM are shown in Figure 11-6 to Figure 11-8; OREAS 50C SRM examples were not available for use at Alex Stewart. The red lines indicate the upper and lower control limits (UCL and LCL) which are defined as ±10% of the accepted value. The coarse blank for copper control chart appears in Figure 11-9. The UCL for blank material is three times the average detection value of all samples analyzed. The Alex Stewart coarse duplicate performance for copper is shown in Figure 11-10. UCL and LCL for coarse material are defined as ± 30% of the relative difference between both assay values.

Figure 11-6: Alex Stewart OREAS 503 SRM Copper Control Chart

Figure 11-7: Alex Stewart OREAS 503 SRM Gold Control Chart

Figure 11-8: Alex Stewart OREAS 503 SRM Molybdenum Control Chart

Figure 11-9: Alex Stewart Coarse Blank Copper Control Chart

Figure 11-10: Alex Stewart Copper Coarse Duplicate Comparisons

All abnormal quality control results were addressed by re-assaying remaining material. Whenever re-assaying was required, the remedial results replaced original assays. No QC failures remain in the database.

Suites of duplicate samples were exchanged between Alex Stewart and ALS. Comparisons of the inter-lab duplicates for copper show very good results with most differences attributed to samples in the low-grade range. Correlations between lab duplicates for gold and molybdenum are not as good, but these results are often attributed to the fact that many of the samples contain very low- grades for these elements. Overall, there is no indication of bias, but precision is generally poor due to the low-grade range of many of the samples. These results have little to no overall effect on the estimation of mineral resources.

Bruce Davis believes Lumina Copper’s drill core sampling protocols at the Project meet accepted industry standards. Recovery is good for the examined diamond drill core and there is no evidence that diamond drill or reverse circulation recovery could materially impact the drill results.

12 Data Verification

12.1 Historic Drilling

Steve Blower, P.Geo, at AMEC’s Vancouver office, visited the Project in April 2003 and authored an NI 43-101 report on the Project dated May 2003. At that time, the data verification consisted of: a collection of several representative samples for independent check assays, a comparison of assay data stored in CASA’s database, and the original assay certificate records.

In 2003, 11 copper and gold check assays were completed on drill core samples collected during AMEC’s visit. The check sample intervals were randomly selected from both hypogene and supergene mineralization, using the same intervals as the previous exploration programs. Robert Sim reviewed this data and concluded both the gold and copper check assay results agree with the original assay results. Differences between the check and original results are small, and the check assays are not systematically higher or lower than the original assays.

Assays for 11 drill holes from Corriente Resources’ spreadsheet database files were checked against the results from the original paper assay certificates. The only certificates available were from Corriente Resources’ third phase of drilling (TK-048 to TK-126). The 11 holes represent 14% of the data in Corriente Resources’ phase 3 campaign, or 7% of the total assay database at the time. All of the assay results in the database were the same as the values on the original certificates.

Robert Sim reviewed AMEC’s validation work conducted on the sample data and he believes that the process and conclusions were appropriate to validate the data.

During property visits in July 2008 and June 2012, Robert Sim compared the assay results and the visual observations of the content of copper and molybdenum-bearing mineralogy in randomly selected intervals from several drill holes. The quantity and type of minerals observed support the assay results in all cases. Drilling activities were observed and numerous drill hole collars from previous drilling programs were seen while visiting the property.

12.2 Lumina Copper Drilling

Robert Sim compared the original assay certificates from ALS Chemex for ten holes from the recent Lumina Copper drilling and the assays listed in the electronic database. There were no errors noted in the database validation.

Robert Sim and Bruce Davis reviewed a series of randomly selected drill core intervals during their site visits in July 2008, January 2011, and June 2012. In all cases, the type and content of observed copper-bearing minerals supported the copper grades present in the database.

Given the assay check results, observation of the drilling and core sampling, and the comparison of certificates to the electronic database, the sample assay data is within acceptable limits of precision and accuracy to generate a mineral resource estimate.

12.3 Conclusions

The Taca Taca database was derived from drilling programs conducted in the 1970s, 1990s, 2008, and 2010-2012. This data was verified using several methods including visual comparisons, resampling, and direct comparisons with assay certificates.

Industry accepted QA/QC programs have only been documented in the drilling programs conducted by Rio Tinto in 1999 and 2008, and by Lumina Copper in 2010-2012. The results of this work indicate that the data is sound. The fact that the drilling results prior to 1999 are not dissimilar to those obtained by Rio Tinto and Lumina Copper indicates this earlier data is also acceptable. Robert Sim and Bruce Davis believe that the sample database is sufficiently accurate and precise to generate estimates of mineral resources.

13 Mineral Processing and Metallurgical Testing

13.1 General

Ausenco has reviewed Taca Taca metallurgical test reports and data to develop process design criteria for the PEA. The test work undertaken is reasonably extensive and considered suitable for this level of study. The comminution data are considered adequate for a conceptual milling circuit design. The design of the processing circuits is based on this test work data in conjunction with assumptions based on typical industry values.

The Taca Taca mineralized material is of moderate competency and hardness, and amenable to grinding in a conventional SAG-ball milling circuit with pebble crushing. The mineralogy is fine grained and test work indicates a requirement to re-grind to a fine particle size to achieve adequate liberation for flotation, as is common practice within the industry.

The Taca Taca concentrator envisions processing 120,000 t/d of ROM mineralized material initially, expanding to process 180,000 t/d in Year 8. Copper and molybdenum concentrates and tailings will be produced. The proposed process includes crushing and grinding of the ROM mineralized material, bulk copper-molybdenum rougher and cleaner flotation, regrinding, copper-molybdenum separation, molybdenum flotation, and dewatering of copper and molybdenum concentrates. The flotation tailings will be thickened before placement in a TSF.

13.2 Metallurgical Testing

The following metallurgical test reports provide the basis for this PEA:

 Plenge Laboratory, Metallurgical Investigation No. 9281-9402, Lumina Copper Corporation Inc. Taca Taca Copper Gold Molybdenum Project Comminution, Copper Molybdenum Separation, Variability Oxide Copper and Gold, October 16, 2012;  Starkey & Associates, S&A Project S117-1, Taca Taca Project SAGDesign® Comminution and Mill Sizing Analysis Report Rev 0, July 11,2012; and,  JK Tech Pty Ltd. SMC Test Report (tested at C.H. Plenge, Lima Peru) Job No. 10199/P2 August 2010.

The Plenge program (October 2012) tested four composites from two material types, S1 and S2 (Supergene) and P1 and P2 (Primary), as defined by the April 2012 resource estimate. The samples were prepared to represent Years 1 to 5, identified as S1 and P1 and to represent Years 6-10, identified as S2 and P2.

The following tests were conducted:

 Bond Crusher Work Index on Supergene and Primary samples  Specific Gravity on Supergene and Primary samples  SAGDesign® Test on Supergene and Primary composites  Unconfined Compressive Strength (UCS) Tests on Supergene and Primary composites

 Bond Rod Mill Work Index on Supergene and Primary composites  Abrasion Index on Supergene and Primary composites  Grindability tests on Supergene and Primary variability samples  Locked cycle tests on each composite S1, S2, P1, and P2 using tap water  Locked cycle test using brine water in rougher flotation and tap water in bulk copper-moly concentrate flotation on each of the composites, and a 1:1 blend of Supergene + Primary (excluding copper and molybdenum separation)  Locked cycle test using tap water on each of the composites, and a blend of Supergene + Primary (excluding copper and molybdenum separation)  Rougher variability tests on 15 Supergene samples  Rougher variability tests on 25 Primary samples  Cleaner variability tests on 15 Supergene samples  Cleaner variability tests on 25 Primary samples  Rougher variability tests on each of the composites, and a 1:1 blend of Supergene + Primary  Batch cleaner flotation tests on Supergene and Primary composites with different water blends for comparison. Varying the water used including 100% tap water, 25% pit water + 75% tap water, 50% pit water + 50% tap water, 75% pit water + 25% tap water  Batch cleaner flotation test with tap water on Supergene + Primary blend composite  Thickening and filtration tests on rougher tailings and copper concentrate products from the tap water locked cycle tests  Thickening and filtration tests on rougher tailings product from Supergene variability tests  Thickening and filtration tests on rougher tailings products from Primary variability tests

Comminution circuit design for Taca Taca was based on results from two JK Drop weight tests conducted in 2010 and the Bond Ball Work Index values from the variability samples and SAG Design®.

Locked cycle tests on each composite using tap water produced copper and molybdenum concentrates. Thickening tests using the Kynch method was performed on each of the copper concentrates and rougher tailings produced in each of the locked cycle tests. Pressure filtration tests were performed on each thickener flotation tailing sample and two tests on each copper concentrate from the locked cycle tests.

Ausenco has reviewed the location of the samples and drill holes and concludes the samples are reasonably representative of the deposit for a PEA given the nature and continuity of rock types and mineralization of the deposit.

Where no test work data are available, reasonable assumptions, based on operating data or test work from other projects has been used to develop the process design criteria used for the plant design.

13.3 Comminution Test Work

Several comminution tests were conducted; the results are summarized in Table 13-1.

Table 13-1: Summary of Comminution Test Results

Supergene Primary Physical Property Unit Average 75th Percentile Average 75th Percentile

Bond Crusher Work Index kWh/t 6.65 - 8.11 -

Bond Ball Work Index kWh/t 17.33 19.06 15.75 16.20 (from SAGDesign®)

Bond Ball Work Index kWh/t 18.95 20.68 16.57 17.76

UCS MPa 12.24 12.71 12.84 14.61

Abrasion Index Ai (g) 0.200 0.213 0.230 0.241

Specific gravity g/cc 2.70 - 2.71 -

SAGDesign® tests were done on four Supergene composites and six Primary composites. The results are summarized in Table 13-2.

Table 13-2: Summary of SAGDesign® Results

SAGDesign® SAGDesign® Sample No. Sample No. (kWh/t) (kWh/t)

Supergene Primary

Supergene Comp 1 16.81 Primary Comp 1 14.99

Supergene Comp 2 15.03 Primary Comp 2 14.94

Supergene Comp 3 18.63 Primary Comp 2 16.08

Supergene Comp 4 20.36 Primary Comp 4 16.60

- - Primary Comp 5 16.29

- - Primary Comp 6 15.58

Average 17.33 Average 15.75

75th percentile 19.06 75th percentile 16.20

Compression tests were done on 25 rocks from each material type composite ranging in size from 13.2 to 16 mm. Bond Ball Work Index numbers were interpreted by comparison from the grindability test data for each material type, 15 Supergene samples and 25 Primary samples.

Abrasion Index tests were conducted on four Supergene composites and six Primary composites.

In 2010, two samples were tested for competency using the JK Drop weight test. The results are summarized in Table 13-3. The value of Axb is a measure of resistance to impact breakage. Due to the low number of samples tested, Ausenco applied a 115% design factor to the Axb for the Primary material which was the most competent. In the future phases of the project there should be more variability testing on material competency to design the SAG circuit which is the higher capital cost area of the plant, as well as in some cases a crucial factor in achieving plant design throughput.

Table 13-3: JKTech – SMC Test Results

Sample No. A b. Axb Primary 66.9 0.91 60.9 Supergene 69.5 0.99 68.8 Design (applying 115% factor) 52.9

13.4 Flotation Test Work

13.4.1 Locked Cycle Tests using Tap Water

Four locked cycle tests using tap water throughout all phases of the test were performed. The results are summarized in Table 13-4.

Table 13-4: Summary of Locked Cycle Tests Using Tap Water

Concentrate Grade Concentrate Recovery % Wt Sample Name Au, % Cu, % Mo, % Au Cu Mo g/t

Supergene S1

Bulk Cu - Mo Concentrate 1.8 5.38 34.5 0.8 60.2 90.6 69.7

Cu Concentrate 1.8 5.6 35.3 0.13 57.8 87.9 8.0

Mo Concentrate 0.02 1.2 1.9 45.3 0.2 0.1 49.6

Supergene S2

Bulk Cu - Mo Concentrate 2.3 4.56 33.3 0.68 67.7 90.2 85.8

Cu Concentrate 2.2 4.5 32.8 0.15 64.1 86.2 13.6

Mo Concentrate 0.02 1.10 0.95 49.6 0.1 0.02 62.3

Primary P1

Bulk Cu - Mo Concentrate 2.11 6.4 28.4 0.7 71 95 80.6

Cu Concentrate 2.1 6.4 28.4 0.7 71 95 7.5

Concentrate Grade Concentrate Recovery % Wt Sample Name Au, % Cu, % Mo, % Au Cu Mo g/t

Mo Concentrate 0.2 0.9 1.3 53.2 0.1 0.0 73.0

Primary P2

Bulk Cu - Mo Concentrate 1.48 5.0 29.3 0.7 60.0 91.4 65.1

Cu Concentrate 1.46 5.0 29.6 0.2 59.8 91.2 18.0

Mo Concentrate 0.02 0.9 4.4 41.5 0.1 0.2 47.2

These four locked cycle tests form the basis for the process design criteria for the Project. They represent the two mineralized material types and Year 1-10 in a conceptual mine plan based on the April 2012 resource estimate, which was developed for internal use. In all four tests copper and molybdenum concentrates were produced. The average copper recovery for Supergene material was 87% with an average copper grade of 34%. The average copper recovery for Primary material was higher at 93.5% with a copper concentrate grade of 29%. The average molybdenum recovery for Supergene material was 56% with an average molybdenum grade of 47%. The average molybdenum recovery for Primary material was similar at 57% with a molybdenum concentrate grade of 47%.

13.4.2 Locked Cycle Tests Comparisons Using Brine & Tap Water

Six locked cycle tests were performed on three samples representing Supergene, Primary, and a 1:1 Supergene and Primary mineralized material blend. In one set of tests, brine water was used in the roughers and tap water in the cleaners. The other set of tests used tap water in both stages of the test. Results are summarized in Table 13-5.

Table 13-5: Locked Cycle Tests on Composites Comparing Use of Brine Water and Tap Water in Roughers

Concentrate Recovery Weight Concentrate Grade Sample Name % % Au, g/t Cu, % Mo, % Au Cu Mo

Brine Water in Roughers & Tap Water in Cleaners

Supergene

Bulk Rougher Concentrate 17.22 0.56 4.15 0.07 72.3 91.9 89.2

Bulk Cu - Mo Concentrate 2.34 3.69 29.2 0.27 63.4 86.5 54.7

Primary

Bulk Rougher Concentrate 13.44 0.82 3.65 0.13 72.0 94.4 92.4

Bulk Cu - Mo Concentrate 1.4 7.35 33.9 0.69 64.9 91.2 52.7

Supergene + Primary

Bulk Rougher Concentrate 12.4 0.83 5.02 0.13 73.9 93.2 90.6

Concentrate Recovery Weight Concentrate Grade Sample Name % % Au, g/t Cu, % Mo, % Au Cu Mo

Bulk Cu - Mo Concentrate 2.1 4.21 29.4 0.64 61.7 88.8 75.0

Tap Water in Roughers and Cleaners

Supergene

Bulk Rougher Concentrate 13.49 0.67 5.37 0.09 65.6 91.4 88.2

Bulk Cu - Mo Concentrate 1.82 4.13 38.6 0.41 51.5 86.4 53.9

Primary

Bulk Rougher Concentrate 10.26 1.1 5.25 0.18 74.0 95.3 92.9

Bulk Cu - Mo Concentrate 1.5 6.91 36.2 1.02 64.7 92.5 72.7

Supergene + Primary

Bulk Rougher Concentrate 11.8 0.9 5.53 0.14 78.5 93.0 91.0

Bulk Cu - Mo Concentrate 1.8 5.3 35.9 0.7 63.5 87.8 67.4

These tests show similar copper recovery whether using brine or tap water in the rougher circuit. One notable difference is that the mass pull in the bulk rougher circuit is higher when using brine by approximately 2% to 4% resulting in approximately 20% greater rougher concentrate production.

The results for molybdenum show a little more variance. For Supergene material, the molybdenum recovery is essentially the same independent of water quality. For Primary material the recovery for brine + tap water is significantly lower by about 20%, while for the blend of Supergene and Primary material, the molybdenum recovery is 8% higher in the case of brine and tap water.

In future study phases of the Project, further test work using brine water should be performed including molybdenum cleaner flotation, which was not included in these tests.

13.4.3 Variability Rougher Kinetic Tests

Rougher kinetic tests were performed on 15 Supergene samples and 25 Primary samples. The results are summarized in Table 13-6 and Table 13-7 and shown in Figure 13-1 to Figure 13-4.

Table 13-6: Variability Rougher Test – Supergene Molybdenum Copper Feed Copper Rougher Molybdenum Sample No. Rougher Grade % Recovery % Feed Grade % Recovery % 9320 1.27 96.8 0.018 79.3 9321 1.66 95.0 0.029 91.5 9322 2.12 97.1 0.026 84.4 9323 1.49 96.3 0.028 88.7

Molybdenum Copper Feed Copper Rougher Molybdenum Sample No. Rougher Grade % Recovery % Feed Grade % Recovery % 9324 1.91 97.5 0.020 87.1 9325 1.21 95.8 0.018 76.6 9326 0.92 98.5 0.010 91.2 9327 1.54 92.1 0.037 93.7 9328 1.64 93.3 0.019 92.5 9329 0.57 93.2 0.010 83.6 9330 0.59 94.4 0.007 90.9 9331 0.92 96.7 0.008 88.7 9332 0.51 83.4 0.004 80.1 9333 0.56 91.6 0.006 80.6 9334 1.52 95.8 0.012 91.5

Average 1.20 94.5 0.017 86.7

Figure 13-1: Copper Recovery vs Copper Feed Grade, Roughers – Supergene (Plenge, 2012)

Figure 13-2: Molybdenum Recovery vs Molybdenum Feed Grade, Roughers – Supergene (Plenge, 2012)

Table 13-7: Variability Rougher Tests – Primary Molybdenum Copper Feed Copper Rougher Molybdenum Sample No. Rougher Grade % Recovery % Feed Grade % Recovery % 9335 0.91 98.0 0.015 89.8 9336 0.47 97.7 0.013 90.6 9337 0.59 95.1 0.006 82.2 9338 0.44 96.2 0.018 93.0 9339 0.78 97.9 0.024 95.0 9340 0.60 97.9 0.021 94.7 9341 0.51 97.6 0.028 97.0 9342 0.63 99.0 0.011 94.9 9343 0.35 97.2 0.032 90.0 9344 0.50 98.2 0.023 94.1 9345 0.46 97.6 0.010 95.7 9346 0.48 98.8 0.013 97.3 9347 0.51 97.9 0.010 87.1 9348 0.62 98.8 0.013 88.9 9349 0.64 97.5 0.013 84.4

Molybdenum Copper Feed Copper Rougher Molybdenum Sample No. Rougher Grade % Recovery % Feed Grade % Recovery % 9350 0.66 99.0 0.009 87.5 9351 0.43 95.2 0.013 96.9 9352 0.33 94.6 0.010 96.1 9353 0.42 94.9 0.012 96.6 9354 0.55 97.2 0.020 98.1 9355 0.32 95.9 0.017 95.1 9356 0.37 97.7 0.013 95.1 9357 0.31 97.6 0.013 96.2 9358 0.38 97.8 0.011 85.7 9359 0.38 97.7 0.009 84.0

Average 0.50 97.3 0.015 92.2

Figure 13-3: Copper Recovery vs Copper Feed Grade, Roughers – Primary (Plenge, 2012)

Figure 13-4: Molybdenum Recovery vs Molybdenum Feed Grade, Roughers – Primary (Plenge, 2012)

The average copper recovery for the rougher variability tests for Supergene material was 94.5% and for Primary material was 97.3%. The average molybdenum recovery for the rougher variability tests for Supergene material was 86.7% and for Primary material was 92.2%. All of these tests used tap water and the results show good recovery for both copper and molybdenum in the rougher stage of flotation.

13.4.4 Variability Cleaner Tests

Cleaner variability tests were performed on 15 Supergene samples and 25 Primary samples. The results are summarized in Table 13-8 and Figures 13-5 and 13-6 (Supergene) and Table 13-9 and Figures 13-7 and 13-8 (Primary).

Table 13-8: Variability Cleaner Tests – Supergene Molybdenum Copper Feed Grade Copper Cleaner Molybdenum Sample No. Cleaner % Recovery % Feed Grade % Recovery % 9320 1.23 94.3 0.018 72.7

9321 1.58 92.1 0.029 90.3

9322 2.02 95.5 0.025 81.5

9323 1.37 94.0 0.027 87.0

9324 1.98 95.8 0.020 79.0

9325 1.17 93.7 0.017 53.4

Molybdenum Copper Feed Grade Copper Cleaner Molybdenum Sample No. Cleaner % Recovery % Feed Grade % Recovery % 9326 0.89 97.2 0.009 84.6

9327 1.60 88.6 0.036 90.5

9328 1.67 89.7 0.019 89.5

9329 0.55 88.8 0.009 80.4

9330 0.61 83.7 0.008 57.4

9331 0.92 94.1 0.009 82.1

9332 0.52 75.2 0.004 66.3

9333 0.55 86.1 0.006 72.3

9334 1.59 93.0 0.013 83.0

Average 1.2 90.8 0.017 78.0

Figure 13-5: Copper Recovery vs Copper Feed Grade, Cleaners – Supergene (Plenge, 2012)

Figure 13-6: Molybdenum Recovery vs Molybdenum Feed Grade, Cleaners – Supergene (Plenge, 2012)

Table 13-9: Variability Cleaner Tests – Primary Molybdenum Copper Feed Copper Cleaner Molybdenum Sample No. Cleaner Grade % Recovery % Feed Grade % Recovery % 9335 0.89 95.7 0.016 81.9

9336 0.45 87.3 0.015 89.8

9337 0.56 90.3 0.006 76.9

9338 0.43 93.0 0.018 87.0

9339 0.74 96.4 0.026 91.5

9340 0.67 96.3 0.020 93.2

9341 0.53 96.9 0.027 95.5

9342 0.60 98.3 0.011 92.3

9343 0.34 96.1 0.037 95.8

9344 0.48 96.3 0.024 89.8

9345 0.45 95.4 0.010 93.2

9346 0.46 97.7 0.013 93.1

9347 0.49 95.4 0.009 92.0

9348 0.61 97.8 0.013 96.1

9349 0.64 95.9 0.012 96.5

9350 0.64 98.1 0.009 95.7

9351 0.42 89.8 0.014 75.5

9352 0.32 91.2 0.010 94.8

9353 0.40 91.3 0.012 95.5

9354 0.54 94.8 0.021 97.4

9355 0.32 89.5 0.018 82.1

9356 0.35 95.4 0.014 86.8

9357 0.31 95.2 0.015 90.1

9358 0.40 96.9 0.013 94.6

9359 96.5 0.009 90.4

Average 94.6 0.02 90.71

Figure 13-7: Copper Recovery vs Copper Feed Grade, Cleaners – Primary (Plenge, 2012)

Figure 13-8: Molybdenum Recovery vs Molybdenum Feed Grade, Cleaners – Primary (Plenge, 2012)

The average copper recovery for the cleaner variability tests for Supergene material was 90.8% and for Primary material was 94.6%. The average molybdenum recovery for the cleaner variability tests for Supergene material was 78% and for Primary material was 90.7%. All of these tests used tap water and the results show good recovery for both copper and molybdenum in the bulk cleaner stage of flotation.

The equations from the cleaner copper recovery versus copper feed grade curves were used to determine copper recovery in the annual concentrate production schedule.

13.4.5 Batch Cleaner Tests Varying Water Combination (Tap and Pit Water)

A series of four batch cleaner tests were performed on two samples representing Supergene and Primary materials to compare the effects of a variation in water quality. The results are summarized in Table 13-10.

Table 13-10: Batch Cleaner Test Results at Different Water Quality Blends Copper Copper Copper Molybdenum Molybdenum Water Molybdenum Feed Rougher Cleaner Rougher Cleaner Type Feed Grade % Grade % Recovery % Recovery % Recovery % Recovery % Supergene

100% Tap 0.8 91.2 75.3 0.012 90.2 34.3 25% Pit + 0.83 91.2 53.5 0.014 86.1 20.8 75% Tap 50% Pit + 0.81 91.5 61.5 0.014 86.7 33.8 50% Tap 75% Pit + 0.8 91.1 63.9 0.014 84.3 42.9 50% Tap Average 0.81 91.3 63.6 0.014 86.8 79.3

Copper Copper Copper Molybdenum Molybdenum Water Molybdenum Feed Rougher Cleaner Rougher Cleaner Type Feed Grade % Grade % Recovery % Recovery % Recovery % Recovery % Primary

100% Tap 0.57 95.5 76.8 0.021 92.4 29.8 25% Pit + 0.53 94.1 47.2 0.020 87.4 17.5 75% Tap 50% Pit + 0.54 94.4 56.7 0.019 89.2 28.4 50% Tap 75% Pit + 0.56 94.4 74.3 0.019 92.0 53.9 50% Tap Average 0.54 94.6 63.8 0.02 90.3 32.4

The copper rougher recovery for Supergene material averaged 91.3% and there was very little difference in rougher recovery at the different water compositions. The copper rougher recovery for Primary material averaged 94.6% and there was little difference in rougher recovery at the different water compositions; however, 100% tap water provided a slight improvement (1.1%) in copper recovery over other blends of water types. This confirms that the use of brine water (or pit water) in the rougher flotation stage does not have a significant impact on rougher flotation as was shown in the locked cycle tests previously discussed.

The molybdenum rougher recovery using tap water appears to be a little better than using pit water for the Supergene material but there was not a noticeable difference with Primary material.

Although rougher flotation has shown little affect due to the water type, the cleaner flotation recovery results were impacted by water type.

The parameters of the brine water used in the test work are summarized in Table 13-11.

Table 13-11: Summary of Brine Water Analyses Salar de Plumas Parameter Unit Pit Water Arizaro Verde pH pH 7.1 8.3 7.05 Conductivity uS/cm >200 000 17 420 241 700 TDS mg/l 255 500 10 700 317 596 Alkalinity mg/l 49 129 - Bicarbonate mg/l 59 158 - Calcium mg/l 2 510 169 1558 Magnesium mg/l 1 350 82 706 Chloride mg/l 153 000 6 120 184 600 Sulfate mg/l 3 900 310 Nitrate mg/l 600 30 Sodium mg/l 84 400 4 120 70 277 Potassium mg/l 3 030 77

Ausenco recommends that more test work, including locked cycle tests using site brine water be conducted to investigate further the use of brine water and site pit water for all flotation processes. This could have a significant impact on capital by reducing the desalination water treatment requirements.

13.5 Sedimentation and Filtration Test Work

Sedimentation test work of copper concentrates and tailings from Supergene and Primary material locked cycle tests were conducted; the results are summarized in Table 13-12 and Table 13-13, respectively.

Table 13-12: Copper Concentrate Sedimentation Results

Settling Requirement Material Type % Underflow m2/(t/d)

Supergene S1 63 0.06

Supergene S2 63 0.04

Primary P1 51 0.10

Primary P2 54 0.06

Average 58 0.07

Table 13-13: Tailings Sedimentation Results

Settling Requirement Material Type % Underflow 2 m /(t/d)

Supergene S1 52 0.08

Supergene S2 54 0.05

Primary P1 50 0.06

Primary P2 50 0.08

Average 52 0.07

Filtration test work of copper concentrates from Supergene and Primary material locked cycle tests were conducted; the results are summarized in Table 13-14.

Table 13-14: Filtration Results for Locked Cycle Test Copper Concentrates

Filtering Rate Material Type % Solids Feed % Moisture m2/(t/h)

Supergene S1 50 7 1.12

60 8 0.95

Supergene S2 50 5 1.12

Filtering Rate Material Type % Solids Feed % Moisture m2/(t/h)

60 5 0.81

Primary P1 51 7 1.11

62 7 0.95

Primary P2 51 8 1.68

62 11 1.35

Average 56 7 1.14

75th Percentile 1.18

14 Mineral Resource Estimates

14.1 Introduction

The mineral resource estimate was prepared under the direction of Robert Sim and assisted by Bruce Davis. Estimates are generated, from three–dimensional (3D) block models bases on geostatistical applications using commercial mine planning software (MineSight® v7.50). The Project limits are based in the UTM coordinate system using a nominal block size of 25 x 25 x 15 m (L x W x H). The majority of drill holes in the main deposit area are vertically oriented with holes spaced on a nominal 150 m grid pattern. At the northern end of the deposit, the final 500 m has been tested with holes that are consistently inclined -70° east.

14.2 Geologic Model, Domains, and Coding

The copper, gold, and molybdenum mineralization on the Taca Taca property is interpreted to be the result of deep-seated rhyodacitic porphyry intrusions. The majority of the rocks that host the Taca Taca deposit are granodioritic in composition, plus minor dacite, diabase, and rhyolite dykes. The deposit is overlain by a leach cap that ranges from 300 m thick in the south to about 150 m in the north. This is underlain by a locally irregular Supergene Zone which varies in thickness from non-existent to greater than 300 m in some parts of the deposit. In some areas, supergene-type mineralization is locally present at depths of greater than 700 m below surface. This variability in supergene thickness is attributed to deep-seated enrichment along fault structures. The Supergene Zone contains varying amounts of chalcocite and covellite. Beneath the Supergene Zone is the Primary Zone domain comprised of varying amounts of pyrite, chalcopyrite, bornite, and minor molybdenite.

A sub-vertical fault is interpreted from drilling on the western side of the deposit area. Generally oriented at 345°, this structure shows variable vertical displacement with approximately 160 m of apparent displacement in the south, but little to no movement in the north. Copper mineralization is present on both sides of the structure, but there appears to be some post-depositional movement along this fault.

The majority of the deposit is hosted within rocks of granitic or granodioritic composition. A series of sub-vertical late/post-mineral dykes occur to the east and southeast of the main deposit area. These dykes have been interpreted based on a combination of drilling results and surface mapping. Dykes of rhyodacitic composition are present primarily in the southeastern part of the deposit and tend to be post-mineral in nature. Several rhyolite dykes are interpreted in the eastern part of the deposit. These rhyolite dykes show mineral trends which suggest they were emplaced prior to or during the mineralizing event.

Immediately east of the deposit area is a salt brine salar in which a series of drill holes were collared from the surface of the salt crust. The base of the salar was interpreted from the results of this drilling and was included as overburden in the model.

The geologic interpretation of the base of Overburden Zone, the Leach Cap Zone, and the Supergene Zone was generated using drilling information. Three-dimensional, wireframe shape domains were generated and represent the extents of these various mineral zones (MinZone). The MinZone domains are summarized in Table 14-1 and shown in Figure 14-1.

Table 14-1: MinZone Domains and Coding

Zone Code Domain Comment Number

Overburden (OVB) 1 Surface soil and gravel, plus the salar.

Leach (LX) 2 Average 300 mV near-surface zone leached of copper.

Supergene (SS) 3 Supergene zone of enriched copper.

Primary (PR) 4 Zone of primary sulphide mineralization.

5 Late/post mineral dykes. Rhyodacite Dykes (DK) Tend to be weakly mineralized or unmineralized.

6 Pre/syn mineral dykes. Tend to be mineralized similar Rhyolite Dykes (RDK) to surrounding host rocks.

Figure 14-1: Isometric Views of MinZone Domains (Sim, 2013)

14.3 Available Data

Delineation and exploration drilling have been ongoing at the Project for several years. Lumina Copper delivered the final database on October 30, 2012 and it included information from 440 drill holes, with a cumulative length of 163,537 m. Drill holes occur over an area that measures 5 km x 5 km, with a few drill holes dating back to the 1970s.

Comments relating to the sample database include the following:

 Most of the holes drilled by Corriente Resources in the 1990s were designed to target near- surface veins and mantos, and, as a result, these do not provide information relating to the deeper-seated porphyry mineralization.  There are 52 holes in the database that do not have any associated sample results. Seven of these were drilled by Lumina Copper for geotechnical purposes and none of those have been sampled or analyzed. The remaining 45 are older drill holes that were either never sampled or the assay data is missing.  Lumina Copper drilled a total of 97 RC holes. Parts of the Leach Zone have been delineated using RC drilling. RC holes have been used primarily in testing the northern and northwestern mineralized areas where mineralization tends to occur to depths of only 300 m. Comparisons between samples from DD holes and RC holes show local variability, but good overall correlation for copper and molybdenum. Gold grades tend to be slightly higher in DD holes than RC holes, but the differences are not considered significant.

The mineral resource estimate is based on a total of 147,449 m of drilling in 310 drill holes that are proximal to the potentially economic mineralization. The associated assay data includes: BHP, 28 holes (9,893 m); Corriente Resources, 18 holes (1,454 m); Rio Tinto, 15 holes (7,608 m); and, Lumina Copper, 249 holes (128,494 m). The distribution of drilling is shown in Figure 14-2.

Figure 14-2: Distribution of Copper Grades in Drill Holes (Sim, 2013)

The previous mineral resource estimate was generated in April 2012. Figure 14-3 shows the distribution of data that was available for the April 2012 resource estimate (shown in blue) compared to the additional sample data available for use in the current resource estimate (shown in red). Significant additional drilling has been completed north and northwest of the main deposit area. Deep drilling has begun to delineate what appears to be the eastern limit of the deposit. Additional holes in the central and southern parts of the deposit have significantly improved the understanding of the deep-seated mineralization in these areas.

Figure 14-3: Distribution of Additional Sample Data since the April 2012 Resource Estimate (Sim, 2013)

The majority of the drill holes that test the deeper porphyry zone are variably spaced between 100 m and 200 m intervals. Initial drilling on the deposit consisted of holes that ranged from 400-600 m long. In 2008, Rio Tinto drilled hole TTBJ0002 to a final depth of 1,153.3 m; this remains the deepest hole on the property. Lumina Copper has intersected appreciable copper mineralization in numerous drill holes pushed to greater than 900 m below surface. The majority of drill holes that test the porphyry are vertically oriented with a few holes drilled at inclinations as shallow as -60°. Recent drill holes located in the northern part of the deposit area are inclined at -70° in an eastern direction to further evaluate a pervasive sub-vertical, north-south series of structures interpreted to be present.

Downhole survey data only exists for holes drilled since 2008; data includes holes drilled by Rio Tinto and all holes drilled by Lumina Copper. Available survey data indicates that drill hole deviations are variable, but can be significant in some cases, especially in the upper leached rocks. It is safe to assume that similar degrees of deviation have occurred in previous drill holes. This is not considered overly significant in respect to a global resource estimate using the current drill hole spacing; however, this fact should be considered as the Project evolves. Some drill holes may have to be re-drilled to confirm the exact location of the contained sample data.

A total of 149,293 m of the drilling was sampled and analyzed for copper and, in most cases, a suite of other minor elements. There are 74,400 individual samples in the database that were tested for copper content, with an average sample length of 2 m. A total of 10 elements were incorporated into the resource model, many of which have no apparent economic influence on the Project. This report describes only the estimation of copper (Cu%), gold (Au g/t), and molybdenum (Mo%) in the resource model. A basic statistical summary of all of the primary sample data is listed in Table 14-2.

Table 14-2: Basic Summary of All Sample Data

Number of Total Standard Element Minimum Maximum Mean Samples Length (m) Deviation Copper (Cu%) 74,400 149,280 0 14.75 0.195 0.337 Molybdenum (Mo%) 73,872 148,229 0 1.000 0.009 0.015 Gold (Au g/t) 73,962 148,410 0 17.490 0.077 0.170

As stated previously, the Taca Taca drilling database includes holes that test for satellite exploration targets and near surface veins and mantos. Of the 440 drill holes in the database, 310 holes are within the immediate vicinity; these 310 holes have potential influence on the mineral resource estimate. A basic statistical summary of these proximal drill holes is listed in Table 14-3.

Table 14-3: Basic Summary of Sample Data Proximal to the Resource Model

Number of Total Standard Element Minimum Maximum Mean Samples Length (m) Deviation Copper (Cu%) 71,144 143,740 0 14.75 0.201 0.341 Molybdenum (Mo%) 70,591 143,625 0 1.000 0.010 0.015 Gold (Au g/t) 70,714 143,713 0 14.300 0.078 0.149

The geologic information is derived primarily from observations during logging, and includes lithology, alteration facies, and mineral zonation type.

14.4 Compositing

Drill hole samples are composited to standardize the database for further statistical evaluation. This step eliminates any effect sample lengths may have on the estimate.

To retain the original characteristics of the underlying data, a composite length that reflects the average original sample length is selected. The generation of longer composites results in some degree of smoothing which could mask some of the features of the data. Sample intervals range from 0.35-6 m long, with an average of 2 m. As a result, a standard 2 m composite length was generated for statistical evaluation and for use in grade estimations in the block model.

Drill hole composites are length-weighted and were generated “down-the-hole”, this means composites begin at the top of each hole and are generated at 2 m intervals down the length of the hole. The contacts of the MinZone domains, listed in Table 14-1, were honoured during compositing of drill holes. Several holes were randomly selected and the composited values were checked for accuracy. No errors were found. Logged lithology, alteration, and MinZone data were assigned to composited intervals on a majority basis to allow for statistical analysis of these variables.

14.5 Exploratory Data Analysis

Exploratory data analysis (EDA) involves statistically summarizing the database to quantify the characteristics of the data. One of the main purposes of EDA is to determine if there is any evidence of spatial distinctions in grade; this would require the separation and isolation of domains during interpolation. The application of separate domains prevents unwanted mixing of data during interpolation; this will result in a grade model that better reflects the unique properties of the deposit. However, applying domain boundaries in areas where the data is not statistically unique may impose a bias in the distribution of grades in the model.

A domain boundary segregating the data during interpolation is typically applied if the average grade in one domain is significantly different from that of another domain. A boundary may also be applied where there is a significant change in the grade distribution across the contact.

14.5.1 Basic Statistics by Domain

The basic statistics for the distribution of copper, gold, and molybdenum were generated by lithology type, alteration facies type, and by interpreted MinZone type. The drill core logs have identified 13 different rock types; 86% are comprised of granite. Although some other rock types may suggest distinct metal properties, they tend to be so rare that it is not practical to use them for resource estimation purposes. The late stage rhyodacite dykes show relatively low copper grades compared to the rhyolite dykes that average 0.30% contained copper.

Comparison of alteration types show propylitic alteration tends to be lower in copper, gold, and molybdenum. This alteration type is present in only a few drill holes that tend to occur around the perimeter of the main deposit area.

The interpreted MinZone domains show copper grades to be highest in the Supergene Zone and lowest in the Leach Zone domain (Figure 14-4). The difference between the two types of dykes is quite evident: the rhyolite dyke grades are similar to the Supergene and Primary Zones, and the rhyodacite dykes show relatively low copper content. There are no significant differences in gold and molybdenum content in relation to the MinZone types.

Figure 14-4: Box Plot Copper by MinZone Domain

14.5.2 Contact Profiles

The nature of grade trends between two domains is evaluated using the contact profile; this profile graphically displays the average grades at increasing distances from the contact boundary. A contact profile that shows a marked difference in grade across a domain boundary is an indication that the two data sets should be isolated during interpolation. Conversely, if there is a more gradual change in grade across a contact, the introduction of a hard boundary (in other words, segregation during interpolation) may result in much different trends in the grade model. In this case, the change in grade between domains in the model is often more abrupt than the trends seen in the raw data. Finally, a flat contact profile indicates no grade changes across the boundary. In the case of a flat profile, hard or soft domain boundaries will produce similar results in the model.

Contact profiles were generated for copper grades between the interpreted MinZone domains. Distinct changes in grade are evident between all MinZones at the domain boundaries. The profiles for the Leach Zone-Supergene Zone (LX-SS) and Supergene Zone-Primary Zone (SS-PR) contacts are shown in Figure 14-5 and Figure 14-6. The results suggest that these boundaries were honoured during the generation of the copper resource model. The rhyolite dyke does not show any significant grade change with the surrounding host rocks.

The nature of molybdenum and gold across domain boundaries was also reviewed and no significant trends or changes were identified.

Figure 14-5: Contact Profiles Copper between LX and SS MinZones

Figure 14-6: Contact Profiles Copper between SS and PR MinZones

14.5.3 Conclusions and Modeling Implications

Boxplots show that differences exist in copper grades between the various MinZone domains, and this is further supported by the contact profile analysis. Late-stage rhyodacite dykes, containing lower copper grades, were interpreted using surface mapping and drill core logging information. All of these domains should be recognized and honoured with hard boundary conditions during block estimations of copper content in the resource model.

The rhyolite dykes show similar mineral content as their surrounding host rocks and, as a result, this unit does not require segregation during block grade interpolation. The areas identified as rhyolite dykes were included within the corresponding Leach, Supergene, or Primary MinZone domains.

There do not appear to be any significant differences in the distribution of gold or molybdenum in relation to any of the geologic domains present in the deposit. As a result, there are no restrictions during block grade interpolations of these elements.

The interpolation domains are summarized in Table 14-4.

Table 14-4: Summary of Interpolation Domains

Zone Code Domain Comments Number 1 OVB (Overburden Zone) – no estimates conducted 2 LX (Leach Zone) -- hard boundary Copper 3 SS (Supergene Zone) – hard boundary 4 PR (Primary Zone) – hard boundary 5 Rhyodacite Dyke – hard boundary Molybdenum 2-5 Combined all domains (excl. OVB) Gold 2-5 Combined all domains (excl. OVB)

14.6 Bulk Density Data

During the 2011-2012 drilling program, samples were sent to ALS Chemex in Lima, Peru for bulk density measurements. To date, 662 samples were selected on approximately 100 m intervals from drill holes spread throughout the deposit area.

Bulk density was measured using the wet method: paraffin wax-coated pieces of core, averaging 10 cm to 15 cm in length, are weighed in air and then again while submerged in water. The following formula for specific gravity was used:

/ /

Note: A = weight of sample in air, B = weight of waxed sample in air, C = weight of waxed sample in water, and D = density of wax.

Although the density of data is considered insufficient to estimate the actual specific gravity (SG) values in model blocks, the data provides a relatively sound basis to determine average ranges.

Table 14-5 lists the average bulk density values assigned to the model block by MinZone domain. These values are similar to the previous densities used in 2008 by Rio Tinto and are considered appropriate based on the rock types and mineralization present in the deposit.

Table 14-5: Summary of Bulk Density by MinZone Domain

Domain SG (t/m3)

Overburden 1.90

LX 2.50

SS 2.62

PR 2.65

DYKE 2.60

14.7 Evaluation of Outlier Grades

Histograms and probability plots were reviewed to identify the existence of anomalous outlier grades in the composited sample database. Following a review of the physical location of these potentially anomalous samples, it was decided that copper and molybdenum data could be controlled through an outlier limitation; this limits their effective distance to a maximum of 50 m during interpolation. Gold data is controlled through a combination of traditional top-cutting of composited samples > 3 g/t Au and an outlier limitation of samples > 1.5 g/t Au to a maximum influence distance of 50 m during block grade interpolation. These parameters are summarized in Table 14-6.

Table 14-6: Summary of Outlier Limits

Number of % Metal Loss Element/Domain Threshold Grade Composites Affected in Model

Copper – LX 1.5% 20 -3.7%

SS 3.5% 44 -2.3%

PR 2.5% 23 -1.2%

DYKE 0.4% 38 -2.4%

Molybdenum 0.25% 19 -1.4%

Top cut to 3 g/t Au, 19 comps >3 g/t Au Gold >1.5 g/t Au limited to 50 m -5.9% 55 comps >1.5 g/t Au influence

Note: 2 m composited drill hole sample data. Outlier data limited to maximum distance of 50 m during interpolation except for gold as noted.

The proportion of copper metal lost in the LX (Leach Zone) domain appears significant, but this domain hosts very low copper grades and, as a result, the potential economic contribution is relatively small. Overall, the combined copper model is reduced by 1.4% as a result of the applied

outlier limitations during block grade interpolations. The gold database shows a relatively skewed distribution, resulting in a moderate reduction in contained metal in the resource model. This suggests that additional drilling is required to fully understand the nature of gold in the deposit.

14.8 Variography

The degree of spatial variability in a mineral deposit depends on both the distance and direction between points of comparison. Typically, the variability between samples is proportionate to the distance between samples. If the degree of variability is related to the direction of comparison, then the deposit is said to exhibit anisotropic tendencies which can be summarized with the search ellipse. The semi-variogram is a common function used to measure the spatial variability within a deposit.

The components of the variogram include the nugget, the sill, and the range. Often samples compared over very short distances (including samples from the same location) show some degree of variability. As a result, the curve of the variogram often begins at some point on the y-axis above the origin; this point is called the nugget. The nugget is a measure of not only the natural variability of the data over very short distances, but also a measure of the variability which can be introduced due to errors during sample collection, preparation, and assaying.

The amount of variability between samples typically increases as the distance between the samples increases. Eventually, the degree of variability between samples reaches a constant or maximum value; this is called the sill. The distance between samples at which this occurs is called the range.

The spatial evaluation of the data in this report was conducted using a correlogram instead of the traditional variogram. The correlogram is normalized to the variance of the data and is less sensitive to outlier values; this generally gives cleaner results.

Variograms were generated using the commercial software package Sage 2001© developed by Isaacs & Co. Multidirectional variograms were generated for copper, gold, and molybdenum in the various domains; the results are summarized in Table 14-7, Table 14-8, and Table 14-9.

Table 14-7: Variogram Parameters - Copper

1st Structure 2nd Structure

Range Range Domain Nugget Sill 1 Sill 2 Azimuth Dip Azimuth Dip (m) (m)

0.335 0.529 0.136 148 193 52 963 186 5

Leach 34 71 22 480 97 -2 Spherical 12 328 29 182 212 -85

0.250 0.316 0.434 47 269 18 1191 82 6

Supergene 27 356 -10 457 352 -4 Spherical 16 59 70 341 293 83

0.253 0.231 0.516 105 88 58 964 222 -20 Primary Spherical 43 351 4 677 186 66

1st Structure 2nd Structure

19 258 31 398 127 -13

0.256 0.056 0.688 64 51 13 2045 8 20

Dyke 59 314 26 1129 92 -15 Spherical 38 346 -61 131 148 64

Note: Correlograms conducted on 2 m composited drill hole sample data.

Table 14-8: Variogram Parameters - Molybdenum

1st Structure 2nd Structure

Zone Nugget S1 S2 Range (m) AZ Dip Range (m) AZ Dip

0.300 0.385 0.315 72 36 7 1671 25 -25 LX+SS+PR+ 37 289 65 1128 104 23 DK Spherical 4 733 338 55

Note: Correlograms conducted on 2 m composited drill hole sample data.

Table 14-9: Variogram Parameters - Gold

1st Structure 2nd Structure

Zone Nugget S1 S2 Range (m) AZ Dip Range (m) AZ Dip

0.300 0.385 0.315 72 36 7 1671 25 -25 LX+SS+PR+ 37 289 65 1128 104 23 DK Spherical 4 733 338 55

Note: Correlograms conducted on 2 m composited drill hole sample data.

14.9 Model Setup and Limits

The block model was initialized in MineSight® and the dimensions are defined in Table 14-10. The selection of a nominal block size measuring 25 x 25 x 15 mV is considered appropriate with respect to the current drill hole spacing and the selective mining unit (SMU) size, which is typical of an operation of this type and scale. The model has not been rotated.

Table 14-10: Block Model Limits

Block Size Number of Direction Minimum Maximum (m) Blocks East 2626000 2630600 25 184 North 7281200 7285800 25 184 Elevation 2450 3770 15 88

Blocks in the model were coded on a majority basis with the MinZone code domains. During this stage, blocks along a domain boundary are coded if > 50% of the block occurs within the boundaries of that domain.

The proportion of blocks that occur below the topographic surface is also calculated and stored within the model as individual percentage items. These values are used as a weighting factor to determine the in-situ resources for the deposit.

14.10 Interpolation Parameters

The block model grades for copper, gold, and molybdenum were estimated using Ordinary Kriging (OK). The results of the OK estimation were evaluated using a series of validation approaches described in Item 14.11. The interpolation parameters were adjusted until the appropriate results were achieved.

In most cases, the Taca Taca OK models were generated with a relatively limited number samples. This approach reduces the amount of smoothing or averaging in the model and, while there may be some uncertainty on a localized scale, this approach produces reliable estimations of the recoverable grade and tonnage for the overall deposit.

All grade estimations use length-weighted composite drill hole sample data. The interpolation parameters are summarized in Table 14-11.

The sub-vertical fault on the western edge of the deposit shows approximately 160 m of vertical displacement of the MinZone domain contacts. As a result, this fault is treated as a hard boundary (in other words, sample data is not mixed across this fault boundary) during interpolation of grades in the model.

Estimation of copper within the Supergene Zone domain was done in two passes. The first pass estimates blocks in the stratabound-type of supergene mineralization. The second pass reduces the mixing of data between the deep-seated, structure-type supergene mineralization by applying an oriented ellipse with shorter search ranges.

Table 14-11: Interpolation Parameters Search Ellipse Number of Composites Element/Domain Range (m) Other X Y Z Min/block Max/block Max/hole Copper – LX 750 750 100 5 60 15 1 DH per quadrant SS - strata 750 750 100 5 80 20 1 DH per quadrant SS - structure 250 250 100 5 36 12 PR 750 750 100 5 48 12 1 DH per quadrant DYKE 750 750 100 5 24 8 1 DH per quadrant Molybdenum 750 750 100 5 120 20 1 DH per quadrant Gold 750 750 100 5 160 20 1 DH per octant

Note: Search ellipse Z-axis is vertical in all estimates except SS-structure. SS-structure ellipse oriented north-south dipping - 70 degrees to the west.

14.11 Validation

The results of the modelling process were validated through several methods including: a thorough visual review of the model grades in relation to the underlying drill hole sample grades; comparisons with the change of support model; comparisons with other estimation methods; and, grade distribution comparisons using swath plots.

14.11.1 Visual Inspection

A detailed visual inspection of the block model was conducted in both cross section and plan to ensure the desired results following interpolation. This included confirmation of the proper coding of blocks within the respective domains and below the topographic surface. The distribution of block grades was also compared relative to the drill hole samples to ensure the proper representation in the model.

Overall, all grade models showed the desired degree of correlation with the underlying sample data. The application of the deep-seated, supergene-type structures has restricted the impact on the high-grade intersections encountered at depth. At this stage of the Project, the nature of these structures is not well known, but this approach retains control of the model in these areas.

14.11.2 Model Checks for Change of Support

The relative degree of smoothing in the block model estimates were evaluated using the Discrete Gaussian or Hermitian Polynomial Change of Support method (Journel and Huijbregts, Mining Geostatistics, 1978). With this method, the distribution of the hypothetical block grades is directly compared to the estimated OK model using pseudo-grade/tonnage curves. Adjustments are made to the block model interpolation parameters until an acceptable match is made with the Herco (Hermitian correction) distribution. In general, the estimated model should be slightly higher in tonnage and slightly lower in grade when compared to the Herco distribution at the projected cut-off grade. These differences account for selectivity and other potential material handling issues that commonly occur during mining.

The Herco distribution is derived from the declustered composite grades which were adjusted to account for the change in support moving from smaller drill hole composite samples to the larger

blocks in the model. The transformation results in a less skewed distribution, but retains the same mean as the original declustered samples.

Herco and model grade-tonnage plots were generated for the distribution of copper in the LX (Leach), SS (Supergene), and PR (Primary) domains, and for molybdenum and gold in the combined domains. Examples for copper in the SS and PR domains, shown in Figure 14-7 and Figure 14-8, exhibit the desired degree of correlation between models. Even though in some of the examples the model grade curves fall above the Herco grade curves, it should be noted that the change of support model is a theoretical tool intended to direct model estimation. There is uncertainty associated with the change of support model, and its results should not be viewed as final and correct values. In cases where the model grades are greater than the change of support grades, the model is relatively insensitive to changes in modelling parameters. Any extraordinary measures to change the grade curves are not warranted.

Figure 14-7: Herco Copper in SS Zone

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Figure 14-8: Herco Copper in PR Zone

14.11.3 Comparison of Interpolation Methods

For comparison purposes, additional models for copper, molybdenum, and gold were generated using both the inverse distance weighted (IDW) and nearest neighbour (NN) interpolation methods. (The NN model was created using data composited to 15 m intervals). The results of these models are compared to the OK models at various cut-off grades in the grade/tonnage graphs shown in Figure 14-9, Figure 14-10, and Figure 14-11. Overall, there is very good correlation between these models. It should be noted that the NN model reflects a different level of selectivity compared to the IDW or OK models. Therefore, it should follow the general trend of the other curves, but no error is suggested if the NN does not closely coincide with the other two models. Reproduction of the model using different methods tends to increase the level of confidence in the overall resource.

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Figure 14-9: GT Comparison of OK, IDW, and NN Copper Models

Figure 14-10: GT Comparison of OK, IDW, and NN Molybdenum Models

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Figure 14-11: GT Comparison of OK, IDW, and NN Gold Models

14.11.4 Swath Plots (Drift Analysis)

A swath plot is a graphical display of the grade distribution derived from a series of bands, or swaths, generated in several directions throughout the deposit. Grade variations from the OK model are compared using the swath plot to show the distribution derived from the declustered NN grade model.

On a local scale, the NN model does not provide reliable estimations of grade, but, on a much larger scale, it represents an unbiased estimation of the grade distribution based on the underlying data. Therefore, if the OK model is unbiased, the grade trends may show local fluctuations on a swath plot, but the overall trend should be similar to the NN distribution of grade.

Swath plots were generated in three orthogonal directions to compare the OK and NN distributions of copper, gold, and molybdenum in the deposit. Examples from the copper model are shown in Figure 14-12, Figure 14-13, and Figure 14-14.

There is good correspondence between the models in most areas of the deposit. The degree of smoothing in the OK model is evident in some of the peaks and valleys shown in the swath plots. Deviations tend to occur on the flanks of the deposit, or at depth, where the density of drilling often decreases.

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Figure 14-12: Copper Model Swath Plot by Easting

Figure 14-13: Copper Model Swath Plot by Northing

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Figure 14-14: Copper Model Swath Plot by Elevation

14.12 Resource Classification

Mineral resources for the Project were classified according to the CIM Definition Standards for Mineral Resources and Mineral Reserves (November 2010). The classification parameters are defined relative to the distance between sample data and are intended to encompass zones of reasonably continuous mineralization.

Mineral resources are limited to the porphyry-style mineralization and exclude any peripheral manto or vein-type mineralization that was intersected in some of the short drill holes that flank the deposit. Although these peripheral veins and mantos represent only a minor fraction of the potential resource, the current drilling density is insufficient to properly evaluate the nature of these occurrences. The extent of the porphyry mineralization is shown in plan in Figure 14-15. Copper resources are also limited to resource blocks that occur within the Supergene and Primary Zone domains. Potentially leachable gold resources are also present in the Leach Zone domain and are included as additional Inferred resources based on gold cut-off grades. The results of the copper grade and indicator variograms, together with experience gained through visual interpretation of the drilling results to-date, form the basis of the classification criteria for mineral resources at Taca Taca. Inferred resources include blocks within a maximum distance of 250 m from a drill hole. Indicated resources are defined by areas delineated by continuous drilling with a maximum grid spacing of 150 m. Portions of the deposit that meet these initial criteria were further reviewed to ensure that they exhibit the appropriate continuity to support the level of confidence required for resources in the Indicated category. Ultimately, a manually generated volume was interpreted and used to code blocks in the model that meet the criteria for Indicated-class resources. There are no resources that meet the criteria necessary to be included in the measured category.

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Figure 14-15: Plan Map Showing Limit of Porphyry Type Mineralization (Sim, 2013)

14.13 Mineral Resources

Although the Taca Taca deposit is primarily a copper deposit, gold and molybdenum are sufficiently present to significantly contribute to the potential value of the deposit. As a result, the mineral resources are estimated based on a copper equivalent (CuEq) cut-off grade. The CuEq grades were calculated in each block in the model based on the following equation and assumptions: % % / ∗ 0.583 % ∗ 6 CuEq% Assumptions: Metal prices: $2.00/lb Cu, $800/oz Au, and $12.00/lb Mo. Mining and metallurgical recoveries are assumed to be 100%.

The estimated Taca Taca mineral resources are summarized at various cut-off grades for comparison purposes in Table 14-12 and Table 14-13; the base case cut-off grades for copper resources of 0.3% CuEq and for gold resources of 0.2 g/t Au are highlighted. The base case cut-off grades are based on assumptions derived from operations with similar characteristics, scale, and location. The base case cut-off grade for the copper mineral resource estimate has been decreased from 0.4% copper equivalent, used in the previous resource estimate, to 0.3% copper equivalent to align the mineral resource estimate with the net smelter return cut-off that has been used to define the parameters of the mine plan in this PEA. There has been no change to the cut-off grade of 0.2g/t gold for the oxide gold resource estimate.

To ensure that the reported resource estimate exhibits reasonable prospects for economic extraction, it is limited within a pit shell generated about copper equivalent grades in blocks classified in the Indicated and Inferred categories at a copper price of $2.00/lb (including a 45° pit slope, $1.50/t mining cost for mineralized material and waste, and $6.00/t total site operating cost).

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This test indicates that some of the deeper mineralization may not be economic due to the increased strip ratios that result from the local topographic configuration. Additionally, the oxide resources are listed in Table 14-13, but for this PEA, this material has been treated as waste. It is important to understand that Table 14-12 and Table 14-13 list mineral resources; mineral resources are not mineral reserves as they do not have demonstrated economic viability. There is no certainty that all or any part of the mineral resource will be converted into a mineral reserve upon application of modifying factors.

Table 14-12: Supergene and Primary Sulphide Zone Mineral Resource Estimate

Cut-off Grade Mt Total Cu (%) Mo (%) Au (g/t) CuEq (%) (CuEq%) (1)

INDICATED MINERAL RESOURCES

0.2 2,817 0.38 0.011 0.07 0.49

0.25 2,484 0.41 0.012 0.07 0.53

0.3 2,165 0.44 0.013 0.08 0.57

0.35 1,851 0.48 0.014 0.09 0.61

0.4 1,545 0.51 0.015 0.09 0.65

0.45 1,249 0.56 0.015 0.10 0.71

0.5 997 0.61 0.016 0.11 0.77

0.55 811 0.65 0.017 0.12 0.83

0.6 660 0.70 0.018 0.13 0.88

0.65 544 0.75 0.018 0.13 0.94

0.7 454 0.80 0.019 0.14 0.99

INFERRED MINERAL RESOURCE

0.2 1,396 0.31 0.010 0.05 0.40

0.25 1,135 0.34 0.011 0.05 0.43

0.3 921 0.37 0.012 0.05 0.47

0.35 731 0.40 0.012 0.06 0.51

0.4 571 0.44 0.013 0.06 0.55

0.45 421 0.48 0.014 0.06 0.59

0.5 302 0.52 0.014 0.07 0.64

0.55 219 0.56 0.015 0.07 0.69

0.6 162 0.59 0.015 0.07 0.73

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Cut-off Grade Mt Total Cu (%) Mo (%) Au (g/t) CuEq (%) (CuEq%) (1)

0.65 117 0.63 0.016 0.07 0.77

0.7 76 0.68 0.016 0.07 0.81

(1) CuEq=Cu% + (Au g/t*0.583) + (Mo%*6). Metal prices: $2.00/lb Cu, $800/oz Au, $12.00/lb Mo (assumes 100% mining and metallurgical recoveries). (2) “Base case” cut-off grade for Supergene and Primary Zone resources is 0.30% CuEq

Table 14-13: Leach Zone Oxide Mineral Resource Estimate

Cut-off Grade Mt Au (g/t) Total Cu (%) Mo (%) (Au g/t) (1)

INDICATED MINERAL RESOURCE

0.1 799 0.18 0.04 0.016

0.15 483 0.22 0.04 0.017

0.2 243 0.27 0.04 0.018

0.25 126 0.31 0.04 0.018

0.3 57 0.34 0.05 0.017

0.35 20 0.39 0.07 0.017

0.4 5.6 0.44 0.10 0.017

INFERRED MINERAL RESOURCE

0.1 213 0.14 0.07 0.011

0.15 66 0.18 0.09 0.011

0.2 17 0.23 0.11 0.008

0.25 3 0.30 0.10 0.010

0.3 1 0.35 0.11 0.010

0.35 0 0.44 0.12 0.005

0.4 0.1 0.51 0.09 0.001

(1) “Base case” cut-off grade for Leach Zone resource is 0.2 g/t Au.

Figure 14-16 shows the distribution of the base case copper resource. Figure 14-17 shows the distribution of base case resources, including the Leach Zone gold resource.

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Figure 14-16: Extents of Base Case Indicated and Inferred Copper Resources (Sim, 2013)

Figure 14-17: Extents of Base Case Resources Including LX Zone Gold Resource (Sim, 2013)

The proportion of Supergene and Primary Zone resources are listed in Table 14-14.

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Table 14-14: Copper Mineral Resource Estimate by Type

Cut-off Supergene Primary Grade Cu Mo Au Cu Mo Au Mt Mt (CuEq%) (%) (%) (g/t) (%) (%) (g/t) Indicated Mineral Resource 0.2 898 0.52 0.008 0.07 1,919 0.32 0.013 0.07

0.25 809 0.55 0.008 0.07 1,675 0.35 0.014 0.08

0.3 701 0.60 0.009 0.08 1,464 0.37 0.015 0.08

0.35 601 0.65 0.010 0.09 1,250 0.39 0.015 0.09

0.4 517 0.70 0.012 0.09 1,028 0.42 0.016 0.09

0.45 448 0.75 0.013 0.10 801 0.45 0.017 0.10

0.5 392 0.80 0.014 0.11 605 0.48 0.018 0.11

0.55 349 0.84 0.015 0.11 462 0.51 0.019 0.12

0.6 318 0.87 0.016 0.12 342 0.55 0.020 0.13

0.65 292 0.90 0.017 0.13 252 0.58 0.020 0.14

0.7 270 0.93 0.017 0.13 183 0.61 0.021 0.15 Inferred Mineral Resource 0.2 123 0.38 0.004 0.06 1,273 0.30 0.011 0.05

0.25 109 0.40 0.005 0.06 1,026 0.33 0.012 0.05

0.3 90 0.44 0.005 0.06 831 0.37 0.012 0.05

0.35 71 0.48 0.006 0.06 660 0.40 0.013 0.06

0.4 56 0.52 0.007 0.06 515 0.43 0.014 0.06

0.45 42 0.57 0.007 0.06 379 0.46 0.014 0.06

0.5 33 0.62 0.008 0.06 269 0.50 0.015 0.07

0.55 26 0.66 0.009 0.06 193 0.54 0.016 0.07

0.6 22 0.69 0.009 0.06 140 0.58 0.016 0.07

0.65 18 0.73 0.010 0.06 99 0.61 0.017 0.07

0.7 14 0.76 0.010 0.06 62 0.66 0.017 0.07

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14.14 Comparison with Previous Resource Estimate The previous mineral resource estimate for the Taca Taca deposit is described in the Technical Report dated June 21, 2012 (effective date: April 11, 2012). The previous and current resource estimates are compared in Table 14-15 for the copper mineral resources and in Table 14-16 for the Leach Zone mineral resources.

Table 14-15: Comparison of Copper Mineral Resources – April 2012 vs. Nov 2012

Cut-off November 2012 April 2012 Grade Cu Mo Au Cu Mo Au Mt Mt (CuEq%) (1) (%) (%) (g/t) (%) (%) (g/t) Indicated Mineral Resource 0.25 2,484 0.41 0.012 0.07 1,060 0.51 0.016 0.10 0.3 (2) 2,165 0.44 0.013 0.08 995 0.53 0.017 0.11 0.35 1,851 0.48 0.014 0.09 918 0.56 0.017 0.11 0.4 1,545 0.51 0.015 0.09 824 0.59 0.018 0.12 0.45 1,249 0.56 0.015 0.10 728 0.62 0.018 0.12 0.5 997 0.61 0.016 0.11 638 0.66 0.019 0.13 0.55 811 0.65 0.017 0.12 558 0.70 0.019 0.14 0.6 660 0.70 0.018 0.13 479 0.75 0.020 0.15 0.65 544 0.75 0.018 0.13 412 0.79 0.020 0.16 0.7 454 0.80 0.019 0.14 358 0.84 0.020 0.16 Inferred Mineral Resource 0.25 1,135 0.34 0.011 0.05 1,778 0.37 0.012 0.06 0.3 (2) 921 0.37 0.012 0.05 1,479 0.40 0.012 0.07 0.35 731 0.40 0.012 0.06 1,196 0.44 0.013 0.07 0.4) 571 0.44 0.013 0.06 938 0.48 0.014 0.08 0.45 421 0.48 0.014 0.06 717 0.52 0.014 0.09 0.5 302 0.52 0.014 0.07 538 0.58 0.015 0.09 0.55 219 0.56 0.015 0.07 410 0.63 0.015 0.10 0.6 162 0.59 0.015 0.07 314 0.69 0.016 0.10 0.65 117 0.63 0.016 0.07 242 0.76 0.016 0.11 0.7 76 0.68 0.016 0.07 189 0.82 0.016 0.11

(1) CuEq=Cu% + (Au g/t*0.583) + (Mo%*6). Metal prices: $2.00/lb Cu, $800/oz Au, $12.00/lb Mo. Assumes 100% mining and metallurgical recoveries as these remain uncertain. (2) “Base case” cut-off grade for Supergene and Primary Zone resources is 0.30% CuEq.

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Table 14-16: Comparison of Leach Zone Mineral Resources – April 2012 vs. Nov 2012

Cut-off November 2012 April 2012 Grade Au Cu Mo Au Cu Mo Mt Mt Au g/t (1) (g/t) (%) (%) (g/t) (%) (%) Indicated Mineral Resource 0.1 799 0.18 0.04 0.016 492 0.20 0.04 0.017 0.15 483 0.22 0.04 0.017 327 0.23 0.04 0.017 0.2 243 0.27 0.04 0.018 198 0.27 0.04 0.018 0.25 126 0.31 0.04 0.018 112 0.31 0.05 0.018 0.3 57 0.34 0.05 0.017 51 0.36 0.05 0.018 0.35 20 0.39 0.07 0.017 22 0.42 0.06 0.018 0.4 5.6 0.44 0.10 0.017 10.0 0.48 0.07 0.018 Inferred Mineral Resource 0.1 213 0.14 0.07 0.011 336 0.17 0.06 0.015 0.15 66 0.18 0.09 0.011 157 0.22 0.06 0.017 0.2 17 0.23 0.11 0.008 81 0.26 0.07 0.017 0.25 3 0.30 0.10 0.010 34 0.31 0.06 0.017 0.3 1 0.35 0.11 0.010 13 0.36 0.06 0.017 0.35 0 0.44 0.12 0.005 5 0.43 0.09 0.016 0.4 0.1 0.51 0.09 0.001 2.5 0.49 0.12 0.016

(1) “Base case” cut-off grade for Leach Zone resource is 0.2 g/t Au.

Drilling conducted by Lumina Copper since the previous resource estimate has resulted in some significant changes to the estimated mineral resources. These changes include the following: (Note that changes to the base case copper resources are shown in Figures 14-18 and 14-19):

 The base case cut-off grade for the copper mineral resource estimate has been decreased from 0.4% copper equivalent to 0.3% copper equivalent to align the mineral resource estimate with the net smelter return cut-off that was used to define the parameters of the mine plan incorporated in the PEA. There has been no change to the cut-off grade of 0.2g/t gold for the oxide gold resource estimate.  Indicated-class copper resource estimate has almost doubled because previously Inferred- class resources have been upgraded in the main deposit area, and new resources were delineated in the northern and northwestern extension areas. The additional resources in the northern and northwestern extension areas tend to be lower grade when compared to the main deposit area.  Inferred-class resources have decreased because delineation drilling has upgraded these resources into the Indicated category.  Leach Zone gold resources have changed because the previously Inferred-class resources have been upgraded to the Indicated category.

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Figure 14-18: Changes to Base Case Indicated Resources Nov 2012 vs. April 2012 (Sim, 2013)

Figure 14-19: Changes to Base Case Indicated & Inferred Resources Nov 2012 vs. April 2012 (Sim, 2013)

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14.15 Discussion of Factors Materially Affecting Mineral Resources

There are no known factors related to metallurgical, environmental, permitting, legal, title, taxation, socio-economic, marketing, or political issues which could materially affect the mineral resource estimates. The Project risks are discussed in detail in Section 25.2.1.

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15 Mineral Reserve Estimates

This item is not applicable for this PEA.

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16 Mining Methods

The Taca Taca deposit is amenable to conventional, large-scale, open pit mining methods. FC evaluations were conducted to determine potentially economic pit limits and the pushback development sequence. This work was based on the deposit block model described in Item 14.0. Six mining phases were designed and a mine production schedule was developed using a declining cut-off grade policy to maximize project present values.

16.1 Economic Pit Limit Evaluations

Table 16-1 summarizes the recovery and economic parameters used in the FC analyses. Areas shaded in gray are not applicable to the referenced zones.

Table 16-1: Floating Cone Recovery and Economic Parameters

Zone Primary Secondary Oxide / Leach Cap Sulfide Sulfide High Cu Low Cu / Hi Au

Cu recovery, % 92% 87% 70.0%

Mo recovery, % (thru Mo con) 60% 55%

Au recovery, % 65% 60% 45% 70%

Ag recovery, % 50%

Cu payable, % 96.5% 96.5% 96.5%

Mo payable, % 85.0% 85.0%

Au payable, % 70.0% 70.0% 70.0% 99.5%

Ag payable, % 98.0%

Cu refining, $/lb payable 0.06 0.06 0.06

Mo refining, $/lb payable 0.00 0.00

Au refining, $/oz payable 7.50 7.50 7.50 2.00

Cu con freight & treatment, $/dmt 155.00 155.00 155.00

Mo con freight & treatment, $/dmt 90.00 90.00

Cu con grade, % Cu 28% 33% 40%

Mo con grade, % Mo 48% 48%

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Zone Primary Secondary Oxide / Leach Cap Sulfide Sulfide High Cu Low Cu / Hi Au

Cu FSR cost, $/lb payable 0.32 0.28 0.242

Mo FSR cost, $/lb payable 0.10 0.10

Royalties (NSR basis): 10.0% 10.0% 10.0% 5.5%

Base mineralized material mining, $/t 1.40 1.40 1.40 1.40

Base waste mining, $/t 1.60 1.60 1.60 1.60

Incremental haulage, $/t/bench below 3605 el 0.035 0.035 0.035 0.035

Sustaining mine capital, $/t mined 0.38 0.38 0.38 0.38

Mineralized material processing, $/t 4.50 4.50 4.50 9.00

General/admin, $/t 0.80 0.80

Internal cut-off, NSR $/t 5.10 5.10 4.30 8.80

Breakeven cut-off, NSR $/t 7.08 7.08 6.28 10.78

Internal cut-off, % CuEq 0.17 0.18 0.18

Breakeven cut-off, % CuEq 0.24 0.25 0.27

Internal cut-off, Au gpt 0.38

Breakeven cut-off, Au gpt 0.46

Low grade sulfide stockpile:

LG stockpile rehandling cost, $/t 1.10 1.10

Model Cu recovery at LG stockpile cut-off 87% 82%

Recovery losses 7% 7%

LG stockpile cut-off, % CuEq 0.24 0.25

LG stockpile cut-off, block model NSR $/t 6.74 6.78

Leach cap (oxide) material above the cut-offs indicated in Table 16-1 was investigated for possible batch processing after sulfide materials are exhausted. This material, along with dyke zones (which have low Cu grades and are mostly from the leach cap horizons of the deposit), would be stockpiled for potential processing at the end of the mine’s life. Low grade sulfide mineralized material stockpile cut-off grades account for rehandling and incremental haulage costs for the stockpiled material.

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W&N, with sub consultants FSR performed geotechnical analyses in support of scoping-level pit slope recommendations for the Project. These analyses were based primarily upon geotechnical data collected by W&N and Lumina Copper personnel during the course of exploration core drilling and dedicated geotechnical drilling in the years 2010 to 2012.

Six slope sectors were defined to control the FC projections and subsequent pushback designs. Overall slope angles for FCs ranged from about 38° in the north walls, 42° in the east and southeast walls, to 44-46° in the southern walls, and about 42° in the northwestern quadrant of the pit.

Recovery and valuation subroutines were developed to compute NSRs, net values, and recoverable metal grades for each block. For this PEA, mineral resources classified as indicated and inferred were allowed to contribute revenues to the economic pit limit evaluations. None of the mineral resources have been classified as measured.

In situ bulk densities range between 1.90 kg/m3 for overburden to 2.65 kg/m3 for primary sulfides. Densities vary by rock type in the deposit model.

The base case is defined by prices of $2.00/lb Cu, $12.00/lb Mo, $1,100/oz Au, and $14.00/oz Ag. Price sensitivities were conducted in increments of $0.25/lb Cu, ranging from $1.25/lb to $3.00/lb Cu; other metal prices were proportioned from the base copper price. An internal NSR cut-off of $5.10/t was applied when estimating contained mineral resources for each pit shell (note that NSR values vary with the metals prices and are recomputed prior to each FC run). A 1.5% per bench discount rate was applied to the net block values in the FC analyses, which is roughly equivalent to 9% per year at a sinking rate of six benches per year. This discounting approximates the effects of time value of money; penalizing zones of marginal mineralized material blocks and/or high incremental stripping ratios in order to maximize present values for the Project.

The $2.00/lb Cu shell with indicated and inferred resources was selected as the basis for subsequent pit designs. A shell at higher Cu prices through $2.75/lb could have been selected given recent market history, more than doubling the contained mineral resource tonnages – but at lower incremental head grades. This would require additional tailing storage capacity and would extend the Project life well beyond 30 years. It was felt that the $2.00/lb Cu shell would capture the bulk of the Project’s potential present value and would be more reasonable to use at this stage of evaluation.

16.2 Mining Phase/Pit Designs

The Taca Taca ultimate pit and internal mining phases were designed to accommodate large-scale mining equipment operating on 15 m benches. This equipment includes rotary blasthole drills capable of drilling holes up to 270-311 mm in diameter, 60 m3 electric shovels, a 42 m3 hydraulic shovel, a 40 m3 FEL, and off-highway haulage trucks with payload capacities of 290-363 t.

Pit walls were smoothed from the basis FC shell to minimize or eliminate, where possible, noses and notches that could affect slope stability. Internal haulage ramps were included to allow for truck access to working faces on each level. The basic parameters used in the design of six mining phases, or pushbacks, are summarized in Table 16-2 below.

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Table 16-2: Basic Pit Design Parameters

Parameters Unit Value Bench height m 15 Haul road width (including ditch & safety berm) m 44 Internal ramp gradient % 10 Minimum pushback width m 90 Typical pushback width m 120+

Inter-ramp and bench face angles vary according to the W&N recommendations summarized in Table 16-3 for six slope design sectors. The sector boundaries were adjusted as necessary to honor the pit wall orientations suggested by the geotechnical analysis.

Table 16-3: Pit Design Recommendations

Slope Sectors Design Parameter 1 2 3 4 5 6 N NE E-Central SE SW NW

Inter-ramp angle (degrees) 43.5 48.5 48.0 50.0 48.5 48.5 Bench face angle (degrees) 55 62 61 64 62 62

Catch bench vertical interval, m 30 30 30 30 30 30 Catch bench width, m 10.6 10.6 10.4 10.5 10.6 10.6

The main ultimate pit is approximately 3,100 m wide E-W and over 2,700 m N-S at the crest. Elevations range from 2,915 m at the bottom to 3,680 m at the highest crest in the concave portion of the southeast wall. The maximum overall wall height is 765 m. The smaller pit to the north, mined in Phase 5 and targeting some lower grade sulfides near surface, is about 720 m E-W and 870 m N-S, with elevations ranging between 3,425 and 3,620 m and a maximum depth of 195 m. The surface areas of the main and north pits are 776 and 67 ha, respectively.

The ultimate pit design is shown in Figure 16-1. Dual ramp access is provided to the 3,080 m elevation. Grid lines are on 1,000 m intervals.

The salar sediments are expected to reach a maximum depth of roughly 50 m along the eastern crest of the ultimate pit. The sediments would be dewatered and excavated to the hard rock contact prior to pit expansions into the salar. Additionally, an impervious barrier would be constructed immediately east of the pit to prevent brine inflows.

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Figure 16-1: Ultimate Pit Design (WLRC, 2013)

The starter pit (Phase 1) lies mostly in the northeast portion of the main ultimate pit area, downslope of the hill covering much of the deposit, and is located where the supergene zone rises closest to the topographic surface. Approximately 230 Mt of preproduction stripping will be required to reach the top of mineralized material in Phase 1. The Phase 1 pit is about 1,070 m wide E-W and roughly 1,000 m N-S, with elevations ranging between 3,290 and 3,740 m for a total depth of 450 m. The surface area of Phase 1 is 126 ha.

16.3 Mine Production Schedule

The proposed development plan for the Project envisions starting with two grinding lines rated at a nominal capacity of 60,000 t/d each, for a total initial milling rate of 120,000 t/d. A third grinding

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circuit is contemplated to be brought on line at the beginning of Year 8, increasing the total milling rate to a maximum of 180,000 t/d.

The basic parameters used to develop the mine production schedule are summarized in Table 16-4 below.

Table 16-4: Production Scheduling Parameters

Parameters Unit Value

Annual target concentrator feed

production rates: Y1 Mt 34.92 Y2-Y7 Mt 43.20 Y8 Mt 62.75 Y9+ Mt 64.80 Daily milling rates: Y1-Y7 t 120,000 Y8+ t 180,000 Operating hours per shift h 12 Operating shifts per day shifts/d 2 Operating days per week d/wk 7 Scheduled operating days per year d/y 360

A nine-month concentrator ramp-up schedule is incorporated into the Year 1 concentrator feed production target. Similarly, the Year 8 production target includes a 4 to 5 month ramp-up period for the third grinding line.

Estimates of the mineral resources contained within the mining phases were based on the November 2012 deposit model, include material classified as indicated and inferred, and were based on the metallurgical recoveries and costs presented in Table 16-1. NSRs were computed at base case metals prices of $2.00/lb Cu, $12.00/lb Mo, $1,100/oz Au, and $14.00/oz Ag.

The Taca Taca deposit is a well-disseminated Cu-Mo-Au mineralized system that has large mineralized zones above the anticipated cut-off grades. The sample compositing and block grade interpolation process used to construct the deposit model are believed to incorporate sufficient dilution and, hence, no additional dilution factors were applied. Mining recovery of mineralized material within the open pit is expected to be virtually 100%. Low grade sulfide stockpile recovery is estimated at 97%.

Bulk densities were stored in the deposit model for each block and vary by geologic zone. Table 16-5 summarizes the zones and the corresponding in situ dry densities. Any undefined blocks were assigned a default density of 2.60 t/m3.

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Table 16-5: Geologic Zones and Bulk Densities

Geologic Zone Model Code Density (t/m3) Overburden 1 1.90 Leach Cap (oxide) 2 2.50 Secondary Sulfide (supergene) 3 2.62 Primary Sulfide (hypogene) 4 2.65 Dyke 5 2.60

A production scheduling analysis indicated that a variable cut-off strategy employing initial sulfide NSR cut-offs of $8.60-12.60/t during the first nine years and declining to internal cut-offs ($5.10/t) by Year 22 would improve the potential present value of profits by about 5%. An internal NSR cut- off of $5.10/t was applied to all stockpiled dyke material. Low copper (≤ 0.02% TCu) oxide blocks were considered for stockpiling and potential precious metal extraction if the computed NSR was above an $8.80/t internal cut-off. High copper oxide blocks were considered for stockpiling and potential batch processing if the NSR value was above a $4.30/t internal cut-off.

The mine production schedule was developed for the Project using estimates of indicated and inferred mineral resources contained within each of six mining phases. Only secondary and primary sulfides above a declining NSR cut-off were considered as potential concentrator feed; all mineralized dyke and oxide material would be stockpiled only. Stockpiled low grade sulfide mineralized material would be reclaimed and fed to the mill immediately after the open pits are exhausted.

Table 16-6 summarizes the mine production schedule based on the declining cut-off grade policy described above. This schedule includes mineralized material processing through a third grinding line in the concentrator beginning at the start of Year 8. Peak material handling rates in the mine, including waste, would be about 508,000 t/d, or 183 Mt per year. The stripping ratio, based on declining cut-offs, is projected at 1.57 (tonnes of waste and stockpiled material per tonne of sulfide material milled).

Mine preproduction stripping will require 2.5-3 years, including site preparation and a gradual build- up of equipment and trained personnel. Mine stripping targets for preproduction Years -3, -2, and - 1 are 15, 86, and 150 Mt, respectively. Total preproduction stripping is estimated at 251 Mt, which includes some advanced stripping in mining Phase 2.

Nearly 21 Mt of dyke material would be stockpiled for potential batch processing after exhausting the sulfide resource. Similarly, about 5 Mt each of oxide copper and oxide gold resources would also be stockpiled. Almost 56 Mt of low grade sulfide material would be stockpiled during Years 1- 15, of which about 54 Mt (97%) would be reclaimed in Year 28. Over 78 Mt of low grade sulfides (above an internal cut-off) would not meet the stockpiling cut-off and are treated as waste in the production schedule.

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Table 16-6: Taca Taca Mine Production Schedule

Cautionary Note: This PEA, including this mine plan and production schedule, is preliminary 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 results of this PEA will be realized. Mineral resources that are not mineral reserves have not demonstrated economic viability.

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A breakdown of the total concentrator feed by indicated and inferred mineral resource classifications is listed in Table 16-7 below.

Table 16-7: Resource Classification of Concentrator Feed – PEA Schedule Concentrator Feed Classification Ktonnes Cu % Mo % Au g/t Ag g/t

Measured 0 - - - -

Indicated 1,545,000 0.46 0.013 0.09 0.6

Inferred 106,000 0.43 0.005 0.09 1.1

The concentrator is expected to operate for approximately 28 years based on the mine production schedule presented in Table 16-6. The mine production schedule presented in Table 16-6 is comprised of 1,545 Mt of indicated mineral resources grading 0.46% Cu, 0.013% Mo, 0.09 Au g/t, and 0.6 Ag g/t and 106 Mt of inferred mineral resources grading 0.43% Cu, 0.005% Mo, 0.09 Au g/t, and 1.1 Ag g/t which would be processed by the concentrator over the life of the Project. The production schedule excludes milling of the dyke and oxide resources.

This 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. Inferred mineral resources have a great amount of uncertainty as to their existence and as to whether they can be mined legally or economically. There is no certainty that the PEA will be realized.

16.4 Waste Rock Storage and Stockpiling

Approximately 31 Mt of mineralized dyke and oxide material will be stockpiled to the west of the main open pit. Over 56 Mt of low grade sulfides will be stockpiled just west of the oxide stockpile. Two waste rock storage facilities (WRSFs) will be required to store an estimated 2,520 Mt of sub- grade material. Of this total, only 414 Mt of waste rock will be extracted from the secondary and primary sulfide geologic zones; the vast majority of waste will be oxide. Figure 16-2 shows the anticipated locations of the low-grade stockpiles and WRSFs.

Portions of the North WRSF and nearly all of the South WRSF are contemplated to be constructed atop the Salar de Arizaro near the open pit. The North WRSF design has a storage capacity of nearly 350 Mt, while the South WRSF, as shown, could store a little over 2,200 Mt of waste rock. The crest elevation of both WRSFs is approximately 165 m above the present salar surface.

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Figure 16-2: Waste Rock Storage and Oxide Stockpile Facilities (WLRC, 2013)

16.5 Mine Equipment

Large-scale open pit mining equipment operating on 15 m benches could be used to extract mineralized material and waste rock for the Project. Electric drills and shovels were chosen because of their lower operating costs and better reliability. The rest of the mining fleet is envisioned to consist of diesel-powered machinery to provide maximum flexibility in pit operations.

The production schedule presented in Table 16-6 was used as the basis for determining equipment requirements. It was assumed that the Owner would be performing all mining services, including preproduction stripping, blasting, and maintenance and repair activities. Mine operations would be scheduled for two 12-hour shifts per day, 7 d/wk, for a total of 360 d/y. A provision was made for five lost days per year due to adverse weather and holidays. Four crews, rotating between day and night shifts, is contemplated to provide continuous operator and maintenance labor coverage for the mine.

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The selected primary mining fleet includes the following:

 Rotary blasthole drills capable of 270- to 311-mm-diameter holes  60 m3 rope shovels  42 m3 hydraulic shovel  40 m3 FEL  363 t off-highway haul trucks  635-kW and 435-kW crawler dozers  600-kW rubber-tired dozer  370-kW rubber-tired dozers  400-kW and 220-kW motor graders  120,000 l (134 t) water trucks

A portable crushing and screening plant with a rated capacity of about 500 t/h is envisioned to produce crushed rock for blasthole stemming and road surfacing throughout the project area. A 30 t vibratory compactor is contemplated to be needed for haul road maintenance/construction and engineered earthworks.

Other auxiliary equipment could be used for miscellaneous earthworks and construction around the mine and WRSF sites, cleaning out minor ditches and sumps, equipment assembly and maintenance, fueling and lubrication services, transporting mine equipment and electric power cables, loading/unloading mining supplies and repair parts, transporting mine personnel, and other mine support duties.

The mine equipment fleet is presented for selected years (1, 7, 13, and 20) in Table 16-8. The peak haul truck requirement of 58 units occurs in Year 7.

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Table 16-8: Mine Equipment Fleet for Selected Years

Equipment Y1 Y7 Y13 Y20 Primary: Crawler Rotary Drills, 270-mm 3466 Crawler Rotary Drills, 311-mm 8865 Secondary Percussion Drills, 89-mm 2222 Electric Shovels, 60-cu-m 5554 Hydraulic Shovel, 42-cu-m 1111 Loader, 40-cu-m 1111 Loader, 12-cu-m (R&D) 1111 Haul Trucks, 363-tonne 38 58 58 51 Lt. Haul Trucks, 91-tonne (R&D) 3333 Dozers, 635-kW 3333 Dozers, 435-kW 7777 R.T. Dozer, 600-kW 1111 R.T. Dozers, 370-kW 5555 Graders, 400-kW 2333 Graders, 220-kW 6888 Water Trucks, 120,000-liter 5555 Auxiliary: ANFO/Slurry Truck, 25‐t3333 Blasthole Dewatering Truck 2 2 2 2 Powder Truck, 1‐t2222 Powder Magazine, 14‐t5555 AN Storage Bin, 60‐t 10101010 Emulsion Storage Bin, 65‐t2222 Integrated Tool Carrier, 105‐kW 3333 Hydraulic Hammer 1 1 1 1 Crawler Excavator, 6‐cu‐m1111 Crawler Excavator, 3‐cu‐m2222 Backhoe/Loader, 1.1‐cu‐m1111 Vibratory Compactor, 30‐t1111 Portable Aggregate Plant, 500 tph 1 1 1 1 Mobile Generator (2 MW) 1 1 1 1 Cable Reel Truck 1 1 1 1 Aerial Boom Truck 2 2 2 2 All‐Terrain Crane, 80‐t1111 Crawler Crane, 200‐t2222 Equipment Transporter, 200‐t1111 Fuel/Lube Truck ‐ Large 2 2 2 2 Fuel/Lube Truck ‐ Small 1 1 1 1 Mechanic Field Service Truck 2 2 2 2 Medium‐Duty Field Truck w Crane 2 2 2 2 Off‐Road Tire Handling Truck 1 1 1 1 Tire Handling Forklift 1 1 1 1 Shop Forklift, 16‐t1111 Shop Forklift, 5‐t2222 Light Plant 16161616 Pickup Truck, 0.5‐t, 4‐WD 32 32 32 32 Crew Bus 5 5 5 5 Ambulance 1111 Truck Dispatch System 1111

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16.6 Mine Workforce

Mine workforce requirements were estimated on the basis of working two 12-hour shifts per day, 7 d/wk, 52 wk/y. A standard, four-crew rotating work schedule for craft labor and front-line supervision will be used for around-the-clock coverage.

It is anticipated that expatriate personnel will be hired to help establish safe and efficient mine operations and maintenance systems, and to train Argentine nationals in proper work procedures. As the skill levels of national workers increase, expatriates would be phased out with the eventual goal of minimal expatriate staffing.

Mine personnel requirements are based on the Owner performing all mining functions. The use of contractors for such activities as supplemental preproduction stripping, down-hole explosives supply and shot services, and contract maintenance and repair have not been incorporated into the workforce and mining cost estimates.

The projected mine workforce is presented for selected years (1, 7, 13, and 20) in Table 16-9. A peak mine workforce of 874 is projected in Year 7.

Table 16-9: Projected Mine Workforce for Selected Years

Position Y1 Y7 Y13 Y20 Operations, Craft Labor 373 450 441 411 Maintenance, Craft Labor 229 274 270 250 Supervision & Technical 150 150 150 150 Total Mine Department 752 874 861 811

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17 Recovery Methods

17.1 Introduction

The PEA production schedule assumes the Taca Taca concentrator will initially process 120,000 t/d of ROM mineralized material, producing copper concentrate, molybdenum concentrate, and tailings. Prior to Year 8, an expansion is contemplated to be undertaken to increase the capacity of the concentrator to 180,000 t/d.

The process includes crushing and grinding of the ROM mineralized material, bulk copper – molybdenum rougher and cleaner flotation, regrinding, copper and molybdenum separation and molybdenum cleaner flotation, and dewatering of copper and molybdenum sulphides. The copper and molybdenum concentrates are envisioned to be transported by truck from the concentrator to the rail load-out facility from where it will be transported by rail to the Port of Mejillones in Chile for shipment to customers. The flotation tailings are planned to be thickened before placement in the TSF.

17.2 Process Design Basis and Design Criteria Summary

The key process criteria for the plant design and operating costs are provided in Table 17-1.

Table 17-1: Summary of Key Process Design Criteria

Criteria Description Units Design Source Crusher Feed t/d 120 000 Lumina Copper (Yrs 1-7) Crusher Availability % 70 Ausenco Crusher Throughput t/h 7 143 Calculation (Yrs 1-7) Mill Throughput Mt/y 43.8 Lumina Copper (Yrs 1-7) Mill/Flotation Availability % 92 Ausenco Mill Throughput t/h 5 414 Calculation (Yrs 1-7) Bond Ball Mill Work Index BWI kWh/t 17.8 Test work/Ausenco SMC Test Axb 52 Test work/Ausenco

Primary Grind Size P80 µm 150 Test work

Concentrate Regrind Size P80 µm 30 Test work % Cu 1.1 Head Grade (Design) % Mo 0.02 Mine Plan Year 1-2 g/t Au 0.12 Copper Recovery % 88-92 Ausenco/Test Work Molybdenum Recovery % 56-57

Gold Recovery % 61-65 Concentrate Filtration Rate t/m2/h 0.85 Ausenco/Test work Concentrates Thickening Flux t/m2/h 0.25 Ausenco Tailings Thickening Flux t/m2/h 0.73 Ausenco/Test work Tailings Thickener Underflow Density % w/w 60 Ausenco

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The key criteria selected for the concentrator design are:

 Treatment of 120,000 t/d in two 60,000 t/d parallel lines (except for the molybdenum circuit which is a single circuit);  Design availability of 92% (after ramp-up), resulting in 8,059 operating hours per year, with standby equipment in critical areas;  An expansion prior to Year 8 adding another 60,000 t/d parallel line consisting of grinding, pebble crushing, bulk rougher flotation, and regrind (bulk cleaner flotation and molybdenum circuit do not require increased capacity);  Sufficient plant design flexibility has been considered for treatment of all mineralized material types at design throughput.

17.3 Flow Sheet Development and Equipment Sizing

The design for the Project was based on the process design criteria and metallurgical test results. An overall process flow diagram for the Project is shown in Figure 17-1. The concentrator is designed to have two parallel lines processing 60,000 t/d each from crushing to bulk copper – molybdenum cleaning; after this stage there is one circuit for molybdenum copper separation and molybdenum cleaner flotation.

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Figure 17-1: Overall Process Flow Diagram

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The key unit process operations that are well proven in the industry are envisioned to include:

 Mineralized material from the open pit is crushed using two 600 kW, 60” x 89”, primary gyratory crushers in parallel to a crushed product size of nominally 80% passing (P80) 150 mm and fed onto the stockpile feed conveyor;  One conical stockpile of crushed material with a live capacity of 24 h, with three apron feeders for each SAG mill train, each capable of feeding 120% of the full mill throughput;  Two 23 MW SAG mills, 12.19 m diameter x 7.01 m effective grinding length (EGL) (40’ x 23’), in closed circuit with pebble crushing;  Pebble crushing will be comprised of two MP1000 cone crushers, crushing to a product size of nominally 80% passing (P80) 12 mm;  Four 17 MW ball mills (two per train), 7.92 m diameter x 12.19 m EGL (26’ x 40’), in closed circuit with hydro-cyclones, grinding to a product size of nominally 80% passing (P80) 150 µm;  Bulk copper-molybdenum rougher flotation consisting of ten 300 m3 forced air tank flotation cells (five per line) to provide a total of 12 mins of retention time;  Bulk copper-molybdenum rougher concentrate regrinding in two 3.0 MW horizontal fine grinding Isamills (one per line) to a P80 of 30 µm;  Bulk copper-molybdenum regrind product thickening in a 38 m diameter high rate thickener (one per line);  Bulk copper-molybdenum first stage cleaner and cleaner scavenger flotation consisting of sixteen 20 m3 (eight per line) forced air tank flotation cells to provide a total of 10 min of retention time;  Bulk copper-molybdenum cleaner scavenger thickening in a 38 m diameter high rate thickener;  Bulk copper-molybdenum second stage cleaner flotation consisting of eight 10 m3 trough shaped flotation cells (four per line) to provide a total of 2.5 min of retention time;  Bulk copper-molybdenum third stage cleaner flotation consisting of six 10 m3 trough shaped flotation cells (three per line) to provide a total of 2.5 min retention time;  Bulk copper-molybdenum thickening in a 32 m diameter high rate thickener (one for both lines);  Copper-molybdenum separation rougher flotation consisting of five 10 m3 forced air tank flotation cells to provide a total of 7.5 min of retention time;  Molybdenum first stage cleaner and cleaner scavenger flotation consisting of five 2.5 m3 forced air tank flotation cells to provide a total of 7.5 min of retention time;  Molybdenum second stage cleaner flotation consisting of three 2.5 m3 trough shaped flotation cells to provide a total of 5 min of retention time;  Molybdenum third stage cleaner flotation consisting of three 2.5 m3 trough shaped flotation cells to provide a total of 2.5 min retention time;

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17.4 Comminution Circuit Design

The test work data and evaluation of bench scale flotation results indicate that the deposit is of relatively moderate competency as indicated by Bond Ball Mill data and the two JK drop weight (Axb) competency tests. An SABC circuit has been selected.

There will be two parallel SABC circuits each processing 60,000 t/d. Product from the SAG mill will gravity flow through a trommel screen and into the cyclone feed pumpbox where it will combine with the product from the two ball mills. The combined SAG and ball mill product will be pumped to two cyclone clusters (one dedicated cluster per ball mill) to classify the product; with the coarse in the underflow returned to the ball mill, and the fines in the overflow flowing via gravity to rougher flotation.

This circuit will be able to achieve the design throughput of 60,000 t/d for mineralized material with an Axb competency of 52 or greater. An Axb of 52 represents a 15% design factor increase in rock competency compared to the measured value of the Primary material of 60. More competent mineralized material will only be able to be processed at lower throughput rates.

The specific energy and mill sizing determined using Ausenco’s in-house method is shown in Table 17-2.

Table 17-2: Mill Design Criteria

Criteria Description Units Design Throughput (Total) t/h 5 435 Throughput (per line) t/h 2 717 Mill Type SAG Grate D/C Number of Mills (Total) 2 Pinion Power required kW 18 917 Mill Speed % Nc 75 Ball Charge Volume Nominal, operating % vol 13 Maximum for design % vol 18 Total Charge Volume Nominal, operating % vol 27 Maximum for design % vol 35 Mill Diameter Inside shell m 12.19 Mill Length EGL m 7.01 Selected Motor Power kW 23,000 Mill Type Ball Number of Mills (Total) 4

Grind Size P80 µm 150 Pinion Power required kW 15 290 Mill Speed % Nc 75 Ball Charge Volume Nominal, operating % vol 33

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Criteria Description Units Design Maximum for design % vol 40 Mill Diameter Inside shell m 7.92 Mill Length EGL m 12.19 Selected Motor Power KW 17,000

The selected ball mill power of 17,000 kW incorporates the allowances for drive train losses to determine the motor power installed from the pinion power required.

The selected motor power for the SAG mill incorporates similar allowances as well as an additional contingency to allow adjustment in the mill operating conditions to handle feed variability. These allowances and contingencies resulted in the selection of a 23,000 kW SAG mill.

17.4.1 Pebble Crushing

The pebble crushing circuit design includes the conveyors feeding the crushers as well as the conveyors returning the crushed pebbles to the SAG mill feed. The pebble crushing circuit is contemplated to be comprised of a single feed bin, with two reclaim feeders feeding two 900 kW pebble crushers for each line.

A pebble circulating load of 25% of the new feed rate has been assumed for the design of the pebble crusher, based on typical industry experience with mineralized material of similar competency. The conveyors are contemplated to be designed to handle peak loads of up to 45% of new feed.

17.4.2 Mill Circuit Classification

The classification circuit has been designed for a maximum circulating load of 250%. This is a typical industry design value for an SABC circuit. The SAG mill discharge slurry passes through a trommel screen with 12.5 mm apertures to remove pebbles, while the screen undersize flows into the cyclone feed pumpbox and combines with ball mill discharge. At the pumpbox, two separate suction pipes would provide feed to the two cyclone feed pumps sending feed to the ball mills.

The ball mills are envisioned to operate in closed circuit with cyclones to reduce the particle size to the desired P80 of 150 μm. Combined SAG and ball mill discharge is pumped to two cyclone clusters of 760 mm diameter hydro-cyclones (14 operating and three spare per cluster) to split the flow to the two secondary mills. Underflow from the hydro-cyclone clusters is envisioned to feed each ball mill operating in parallel for further grinding.

Overflow from the cyclone clusters flows via gravity to the flotation circuit.

17.5 Bulk Copper-Molybdenum Flotation Circuit Design

Mineralogical examination has highlighted that the copper mineralogy is fine grained. The flow sheet selected consists of rougher flotation, rougher concentrate regrind, and three stages of cleaner flotation. The mineralized material will require a fine regrind of the bulk rougher concentrate to a P80 of 30 µm to achieve adequate recovery. After regrinding, conventional flotation would be used to enable the production of saleable grade concentrates.

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17.5.1 Bulk Flotation Circuit Configuration

A rougher flotation circuit consisting of five 300 m3 tank flotation cells per train has been selected. The flotation circuit consists of a stepwise cell installation with one cell per step.

The first cleaner flotation circuit, including cleaner scavengers consists of seven 20 m3 tank flotation cells. The second cleaner circuit consists of four 10 m3 trough flotation cells. The third cleaner circuit consists of three 10 m3 trough flotation cells.

In the 2nd and 3rd cleaner applications smaller cells have been selected to avoid short circuiting with fewer cells, while the trough cell design provides a higher lip length and cell surface area to volume ratio making this style of cell more suited for this application than tank cells.

Subsequent optimization work may identify better alternate configurations including the potential use of column cells and/or Jameson cells.

17.5.2 Regrind Circuit Configuration

Rougher concentrate will report to the regrind cyclone feed pumpbox. The regrind cyclones will target a cut size P80 of 30 µm which will be achieved with a cluster of 250 mm cyclones, including 18 operating, 2 spares, and 3 blank nozzles for future additions. The cyclones will report to an M1000 Isamill. The Isamill discharge is combined with the cyclone cluster overflow and reports to the first cleaner pumpbox.

17.6 Copper-Molybdenum Separation and Molybdenum Cleaning

The flow sheet selected to separate the bulk concentrate consists of rougher flotation and four stages of cleaner flotation. The tailings from the rougher flotation is expected to be the final copper concentrate. Concentrate from the fourth cleaner flotation cells would be the final molybdenum concentrate.

17.6.1 Molybdenum Rougher Flotation

Molybdenum rougher flotation cells are planned to consist of a series of five 10 m3 forced air trough flotation cells. The cells would be arranged in a step configuration between each pair of cells. The tailings from the molybdenum roughers would be the final copper concentrate. Molybdenum flotation is planned to be done with nitrogen gas (N2) to minimize oxidation and consumption of NaHS.

Concentrate from the molybdenum rougher cells is planned to flow via gravity to the molybdenum rougher concentrate pumpbox and would then be pumped (one duty and one standby) to the molybdenum first cleaner flotation cell.

17.6.2 Molybdenum Cleaner Flotation

The molybdenum first cleaner is planned to consist of three 2.5 m3 forced air tank flotation cells, followed by molybdenum first cleaner scavengers consisting of two 2.5 m3 cells.

The molybdenum second cleaner is planned to consist of three 2.5 m3 forced air tank flotation cells, while the molybdenum third cleaner and fourth cleaner would each consist of three 2.5 m3 forced air tank flotation cells.

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Concentrate from the molybdenum fourth cleaner cells is planned to be pumped to the molybdenum concentrate thickener feed box. Tailings from the molybdenum fourth cleaner cells would be pumped to the molybdenum third cleaner cells.

17.7 Copper Concentrate Thickening and Filtration

Copper concentrate from the molybdenum rougher tailings pumpbox would then be pumped to a 32 m diameter high-rate thickener and would be thickened to approximately 60% w/w solids. Thickener underflow would be pumped to the copper concentrate filter feed tank.

Thickened copper concentrate slurry would be delivered to one of two copper concentrate filters. The filter presses are designed to reduce the moisture content of the concentrate prior to transport. Filter cake would be discharged through the floor of the filter building into a concrete bunker area underneath the press. Filter cake could then be removed from the bunker by an FEL and stored in the covered concentrate storage shed. Concentrate would then be reclaimed by an FEL from the stockpile and loaded into 20 t tipper trucks for transport to the rail load-out facility.

17.8 Molybdenum Concentrate Thickening and Filtration

Molybdenum flotation concentrate from the fourth cleaner cells would then be thickened to approximately 60% w/w solids in a 4 m diameter high-rate thickener. Molybdenum concentrate storage is envisioned to provide surge capacity ahead of the molybdenum concentrate filter.

The molybdenum concentrate filter has been sized based on a head grade of 0.02% Mo and a final concentrate grade of 53% Mo. As with the copper filters, the molybdenum filter size is based on filtration rates determined from previous experience.

17.9 Concentrate Rail Load-Out Facility

Copper concentrate would then be trucked from the concentrator to the rail load-out facility approximately 2.5 km to the northeast, adjacent to the new rail spur.

An FEL would be used to manage concentrate in the storage facility and dump concentrate into a feed bin from which a feed conveyor would then take the concentrate to a 500 t concentrate storage bin located over the rail spur. The storage bin would be equipped with vibrators and three hydraulically actuated clamshell type gates from which rail cars could be loaded as they are positioned under the concentrate storage bin.

17.10 Tailings Processing

The design basis chosen for this level of study includes a tailings thickener, with storage of thickened tailings in a nearby TSF and recovery of water from the TSF surface. Due to the low levels of pyrite in the feed, it is assumed that overall the tailings will be non-acid-generating (NAG) and separation of pyrite from the tailings stream for storage in a separate TSF is not required.

Ausenco sized the tailings thickener for the full expanded tonnage rather than build a new smaller thickener later, which resulted in the requirement for a 120 m diameter tailings thickener. Thickener underflow would be pumped to the TSF, while thickener overflow would gravity flow to the process water tank.

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17.11 Brine, Desalinated, and Process Water

Under the current design plan, the water systems will consist of three water types as follows:

 A brine water system for the grinding and rougher flotation areas  A desalinated water system for the bulk cleaner and molybdenum flotation circuits  A filtered brine water system for gland seal water and rougher launder water

The majority of the plant water requirements are envisioned to be met from the plant process water systems, which will be composed of recycled water streams i.e., thickener overflows and TSF return water, supplemented with brine and desalinated water as required. Process make-up water requirements and systems are discussed under Item 18.8.2

17.12 Reagents

Estimated reagent consumptions are shown in Table 17-3.

Table 17-3: Summary of Reagent Consumption

Criteria Description Units Design Source Collector Consumption (PAX) g/t 30 Test work Collector Consumption (Z6) g/t 5.2 Test work Collector Consumption (Z14) g/t 0.6 Test work Molybdenum Emulsion g/t 8 Test work Frother Consumption (MIBC) g/t 12 Test work Sodium hydrosulphide Consumption (NaHS) g/t 8-10 Test work Lime Consumption kg/t 2.28 Test work Flocculant Consumption (concentrate and tailings) g/t 37 Ausenco

17.13 Plant Layout & Description A layout has been prepared, placing the concentrator as reasonably close to the pit as possible. The layout is shown below in Figure 17-2. The concentrator layout has taken site topography and concession boundaries into account and works within the bounds imposed by preliminary locations of the pit, stockpiles, and WRSFs.

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Figure 17-2: Pebble Crushing – Mill Building Plan and Elevations

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18 Project Infrastructure

18.1 Overview

The Project is located in a remote area of northwestern Argentina, however there is reasonably good access to regional infrastructure. A network of paved and gravel roads from Salta to the towns of San Antonio de los Cobres and Tolar Grande provide road access to the Project. The road continues west beyond the Project to the Socompa Pass on the Chilean border and eventually to the port city of Antofagasta, Chile. The road conditions are generally good, but some upgrades to portions of the road closer to the Project will be required.

The Project is located within 10 km of the railway line that connects Salta with Antofagasta. It is expected that this railway will be used for concentrate transport to the Port of Mejillones, located approximately 65 km north of Antofagasta, and that consumables may also be transported by rail by adding tank or flatbed rail cars to the empty concentrate trains returning to the Project. Construction of a new rail spur, a new maintenance and repair facility for locomotives and railcars adjacent to the concentrate load-out facility, and rehabilitation of approximately 40 km of track between the Project and Socompa are envisioned to be required.

18.2 Roads

18.2.1 Mine Access Road A preliminary assessment was performed of the existing conditions of the road from Salta to the Project that will be used as the primary access during construction and operations (Figure 18-1). Access from Salta to the Project is along existing public roads that include Highway 51 from Salta to the Cauchira Salt Flat, approximately 223 km, and Road 27 from Cauchira to the Project, approximately 158 km, for a total distance of 381 km.

Access along Highway 51 consists of both paved and gravel road sections that are in good condition and are maintained by the government. The typical road width is approximately six to seven meters. Based on the general visual conditions of this section, no upgrades are required. From Highway 51 at Cauchira Salt Flat, the Project is accessed by Road 27, which is a gravel and dirt road. This road is approximately six to eight meters wide. The government also maintains this road; however, the section from the Pocitos Salt Flat to the Project is not as well maintained and may require some light maintenance by Lumina Copper during construction and operation. Also, along this section, there is a segment that contains several switchbacks, which is considered a local landmark. Prior to construction, Lumina Copper may need to develop a by-pass to preserve the switchback and increase the road width and turning radii.

The other section of Road 27 that will need to be upgraded lies between Tolar Grande and the Project, which is approximately 34 km long (Figure 18-2).

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Figure 18-1: General Mine Access Road (Ausenco, 2013)

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Figure 18-2: Mine Access Road (Ausenco, 2013)

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Design criteria for the roads is based on the proposed traffic volume, vehicle types, and desired road surface.

The mine access road connects the mine infrastructure roads at the rail load-out, just north of the open pit, to the town of Tolar Grande. The mine access road will need to be built to Argentinian standards for gravel roads with an approximate width of six meters and one meter wide shoulders. The majority of the road in this section crosses the Salar de Arizaro. To reduce differential settlement from vehicle traffic crossing the salar, a half-meter of structural fill will need to be placed on the prepared salt crust of the salar and topped with 200 mm of road base.

18.2.2 Site Roads

Three types of site roads were envisioned for the Project: out-of-pit haul roads, high traffic roads, and a light vehicle maintenance road.

The out-of pit haul roads would include roads from the pit to the WRSFs, the crusher, the truck shop, and the concentrator feed stockpile (Figure 18-3). There are approximately eight km of out- of–pit haul roads and these need to be designed for two-way mixed haul truck traffic.

Similarly, the high traffic infrastructure-connecting roads will need to join the truck shop to the concentrator and the concentrator to the rail load-out facility. It has been assumed there will be a high volume of light to heavy vehicle traffic, typically less than 40 t, traveling these roads. There are approximately 4 km of infrastructure-connecting roads that have been designed for two-way mixed traffic, excluding haul trucks.

It is planned to have one low traffic maintenance road for light vehicles around the TSF, a distance of 18 km. This road will access the cyclone and water reclaim facilities and allow one-way traffic access to any point around the TSF.

The concentrator facilities are assumed to have a number of internal roads to transport materials and personnel between facilities.

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Figure 18-3: Out of Pit Haul Roads (Ausenco, 2013)

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18.3 Power Supply

Hugo Gil Figueroa & Asociados (Hugo Gil) of Buenos Aires completed a preliminary power supply study in July 2011 and an update to this study was prepared in January 2013 using revised estimated electrical loads and schedule dates. The updated study addresses the below-listed items, which are discussed further in the following sections:

 Confirm the power supply source;  Confirm the tap point on the existing transmission line;  Confirm the voltage of the new transmission line that will supply power to the Project;  Confirm the routing and length of the new transmission line;  Update the estimated capital cost to complete the new transmission line, including the switching station at Olacapato and the main substation at the Project; and,  Update the estimated energy supply and transmission cost per-kilowatt-hour (kWh), and the transmission line operation and maintenance cost; both to be used in developing Project capital and operation costs for use in evaluating potential Project economics

18.3.1 Power Supply Source, Tap Point, Routing, and Length

The 345 kV transmission line from the Guemes generating station in Salta to Chile will be tapped at Olacapato where a switching station will be installed. The prospective routing of the transmission line from Olacapato to the Project is shown below in Figure 18-4. The overall transmission line length is 144 km.

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Railway route

Proposed line route

Figure 18-4: Transmission Line Routing from Olacapato to Taca Taca (Hugo Gil, 2013)

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18.3.2 Transmission Line Voltage

The existing Guemes-Chile transmission line is 345 kV. The transmission voltage from Olacapato to the Project is anticipated to be 345 kV.

18.3.3 Estimated Costs

The estimated capital cost of the power supply system is $146,806,000, which includes all site civil, earthworks, and other miscellaneous items for a complete, functioning switch station and main substation. This cost is included in the Project capital costs described in Item 21.0 of this Technical Report.

The estimated average cost to supply power to the Project is $0.085 kWh (rate includes estimated line losses). This unit cost has been used in estimating mining, processing, and General and Administrative (G&A) operating costs by applying this unit rate to estimated power consumption by each area.

Annual switching station, main substation, and transmission line operating and maintenance costs are estimated to be $1,468,000. This cost has been included as a component of G&A operating costs.

18.4 Site Electrical Power Distribution

The Taca Taca main substation is located directly adjacent to the main concentrator building. From the Taca Taca main substation, medium voltage electrical power will be distributed to the concentrator and the rest of the site at either 33 kV or 13.8 kV depending on the length of cable runs or the end user demand.

Electrical distribution to the Project areas such as the mine ring main, permanent camp, and water supply stations is envisioned to be on 33 kV overhead transmission lines run along poles. The longest transmission line is expected to be 50 km to the furthest wellfield. The rest of the distribution, mainly to the concentrator electrical rooms as well as ancillary areas such as the administration building, truck shop, tailings water return, rail car loading, and desalination plant will be on 13.8 kV transmission lines.

An allowance for three 1.25 MVA diesel emergency generators on skids complete with fuel tank switchgear, voltage monitors, and switchroom has been provided.

18.5 Tailings Storage Facility

18.5.1 Introduction

The mineral extraction processes for the Project are planned to be grinding and flotation that produce copper and molybdenum concentrates and tailings. The flotation tailings should be chemically benign and are anticipated to be non-acid generating based on preliminary acid based accounting (ABA) testing by Plenge.

The ultimate tonnes of tailings anticipated to be stored in a containment facility are approximately 1.792 Bt. Various disposal options and configurations were considered for the siting of the TSF. Based on field reconnaissance, Ausenco recommended an area northwest of the pit, which has a

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small impounding area as shown in Figure 18.10. This site is composed of sandy alluvium of variable thickness overlaying sedimentary bedrock, with some weathered outcropping.

18.5.2 Tailings Storage Facility Design

Tailings are anticipated to be discharged into the TSF as a thick slurry, a mixture of finely ground rock and water. Preliminary physical characterization of tailings has been performed in Ausenco’s laboratory in Peru. The laboratory test work showed the whole tailings to have a settled density of approximately 1.3 t/m3 and a maximum consolidated density of 1.6 t/m3. For the PEA design, conservative values of 1.2 and 1.4 t/m3, respectively, were used. For more efficient water recycling, high capacity thickeners are anticipated to be utilized to increase the solids content to approximately 55% solids in the waste stream going to the TSF.

A siting study was performed that included considerations of centerline and downstream embankment construction along with construction materials that included waste rock and coarse tailings from the cyclone underflow. Due to the low storage efficiency of the dam (tailings storage/embankment volume), a centerline cyclone tailings dam was selected as the best option.

The typical construction of a sand dam requires there to be sufficient volume of coarse sand from the cyclone underflow in order to construct the dam in parallel with impounding the balance of the tailings. A rockfill starter dam is planned to ensure there is sufficient capacity for the fine fraction of the cyclone tailings during normal operation and whole tailings overflow during construction of the sand dam. The starter dam is anticipated to be approximately 55 m high with a crest elevation of 3,585 masl and an impoundment footprint of 254 ha. Both the upstream and downstream faces of the starter embankment are envisioned to be constructed with 2:1 (H:V) slopes. The upstream face of the starter dam is expected to be lined with a low permeability soil liner and geomembrane to ensure initial containment of tailings and free water until the tailings beach is established and the decant pool is pushed to the back of the impoundment. Between the rock fill and soil liner zones of the dam, a series of granular filters will be installed to prevent piping of materials through the dam if there is a leak in the geomembrane liner. The starter facility has a capacity of approximately 63 Mt of tailings including 3 m of freeboard.

The tailings dam is planned to be raised using coarse cyclone tailings and centerline construction methods using a hydraulic jacking header system and a second jacking system is envisioned to be used to distribute the cyclone overflow (fine fraction and whole tailings) into the impoundment area. In addition, the centerline method is able to accommodate the anticipated rapid rate of rise of the tailings surface expected during the early years of the proposed Project production schedule. The ultimate height of the tailings dam is expected to be approximately 175 m with a crest elevation of 3,696 masl and an impoundment footprint of 1,754 ha.

18.5.3 Cyclone Station

The sands from the tailings stream are planned to be utilized in the construction of the tailings dam. To separate the sand, a cyclone station is envisioned to be located above the TSF at an elevation of approximately 3,720 masl. The purpose of the station is to separate the coarser tailings from the whole tailings. The starter cyclone facility has capacity for 120,000 t/d and in Year 8 an additional cyclone station is planned to accommodate an additional 60,000 t/d. For the PEA, a two-stage cyclone system with approximately 20% underflow efficiency and an average cyclone utilization time of 85% has been assumed. The tailings from the concentrator are envisioned to be pumped to the cyclone station. In the initial years, the underflow and overflow are planned to be transported to the jacking header tailings distribution systems by gravity. A pumping system has been planned in the later years.

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18.5.4 Tailings Delivery System

The selection of a jacking header system to accommodate efficient raising of the underflow and overflow header pipelines along the embankment crest was a critical design decision because of the rapid raise rates in the first few years of operation. The jacking header consists of a steel frame that is mounted on steel posts on which it can freely slide to accommodate vertical raising by a system of jacks. One jack is located on each set of two posts (Figure 18-5). The underflow is discharged from the underflow delivery line. Raising the jacking header is coordinated with the embankment rate of rise and the deposition sequence. Sufficient clearance needs to be maintained for compaction equipment to move beneath it and compact the underflow sand.

The underflow from the jacking header would be discharged on the downstream side of the embankment crest. During initial operations the underflow will be discharged at the base of the facility and equipment will be used to compact and build a 3:1 (H:V) slope. Once the slope reaches the starter dam crest the underflow will be placed in compacted thin lifts down the slope to prevent liquefaction during a seismic event.

Figure 18-5: Tailings Discharge System – Jacking Headers

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18.5.5 Underdrain System

A drainage network underlies the sand portion of the embankment, which is a key component of the design. The drain facilitates drainage of the cyclone tailing sands, maintaining a low phreatic surface in the lower section of the embankment. Water that drains from the underflow sands is collected and returned to the process plant for reuse. The embankment is raised in lifts of compacted cycloned tailing sands while filling the impoundment. As the embankment is raised in height, its footprint expands downstream necessitating phased expansion of the underdrain system.

18.6 Reclaim Water System

Tailings water is planned to be reclaimed from the TSF from a natural ponding location that is expected to be on the western edge of the facility. Tailings reclaim water at a rate of approximately 960 m3/h in the initial Years 1 to 7 is planned to be pumped from the pond and sent via pipeline approximately 9 km around the periphery of the TSF to the process water tank and the concentrator. It is anticipated that there will be a significant quantity of water that is lost to interstitial water retained in the settled tailings, as well as water lost to evaporation and seepage from the TSF. This has been accounted for in the Project water balance.

18.7 Hydrogeology and Raw Water Supply Sources

During 2011 and 2012 Ausenco’s hydrogeologists conducted an extensive hydrogeological investigation to identify and characterize potential water supply sources for the Project, and to quantify potential pit dewatering requirements (Ausenco, 2012). The methodology and results from the investigation are summarized below.

The climate of the western half of Salta Province is very arid with a very low average annual precipitation of 98.4 mm/y and very high annual potential evaporation rate of 2,253 mm/y (calculated). No perennial surface waters exist at the site. The very arid climate results in no sources of sustained runoff and any groundwater discharge tends to evaporate quickly. However, geomorphic evidence in the small mountains west of the site suggest runoff events do occur occasionally for reported periods of 10 to 15 min during the months of January and February.

18.7.1 Hydrogeology

Following a review of geologic mapping and satellite and ground reconnaissance, prospective water supply areas were identified. The areas of interest, located both north of the Project and further west and northwest, were selected because they are outwash deposits and alluvial fans.

Ausenco supervised the installation of 18 2-inch diameter Schedule 80 PVC standpipe piezometers using RC and mud rotary drilling methods. Slug tests were conducted at six piezometers.

Groundwater quality data indicate that there are essentially three main groups of water represented by samples from all the piezometers.

 Group I comprises fresh groundwater and may be classified as a sodium bicarbonate NaHCO3 type.  Group II is comprised of brackish groundwater and is a sodium chloride type that is similar to the water at the Salar de Arizaro.  Group III is represented by fresh groundwater and is a sodium chloride type but shows higher sulfate and magnesium concentrations than water from Group II.

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The direction of groundwater flow in Lumina Copper’s northern concessions is towards the Salar de Taca Taca from the Aracar Volcano. The phreatic water level surface in this area ranges from approximately 36-120 m below ground (bg) dependent upon surface topography.

Ausenco utilized two empirical methods to estimate the available groundwater upgradient of the Salar de Arizaro: the Darcy flow calculation and the mass balance approach. Both methods supported the presence of a substantial mixing zone of fresh and brackish water upgradient of the salar. This mixing zone is the planned source of raw water feed to the water treatment plant.

18.7.2 Pit Dewatering

Multiple 2-inch diameter piezometers and two 8-inch diameter test wells were installed in the Taca Taca pit area using RC and mud rotary drilling methods. Both test wells were installed to conduct aquifer pumping tests to determine aquifer properties of hydraulic conductivity and storativity.

A groundwater contour map of static water levels in the pit area indicates that the phreatic groundwater surface generally follows surface topography and that the direction of groundwater flow is towards the southeast, towards the Salar de Arizaro. This indicates the area of groundwater discharge is most likely the salar. Groundwater samples collected from stand-pipe piezometers within the mine pit area indicated that the groundwater is similar to the sodium chloride brine of the Salar de Arizaro. The groundwater has a high TDS ranging from 253,000 to 310,000 mg/l and is similar to the brine water sampled from the nearby Salar de Arizaro, which has a TDS of 255,500 mg/l.

Two predictive models were run for the mining phase of the Project, a passive dewatering scenario (free draining pit wall surface) and an active dewatering scenario. The active dewatering scenario included a combination of vertical interceptor wells installed around the circumference of the pit and horizontal drains installed in the pit walls. Active dewatering simulations indicated that the predicted average dewatering rates ranged from 26 l/s to 119 l/s, with the greatest flows entering the pit during the first years of mining and decreasing for successive years. The predicted effects of an aggressive pit depressurization program using wells and horizontal drains indicate that pore pressures in and around the proposed pit slopes could potentially be significantly reduced.

Based on the results of the groundwater modeling, costs were developed for the installation of vertical dewatering wells and installation of horizontal drains in the pit walls. These costs are included in the capital and operating cost sections, discussed in Item 21.0 of this PEA.

18.7.3 Raw Water Supply

It is envisioned that water could be provided to the concentrator from wellfields located at a minimum of three locations in the region, likely, but not definitely, including the following:

1. 50 km to the southwest of the Project

2. 25 km to the northwest of the Project

3. 20 km to the west of the Project

Additionally, brine water from the pit dewatering and the nearby salar is anticipated to be utilized.

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It has been assumed that water with varying amounts of salinity and TDS will be pumped from multiple wells at each location to a common collection tank at each site, before being pumped to the concentrator. Power will be provided to each water supply site by overhead transmission lines.

Some of the raw water will require treatment and will be pumped directly to the water treatment plant located adjacent to the concentrator.

18.8 Process Water Supply

18.8.1 Introduction

Schlumberger Water Services USA Inc. (Schlumberger) completed a scoping level assessment and design of a water treatment plant to meet Project make-up water requirements (Schlumberger, 2013).

18.8.2 Make-up Water Requirements

The process water make-up requirements for the Project at 120,000 t/d throughput are estimated to be 2,100 m3/h (approximately 600 l/s) of untreated brine water and 800 m3/h (approximately 200 l/s) of fresh water. The water quality of the fresh water make-up under the PEA requires a TDS concentration below 1,500 mg/l. This fresh water is anticipated to serve the purposes of cleaner flotation, copper molybdenum separation flotation, gland seal water, cooling, reagent mixing, and concentrate washing. An acceptable TDS of the brine process stream component is estimated to be in the range of 200,000 mg/l to 300,000 mg/l, as determined by preliminary metallurgical testing, as described previously in Item 13.4.5. The brine process water component will satisfy requirements associated with grinding, rougher flotation, tailings losses (evaporation, seepage, and retention), and contingency. At 180,000 t/d throughput in Year 8, the total estimated demand is expected to increase to approximately 2,900 m3/h.

18.8.3 Quantity and Quality of Available Water Resources

The estimated water treatment requirements for provision of 800 m3/h of fresh water make-up were evaluated through examination of quantity and quality of available groundwater resources. The objective of this evaluation was to minimize the flow rate and TDS of the raw water stream entering this treatment system and to provide as much of the make-up water through blending of the available fresh groundwater sources in the Project area, as possible. The application of conventional desalination equipment requires a raw water feed TDS to be below 50,000 mg/l (Guyer, 2010). Based on scoping level groundwater investigations, water sampling results and estimates of well and aquifer yield, Ausenco identified the following potential estimated yields and qualities of raw water, grouped by TDS ranges (Ausenco, 2013):

 600 m3/h of 250,000 – 300,000 mg/l TDS  150 m3/h of 10,700 mg/l TDS  90 m3/h of 440 – 830 mg/l TDS

Given the available data, the proposed water blending and treatment scenario to provide the Project make-up water requirement is presented in Table 18-1 and Table 18-2. Hence, to meet the fresh water make-up quality requirements, groundwater derived from the mixing zones will be the only source requiring treatment.

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Table 18-1: Components of 600 l/s of Untreated Brine Water Make-Up

Source Proposed Flow (l/s) TDS (mg/l)

Salar de Arizaro 460 250,000

Pit dewatering discharge 50 250,000

Concentrate from desalination 90 23,350

Total composite brine water 600 216,000

Table 18-2: Components of 200 l/s of Fresh Water Make-Up

Source Proposed Flow (l/s) TDS (mg/l)

Wellfield No. 1 (50 km southwest) 30 440

Wellfield No. 2 (25 km northwest) 30 830

Wellfield No. 3 (20 km west) 30 635

Mixing Zones (post-desalination) 110 1,815

Total composite fresh water 200 1,285

The above water treatment and blending scenario accounts for a maximum potential desalination system efficiency of 55%, which requires 200 l/s of groundwater to be derived from the mixing zones to produce 110 l/s of treated fresh water.

Therefore, because the estimated yield from the mixing zones has not been fully proven, it is uncertain that the desalination system raw water feed rate can be met on a sustained basis. To help reduce the treated water demand (i.e., mixing zone water demand) additional yield from the fresh sources would be required. For instance, if a total of 150 l/s can be derived from the three proposed wellfields combined (the upper estimated ranges of 10-50 l/s for each), the water treatment requirement would be reduced to 50 l/s, thus lowering the desalination raw water feed from the mixing zones to 90 l/s.

For the purposes of the PEA, the water treatment design and costing was kept conservative in terms of process equipment sizing. Therefore, it was assumed that the 200 l/s mixing zone raw water demand could be met.

18.8.4 Proposed Water Treatment System

The proposed water treatment system is anticipated to consist of the following components:

 Stabilization ponds  Prechlorination dosing systems

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 Multi-media filtration systems  Dechlorination systems  Antiscalant dosing systems  Reverse osmosis systems

To facilitate the use of conventional reverse osmosis membrane technology, the treatment system is assumed to consist of four treatment modules operating in parallel. Two additional modules are envisioned to be required as standby units to allow cycling of modules as part of the system maintenance program.

18.8.5 Projected Water Quality of Treated Process and Waste Streams

The estimated water quality of the desalination system permeate is provided in Table 18-3. The water quality projections were made utilizing Hydranautics membrane Solutions Design Software, v 2012.

Table 18-3: Estimated Permeate and Concentrate Chemistry

Parameter Permeate (mg/l) Concentrate (mg/l) Calcium 6 368 Magnesium 3 179 Sodium 689 8313 Potassium 16 152 Sulfate 18 667 Chloride 1,081 13,631 Fluoride 0.5 2.2 Boron 0.6 0.8 Silica 1.7 33 TDS 1,816 23,347 pH 8.17 8.44

18.8.6 Disposal of Waste Streams

The waste stream generated by the desalination system is anticipated to include the concentrate stream as the primary component, with secondary waste streams consisting of backwash water from the multi-media filtration system and backwash water from membrane system maintenance. Because the water quality of these waste streams is anticipated to have a relatively low TDS concentration, it is assumed they will be blended with the mine dewatering discharge and/or groundwater derived from Salar de Arizaro to satisfy the brine make-up water supply requirement.

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18.8.7 System Operating and Maintenance Requirements

The operation of the desalination system is estimated to require 1.5 full time staff to conduct monitoring, prepare reagents, and conduct routine maintenance.

Operation of the primary components of the desalination plant is planned to be automated, including backwashing of the multi-media filtration system. However, system adjustments to maintain optimal performance may be required based on monitoring of parameters including flow, pressure, and water quality.

The total power requirement for operating the desalination system is estimated to be 777 kW.

18.8.8 Capital and Operating Costs

The total capital cost of the desalination plant is estimated to be $5,276,000, including engineering, freight, installation, and commissioning. The annual operating cost of the system including labor is estimated to be $1,089,000, which is equivalent to $0.31 per cubic meter of fresh-make up water produced. The operating cost is based on a power cost of $0.085 kWh. Both capital and operating cost estimates are +/-25% of actual costs. These costs have been utilized in the PEA cash flow model and economic analysis.

18.9 Potable Water

An allowance has been made for a potable water system including an HDPE piping network throughout the Project including the concentrator and ancillary buildings. A total demand of 20 m3/h has been assumed, based on potable demands for similar sized plants with similar staffing requirements. It is anticipated that a pre-engineered packaged system will be utilized.

18.10 Logistics and Transportation

18.10.1 Transport Requirements

During construction, permanent plant materials and equipment will need to be transported from their points of origin to the Project. During operation of the mine and mineral processing plant, consumables and supplies, such as fuel, explosives, process reagents, equipment spare parts, and food will need to be transported to the Project.

During both construction and operation, workers will need to be transported between the Project and their home bases, assumed to be Salta or elsewhere in Argentina, for scheduled rest and relaxation (R&R) breaks.

Final mineral processing products, bulk copper concentrate and molybdenum concentrate packaged in one tonne bulk bags and containerized, are assumed to require transport to their final markets in Asia and Chile, respectively.

18.10.2 Background

In order to address the transport options to meet the needs described above, the following studies were completed:

 Transport Survey: Taca Taca Mine, Transportes Universales S.A., June 30, 2011.

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 Taca Taca Mine Access Road Photographic Survey, Ausenco, December 18, 2013.  Taca Taca Project Report Addendum to the Transport Survey: Taca Taca Mine, Transportes Universales S.A., August 19, 2011.  Logistic Preliminary Report for Taca Taca Project – Argentina: Ports, Airports & Road Restrictions in Chile, Agility Logistics Chile S.A.  Taca Taca Project Rail Study 2012/2013 Scoping Study, TFP Construcciones SRL, March 2013.

18.10.3 Argentinian Port Facilities

Two ports were considered: the Buenos Aires Port and the Zarate Port. The Buenos Aires Port was dismissed by Transportes Universales Logistica Integral (TUSA), primarily because it is located in the center of the city and is more expensive than the other option considered, Zarate Port.

The Zarate Port, operated by Terminal Zarate S.A., is located along the right riverbank of the Parana River at km 110, approximately 100 km upriver from Buenos Aires. The natural draft at the 245 m long pier is 35 ft. It also has mooring towers at 50 m, allowing use by multi-purpose vessels such as container carriers, car carriers, or general cargo vessels up to 270 m long.

The Zarate Port occupies an area of 120 ha, with ample container storage area, 8,900 m2 of warehouse storage, and substantial land for expansion.

Two portainers are each capable of unloading 25-30 containers per hour, with maximum lifts of 40 tons at 36 m horizontal distance. Heavier lifts, and break bulk items, can be offloaded using the vessel’s cranes, if available, or by a mobile crane located on the pier.

The port also has very good road access, located just 9 km from the Pan American highway (National Route 9) and access to the rail transfer terminal located only 1 km from the port.

18.10.4 Road and Rail Transport from within Argentina to Taca Taca

The vast majority of items requiring transport to Taca Taca would either originate at the Zarate Port or from the greater Buenos Aires metropolitan area.

While rail transport is technically possible, TUSA has concluded that it is not the best alternative. Rail transport in Argentina has historically been unreliable and expensive.

For discussion purposes, road transport is divided into two segments: Zarate Port to Salta and Salta to Taca Taca.

 Zarate Port to Salta Transport times are expected to range from three days for conventional trailers (less than 5 m wide and 5 m high), to 5-6 days for oversized loads requiring detours and temporary raising of utility lines, to 7-8 days for overweight loads (greater than 100 tons) requiring detours around weight restricted bridges.  Salta to Taca Taca Two options exist for transport between Salta and Taca Taca, one for conventional loads and one for oversize or overweight loads.

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Conventional cargo may travel from Salta to San Antonio de los Cobres via Provincial Route 51, on to Cauchari via Route 51, and then to Taca Taca via Route 27. Oversize and overweight loads will need to utilize Routes 8, 52, and 40 to reach San Antonio de los Cobres. The remainder of the trip would be the same as for conventional cargo. One segment of the existing route between Cauchari and Tolar Grande, known locally as “the Seven Curves”, will be preserved in its current state for cultural/social reasons. A bypass around this area, planned for construction during Year -4 in the Project schedule, has been included in the Project capital cost estimate.

18.10.5 Chilean Port Facilities

Agility Logistics completed an evaluation of the three ports in the Antofagasta area, the major points of which are summarized below:

 Antofagasta Terminal International (ATI) ATI is a privately operated terminal with four berths ranging from 130-220 m in length and drafts of 9.1 to 11.6 m. Surface area totals 8.5 ha. Cargo can be offloaded using ATI’s two 100 MT cranes, or vessels’ onboard cranes. ATI receives break bulk and containerized cargo via regularly scheduled port calls by numerous shipping companies, both coastal and transoceanic. ATI also serves a large segment of the northern Chile mining industry by storing, handling, and loading mineral products, such as copper concentrates. The drawback to the ATI port is its location in the center of Antofagasta. Agility recommends ATI for container freight only, even then with an area outside the city for staging ongoing transport to Taca Taca to avoid demurrage charges, and to store empty containers on return. Oversize and/or overweight freight would present difficulties exiting the port and transiting the city due to narrow streets, overhead utilities, and traffic. Rail transport is not available.  Empresa Portuaria de Antofagasta (EPA) EPA is a public wharf with three berths ranging from 180-210 m in length and drafts from 9.14 to 10.0 m. Surface area is only about 4,000 m2. EPA has no cranes for heavy lift or break bulk project cargo, instead relying on vessels’ cranes, if available. Due to EPA’s location (same as ATI), it has the same issues mentioned above regarding congestion, difficult egress from the city, etc.  Angamos Port – Mejillones Mejillones is located 65 km north of Antofagasta and is accessed by excellent paved roads, as well as rail. For road transport between Mejillones and the Project, it is not necessary to traverse the city of Antofagasta. The berths are sheltered from the effects of bad weather by its natural bay and a 192 m long breakwater which has been constructed parallel to the shore. Mejillones has four berths, ranging in length from 180-225 m with drafts from 10.7 to 12.5 m. Offshore depths facilitate access to the berths without undue maneuvering by vessels. Mejillones receives regularly scheduled port calls by liners providing break bulk, container, and bulk carrier services.

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Unlike both ATI and EPA, Mejillones has more than adequate areas for temporary storage of offloaded freight, including enclosed warehouses and empty containers. Additionally, Mejillones has substantial area for expansion to facilitate storage of bulk concentrates.

18.10.6 Transport Chilean Port to Taca Taca

Depending on availability within Argentina and pricing, certain bulk consumables for operation may be procured in Chile or purchased elsewhere and delivered to a Chilean port. All, or a portion, of these consumables may be transported by rail by adding tank or flatbed rail cars to the empty concentrate trains returning to the Project. The rail option will be discussed in Section 18.11.4.

Permanent plant equipment and materials for development and construction of mining and mineral processing facilities at the Project will likely be purchased worldwide. For equipment or materials originating in Asia or the west coast of the United States, offloading from ships will likely occur at Mejillones, north of Antofagasta.

TUSA identified two potential truck transport routes, depending on Chilean customs requirements, although subsequent discussions with Chilean officials have indicated that customs clearance at the Project may be possible. Access from the Project to Paso Sico is shown below in Figure 18-6 and the prospective routes, shown below in Figure 18-7 and Figure 18-8 are:

 Antofagasta - Paso Sico (via Lacos, Socaine, and Peine) – Taca Taca (via Route 51 and Route 27)  Antofagasta – Calama – San Pedro de Atacama – Paso Sico – Taca Taca (via Route 51 and Route 27)

The first route described above is approximately 200 km shorter than the second route. Travel by both of these routes would require approximately two days.

Figure 18-6: Exit from the Project along Ruta 27 and Access to Ruta 51 to Paso Sico (TUSA, 2011)

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Figure 18-7: Travel from Paso Sico to Antofagasta through Lacos, Socaine, and Peine (TUSA, 2011)

Figure 18-8: Travel from Paso Sico, to San Pedro de Atacama and Calama, arriving to Antofagasta (TUSA, 2011)

18.10.7 Transport of Workers

The Project is located approximately 230 km west of the city of Salta, and approximately 400 km by road. During both construction and operation, workers will be transported between Salta and the Project by road at specified intervals for their R&R. The travel time is approximately 8 hours, depending on vehicle type, stops, and road conditions.

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If commercial rail operations between the Project and Salta are re-established by a rail operating company in the future, employee transport could be by rail.

During the construction period, construction contractors will bear the responsibility for transporting their workers between the Project and Salta. For the Project’s permanent operations workforce, light vehicles, including buses, vans, and pickups with trained drivers, have been included in estimated capital and operating costs for employee transport. In both cases, workers whose home bases are not in Salta have been assumed to travel beyond Salta at their own expense.

18.11 Concentrate Storage, Handling, and Transport

18.11.1 Introduction

Throughout the 28-year life of the mine, the Project is estimated to produce approximately 21.6 dmt of copper concentrates and approximately 228,000 dmt of molybdenum concentrates. Annual totals of copper concentrate are estimated to range from 511,782 dmt in the first year, to peaks of 1,025,680 dmt and 1,094,594 dmt in Years 2 and 10, respectively.

Molybdenum concentrates are estimated to total 4,318 dmt in the first year, and then range between 13,484 dmt and 3,793 dmt in Years 10 and 16, respectively.

The PEA contemplates bulk copper concentrates will be transported by rail to the Port of Mejillones, located approximately 65 km north of Antofagasta, Chile. Figure 18-9 shows the route from Taca Taca to the Chilean border at Socompa, and on to the Port of Mejillones.

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Figure 18-9: Map of Railway Routes between Taca Taca and Mejillones (TFP, 2013)

Copper concentrates are envisioned to be stored while awaiting regularly scheduled vessels for ongoing shipment to smelters in Asia.

Molybdenum concentrates are anticipated to be packaged in one tonne bulk bags, loaded into sea containers, and also transported by rail to the Port of Mejillones. At Mejillones, the molybdenum concentrates may be transferred to trucks for the final leg of transport to a purchaser in Chile.

The infrastructure associated with onsite transportation, onsite storage and loading, train operation and unloading, storage and ship loading at Mejillones is described below.

18.11.2 Onsite Transportation

After thickening and filtering to achieve a moisture content of approximately 9% (w/w), copper concentrate is planned to be temporarily stockpiled in the copper concentrate handling building, located adjacent to the concentrator. An FEL and five 20 t trucks are planned for transporting the concentrate to the railcar load-out facility, located approximately 2.5 km to the northeast. Refer to Figure 18-10 for the relative locations of the facilities.

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Figure 18-10: Site Layout

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18.11.3 Onsite Storage and Loading

The PEA contemplates that copper concentrate will be stockpiled in a concrete-walled enclosed storage area. An FEL is envisioned to maintain the stockpile, as well as transfer the concentrate to a feed hopper on the tail end of a transfer conveyor. The transfer conveyor is planned to convey the concentrate to an elevated load-out bin which straddles the rail spur. The load-out bin is envisioned to have three hydraulically operated clamshell-type gates through which the concentrate would flow into the railcar being loaded. The gates, operated by an operator manning a control panel in an enclosed control room located adjacent to the load-out bin, are envisioned to be cycled in a sequence to equally distribute the concentrate in the railcar. Railcar weight during the loading process is planned to be monitored by track mounted load cells, allowing the operator to maximize loading without exceeding rail system limits. Indexing of empty cars into loading position would occur automatically, allowing a 28-car train to be loaded with 48.5 wmt of copper concentrate per car in a total elapsed time of 1 h and 10 min.

Additional time may be required for inspection of railcars (allowed at 2 h and 20 min) and train formation (estimated at 1 h). Up to four hours have been allowed for minor repairs, as required.

18.11.4 Evaluation of Existing Rail Infrastructure and Rolling Stock

TFP Construcciones SRL (TFP) completed a rail study to:

 Evaluate the condition of the existing rail line between Salta and Socompa, located on the border with Chile, a distance of 571.1 km.  Evaluate the extent and condition of Belgrano Cargas S.A.’s (the current rail operator) rolling stock.  Estimate the cost to rehabilitate the rail system, particularly from the Project to Socompa, to a condition which would support the frequency and weight of the anticipated trains to meet the Project’s needs.  Assess the potential of rehabilitating the abandoned maintenance facilities located in Tolar Grande for the Project’s use, or, alternatively, estimate the cost to construct new rolling stock maintenance facilities at the Project.  Evaluate the assessment of the Chilean rail segment (394 km of the total 529 km distance from the Project to Mejillones), by the Chilean rail company, Ferrocarril de Antofagasta (FCAB), and incorporate appropriate information into TFP’s overall evaluation, cost estimates, and recommendations.  Identify potential commercial options for rail transport between the Project and Mejillones, and recommend the most favorable option.

Major conclusions of TFP’s report are briefly summarized below:

 Substantial rehabilitation of the existing rail line is required for the service planned. Scope was developed and costs estimated for specific sections of the track between the Project and Socompa.  Belgrano Cargas’ existing rolling stock is old, in poor condition, and is not well-suited for the Project application.

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 While the Tolar Grande abandoned maintenance and repair facilities could be rehabilitated, the result would not be as functional as a new facility at the Project. Obtaining the right to use the facility from Belgrano Cargas also may not be possible.  Currently, it does not appear that Belgrano Cargas will be able to provide contracted third party rail services to meet the Project’s currently anticipated needs.  Three independent rail companies own portions of the rail lines between the Project and Mejillones: Belgrano Cargas – the Argentinian section; Ferronor – the initial segment within Chile; and FCAB – the final segment arriving at Mejillones. Rather than operation by three independent rail companies and the coordination difficulties and inefficiencies that would certainly result, TFP recommends a single operator for the entire trip.  There is no existing, or even potential, independent third party operator capable of providing the complete freight service needs.  Primarily for the preceding reasons, but also for better control of product shipment and cost effectiveness, TFP recommends purchase of new rolling stock (engines and railcars) by Lumina Copper and operation of its own trains between the Project and Mejillones.

18.11.5 Train Components, System Description and Operation Equipment (Rolling Stock)

The initial capital and sustaining capital cost estimates discussed in Item 21.0 of this PEA, include purchase of the following new rolling stock in the Project years indicated:

Item Number Units Project Year

Initial 8 -1 Locomotives

Additional 7 1 Locomotives

Initial Railcars 139 -1

Additional 139 1 Railcars

18.11.6 Infrastructure

New or rehabilitated infrastructure included in Project initial capital and sustaining capital estimates includes the following items:

 Construction of a new rail spur, sidings (3,465 m of rail in total) and associated switches to connect the Project concentrate load-out facility to the existing main line in project Year -1.  Construction of a new maintenance and repair facility for locomotives and railcars located adjacent to the concentrate load-out facility and rail spur. This facility would be constructed in project Year -1.  Rehabilitation of 40.5 km of track between the Project and Socompa which have been identified as requiring upgrading to meet the frequency and loading of Taca Taca rail traffic. This work would be completed by the Project, in equal increments over five years (Project Year -1 through Year 4). The cost in Year -1 has been included in initial capital, with the remaining four years included in sustaining capital.

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18.11.7 System Description and Operation

Each train is planned to consist of 28 wagons (railcars), each loaded with 48.5 wmt of copper concentrate. Each train would therefore transport 1,358 wmt. Wagons are covered to meet environmental requirements, to avoid loss of concentrates, and to avoid introduction of moisture during winter months in the mountain passes.

Locomotives could be added or removed from trains at different points along the route, to either the head or tail of the train, to provide additional power or braking capacity, respectively. Of the 15 total locomotives, 14 are planned to be active at all times, with one spare unit.

18.11.8 Operating Schedule

Highlights of the system operation are:

 Trains are scheduled to depart the Project approximately every 8 h throughout the year, including Sundays and holidays.  15 d/y have been allowed for rail service interruption due to severe weather (blizzards in the mountains) or landslides onto the tracks.  During peak concentrate production in operating Years 2 and 10, 975 trains per year would run.  As currently scheduled, each train would take 20 h to reach Mejillones and an additional 20 h to return. With unloading and an allowance for miscellaneous repairs, if required, a train will complete one cycle approximately every three days.

Rolling stock, track, and spur maintenance, are included in annual operating costs, as detailed in Section 21.15 of this PEA.

18.11.9 Unloading, Storage, and Ship Loading

Lumina Copper management and MTB staff toured the Mejillones port facilities in August 2012 and met with port management personnel. During the visit, Mejillones management informed the visitors that Mejillones has a completed design and plans to complete additional infrastructure, which could be used to unload railcars, temporarily warehouse, and load vessels with Project concentrates. This information was reconfirmed by Mejillones when recontacted by H&H in February 2013 to request port fees. Port fees subsequently provided were incorporated in the cash flow model described in Section 22.0 of this PEA.

Reportedly, similar railcar unloading operations have demonstrated unloading rates of approximately 500 tph.

Concentrate would be conveyed to a 50,000-70,000 wmt enclosed stockpile for temporary storage pending arrival of scheduled vessels. For vessel loading, concentrate would be transferred to a feed hopper for an overland conveyor, discharging to a telescoping (and perhaps slewing) shiploader. Depending on the shiploader’s characteristics, the vessel may or may not have to be repositioned for discharging to different holds. Loading rates will allow loading vessels in less than two to three days to avoid additional charges by the shipping line.

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18.12 Ancillary Site Buildings and Facilities

Various ancillary facilities would be located near the concentrator. The buildings and facilities include the following:

 Single story administration building  Assay lab and a small metallurgical lab facility  Changehouse for concentrator personnel  First aid or clinic building  Gatehouse at the entrance to site  Two-story concentrator warehouse and an attached workshop building  Potable water system  Sewage treatment plant

Mine ancillary facilities would be located near the primary crusher approximately 1 km west of the concentrator. These facilities include:

 Truck shop with eight truck bays and six bays for light vehicles. Prior to Year 8, an additional 4 truck bays are planned to be added. The facility includes two overhead cranes and a lube storage facility.  Two-story warehouse with offices  Tire shop  Maintenance and welding shop  Truck wash bay  Fuel storage depot  Effluent treatment facility  Mine explosives storage facility located approximately 1 km to the west of the ROM storage area. Commercially available packaged sewage treatment plants will need to be installed at the main concentrator facilities, the mine support facilities, and the camp. Effluent will need to meet Argentinian regulations for discharge, or undergo additional treatment until it does or will be recycled for reuse. Technical details will be developed during the next phase of Project development.

18.13 Permanent Camp

18.13.1 Introduction

Due to the remote location of the Project and the absence of any nearby infrastructure for providing accommodations and meals for the permanent workforce, plans call for construction of a permanent camp. The specific location for the camp has not been determined, but as a basis for the PEA, it has been assumed to be located within 10 to 12 km of the Project, easily accessible from the main access road.

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A leading Argentinian camp and catering contractor supplied separate proposals for designing, manufacturing, transporting, and erecting a modular camp; and for providing catering, camp maintenance, and cleaning services. 18.13.2 Basis for Sizing of Facilities

The facility has been sized for 752 inhabitants. The total permanent Taca Taca workforce is estimated to be 1,100 personnel, of which 21 would be based in Salta. One hundred and nine management, technical, and administrative personnel are assumed to work dayshift and return to their home base one week per four week period. These individuals’ accommodation units would be assigned to them, i.e., they will return to the same unit after their R&R.

Technical support staff, operators, mechanics, skilled craft labor, and laborers are expected to form four crews, with two crews onsite at any time working 12-hour shifts and the remaining two crews on rotation. The specific rotation period has not been established at this stage of the study, but is expected to be in the range of two to four weeks on and two to four weeks off. The exact rotation cycle is anticipated to be determined at a later stage of Project development, considering worker safety, travel distances/time to return workers to their home bases, and conditions necessary to attract and maintain a skilled and motivated workforce.

Two crews of shift workers, or 526 to 534 workers, are envisioned to be onsite and resident in the camp at any one time. The remainder of the available beds (approximately 59) has been included for visitors, such as Lumina Copper management, Salta-based employees, vendors, subcontractors, delivery truck drivers, train crews, etc. 18.13.3 Description of Facilities A fresh water supply and power supply are planned to be provided to the facility. Otherwise, the camp, consisting of the following components, would be self-sufficient:

 Refrigerated, frozen, perishable, and non-perishable food storage facilities  Fully equipped kitchen for food preparation  Common dining facilities, including furniture and supplies, to seat the scheduled workforce in two sittings  Recreational facilities  Laundry facilities  Potable water treatment plant, including storage  Waste water treatment plant  Sewage treatment  Communications system  Four levels of fully furnished and supplied accommodation units, as listed below, to be assigned based on job, work hours, and seniority. Table 18-4: Camp Arrangement Type Unit Beds per Room Bathroom Arrangements Number of Beds Private 1 Private 48 Semi-Private 2 Share by 2 rooms 128 Dormitory 2 32 rooms/Ablution Block 576

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19 Market Studies and Contracts

19.1 Introduction

H&H completed a marketing study for the Project. Using contacts with smelters and recent data from other projects, H&H analyzed the following marketing aspects in its report: supply and demand for both copper and molybdenum concentrates, including anticipated production from new projects in the construction or planning stages; smelter capacities, utilization, and the impacts on TCs, RCs, and payfors; ocean freight costs; and future metals pricing.

These marketing aspects are summarized below.

19.2 Supply and Demand

19.3 Smelter Capacities and Utilization

H&H’s forecasts for TCs, RCs, and payfors have been incorporated in the Project cash flow.

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19.4 Ocean Freight

19.5 Future Metals Pricing

After much analysis and discussion, H&H forecast the following metal prices for the Project as shown below in Table 19-1.

Table 19-1: Metal Price Forecast

Product Price Range/lb Recommend Price/lb

Copper $2.90 - $4.10 $3.10

Molybdenum $11.00 - $15.00 $13.00

Gold $1,350 - $1,900 $1,650

Silver $26.00 - $38.00 $35.00

In addition to the H&H forecast, consideration was given to a $2.89/lb consensus long-term price forecast made by 21 financial institutions (Thomson One Analytics, March 2013).

The Qualified Person of this section has reviewed the above pricing forecast studies and analyses and is satisfied that the results of these studies and analyses support the pricing forecast used in the economic analysis of the PEA. The PEA economic analysis utilizes the following more conservative pricing forecast (Table 19-2), which has been applied throughout the projected 28 year mine life of the Project.

Table 19-2: Project Metals Pricing

Product Price

Copper $2.75/lb

Molybdenum $12.00/lb

Gold $1,200/Toz

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20 Environmental Studies, Permitting, and Social or Community Impact

20.1 Environmental Baseline

Lumina Copper has initiated environmental baseline data collection to support exploration, development, and permitting for the Project. The intent of baseline data collection is to first and foremost document existing conditions at the Project. To the extent possible and available, baseline data will also include historical information from adjacent/nearby activities, operations, and other sources (i.e., governmental agencies). Lumina Copper will utilize this information to properly design the Project, develop environmental protections, design appropriate mitigation measures, and document compliance necessary to satisfy Argentinian national, provincial, and local regulatory authorities/requirements. Environmental baseline data will also be collected to compare against international lending (International Finance Corporation (IFC)/World Bank) standards and requirements to support Project financing options and alternatives.

By utilizing IFC/World Bank standards as a guide, Lumina Copper is underscoring the importance of managing environmental and social performance throughout the life of a Project. Doing so will aid in developing an effective Environmental and Social Management System (ESMS) that is recognized as a dynamic and continuous process that will be initiated and supported by management, and involves engagement between Lumina Copper, its workers, local communities directly affected by the Project (the Affected Communities) and, where appropriate, other stakeholders. The ESMS will entail a methodological approach to managing environmental and social risks and impacts in a structured way on an ongoing basis. A good ESMS appropriate to the nature and scale of the project will promote sound and sustainable environmental and social performance, and can lead to improved financial, social, and environmental outcomes.

To support this effort, Lumina Copper has contracted the team of Vector Argentina, members of an internationally recognized mining and environmental consulting firm, to design and implement the environmental baseline program.

The following environmental baseline disciplines will be collected and/or assessed:

Physical media:

 Soils and their uses  Geology (to be developed in coordination with the exploration team of Lumina Copper)  Geomorphology (including glacial studies, since there is new legislation on this topic)  Hydrology  Hydrogeology (program already started by Vector Argentina)  Basin water balance  Quality of water (surface and groundwater coordinated with hydrogeological study already started)  Meteorology

 Air quality (PM10, NOx, SOx, SH2, and CO2).

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 Landscape  Seismology  ARD (acid rock drainage potential analysis of waste rock, mineralized material, and tailings)

Biotic Media:

 Flora  Fauna  Limnology  Ecosystem characterization

Cultural Media:

 Socio-economic study  Archaeology  Paleontology

Environmental baseline data will be collected for a minimum of one year to include seasonal variation(s). Once baseline is fully established, and the site is permitted and constructed, Lumina Copper will perform compliance monitoring. The individual scope of these baselines is provided in the Taca Taca Permitting Handbook.

20.2 Permitting

The permitting requirements anticipated to develop and operate the Project were developed by Vector Argentina with input from Argentine legal counsel. Together, they developed the “Taca Taca Permitting Handbook”, which is intended as a guide to permitting. The overall permitting timeframe is expected to take 12-18 months.

20.3 Conceptual Closure

An essential part of the Project is the development of a closure plan that outlines activities for decommissioning, and mitigation of impacts during operation and once mining and processing activities have ceased. The preparation of a rehabilitation and closure strategy prior to the development of the Project is an integral part of the closure design process. This approach to project planning recognizes that mining represents a temporary use of the land and that appropriate closure of the operation is in line with the sustainable use of available resources.

The Taca Taca closure plan will focus on safety, stabilization of the land surfaces, post mine utilization of facilities and structures, and protection of the environment. Since the Project is located in an extreme arid, high altitude environment, revegetation is considered impractical and not conducive to the surrounding environment.

The following sections focus on specific closure plans/activities that are contemplated to be developed for the proposed mine, processing facilities, and associated infrastructure. To the extent possible, closure will be concurrent with the mining activities, with final closure occurring once mining ceases. Post closure monitoring will also take place to document and ensure proper reclamation of all impacted areas.

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The following sections present a description of the proposed closure activities that will be developed for the mine, overburden storage, mineral processing facilities, and infrastructure for the Project.

20.3.1 Mine Closure Plan

Mined-out areas of the open pit will not be reclaimed. However, Lumina Copper will evaluate opportunities in mine plan design for in-pit storage of waste rock material to minimize hauling. This opportunity will not be present until the later years of the mine life. Once the open pit has reached ultimate surface area, a safety berm and/or fencing will be installed to deter access into the open pit.

In advance of the open pit encountering the groundwater table, Lumina Copper will begin dewatering activities to allow for access and pit wall stability. This water will be incorporated into the overall water management plan to reduce water needs/pumping from distant sources.

At the end of mining, water will be allowed to fill back into the open pit and the newly formed pit lake will act as a hydraulic sink not requiring long-term maintenance.

20.3.2 Waste Rock (Uneconomic Material) Storage Facility Closure Plan

The main target of closure of the WRSFs is the physical and chemical stabilization of the waste rock materials and the rehabilitation of the habitat. The closure of the WRSFs will be initiated during operation (progressive closure), since one of the WRSFs will be completed before the end of the mine life.

Overburden facilities are planned to be built in lifts with an overall 2.5:1 outer slope. No grading of these facilities is proposed as the final slopes are intended to represent a stable land form. Preliminary assessment of waste rock material indicates that a portion of the material has the possibility to generate acid. This material can be isolated/encapsulated/blended with non-acid generating material. However, it should be pointed out that the lack of precipitation at the Project minimizes the risk of “drainage” from the storage areas.

20.3.3 Tailings Storage Facility Closure Plan

The TSF will include the construction of a tailings dam with a lined inner slope/face with tailings deposition designed to direct and collect water at the back of the dam for reuse. The outer face will be constructed of clean overburden rock to create a stable 3:1 slope. Towards the end of mining and milling, it is proposed that an over-liner of clean overburden be placed on the surface of the TSF.

The flotation tailings are expected to be chemically benign, non-acidic, and unlikely to generate acid according to initial ABA testing results.

The Project is located in a very dry environment, which is expected to make closing the TSF easier. A spillway to pass surface water runoff from the impoundment in a controlled manner to prevent erosion of the embankment will have already been installed during the last embankment raise. Ausenco has assumed that all mechanical and electrical equipment, such as the cyclone station and jacking header system for tailings disposition, will be demolished, and that the salvage value of the material equals the cost of removal. Based on the geochemistry to date, any water seeping from the TSF should be clean water. Therefore, the underdrain ponds are planned to be removed and a canal to be excavated to the salar, allowing any seepage from the facility to discharge directly into the salar.

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20.3.4 Mineral Processing Facilities Closure Plan

Decommissioning and demolition of industrial structures related to the mine support and processing areas of the Project will generally follow this sequence:

 Completion of all electromechanical installations/equipment decommissioning and removal of such equipment from the Project site;  Dismantling of internal metallic structures (e.g., walkways, hand rails, and platforms) by dismounting or by oxy-acetylene or arc cutting of reinforcing or support elements;  Removal of metal shrouds or shields by dismounting or by oxy-acetylene or arc cutting;  Removal of catwalks, ladders, and stairways;  Removal of insulation materials;  Removal of metal-framed doors and windows;  Removal of metal sheathing and roofing;  Disassembling of structural steel beams and joists; and  Demolition of vertical walls, cut-off walls, or berms made of concrete slabs or blocks via jackhammers or picks mounted on an excavator arm or other hydraulic equipment.

The onsite electrical power lines are planned to be maintained during closure as long as necessary for the operation of several facilities, including mine support. Once the electrical power lines no longer provide any use to the Project, they would be dismantled, salvaged, and the area reclaimed. However, the power lines may be left in place for use by the community, if needed.

20.3.5 Mine Site Roads Closure Plan

After operations cease, the out-of-pit mine haul roads are envisioned to be decommissioned. All other site roads are envisioned to be left in place to use to monitor the facility during and after closure.

20.3.6 Safety and Security

Site maintenance and security are expected to be ongoing activities during closure and would be subject to periodic monitoring. Drainage/diversion systems would be installed to ensure continued integrity and proper function of closed facilities.

Site security during the early years of closure will be an important concern, with varying numbers of contractors and workers onsite, and fewer mine employees overseeing the Project. Security measures may include the use of lockable gates and signs, as well as regular security checks.

20.3.7 Final Closure and Environmental Monitoring

Final closure activities are contemplated to commence when facilities are no longer needed for operation. The environmental monitoring program developed for baseline and operations would be modified for closure to:

 Assess compliance;  Ensure closure activities were performed as required;

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 Determine effectiveness of mitigation measures;  Evaluate the accuracy of the impact predictions;  Evaluate the accuracy of risk predictions;  Compare changes in the environment with existing baseline (pre-development) conditions and distinguishing project-related impacts from natural events;  Detect any unacceptable impacts to enable the implementation of supplementary mitigation and/or contingency measures in a timely manner;  Provide project-specific data on field performance of various cover materials, combinations, and thickness;  Determine the effectiveness of reclamation measures carried out as part of closure;  Ensure compliance with applicable environmental regulations and guidelines (local and international);  Ensure compliance with permit/license requirements;  Ensure accountability through a system of routine reporting to mine management, including summary reports being sent to applicable regulatory agencies;  Investigate environmental incidents or accidents, and identify follow-up requirements; and  Document and respond to public or regulatory agencies’ concerns.

20.3.8 Schedule

The final closure is estimated to require approximately two years following completion of mining and processing operations. After closure, a period of at least five years is estimated to be needed for post-closure monitoring and maintenance.

20.3.9 Closure Cost Estimate

A cost estimate for the proposed closure activities has been prepared, based on current Project definition. Activities described in the closure plan include: structure dismantling, site grading, tailings covering, contingency efforts, and monitoring. Estimated disturbed areas have been determined from the Project drawings. A unit cost per hectare of disturbed area, determined from historical data from similar projects, has been applied to the total disturbed area and used in the Project cash flow model.

20.4 Socioeconomic Conditions

20.4.1 Current Project Status

Exploration activities have been ongoing for several years; however, Lumina Copper began to work intensively at the Project in 2010. As part of its exploration activities and in preparation for the PEA, the Taca Taca team has been addressing a number of social issues. In particular, it has been managing local employment associated with the exploration, conducting a modest amount of purchasing from local and provincial providers, providing donations and social investment support to Tolar Grande, and managing ongoing engagement with key stakeholders. There are no local land users, and the Province of Salta is the owner of all of the surface rights within the Project area; as a result, the Project does not anticipate needing to manage either local land acquisition or resettlement issues.

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Senior management of Lumina Copper has been involved in assessing, planning, and implementing social management activities. As the Project has advanced, it has also dedicated several team members to managing the key social issues. Social Capital Group (SCG) has been providing advisory support to the Project with regards to social management.

20.4.2 Legal Requirements

The social legal and regulatory framework for the development of the Project on a national and international level is described as follows:

a) Articles 246 and 268 of the Mining Code of the Republic of Argentina require Environmental and Social Impact Reports for mining exploration and exploitation. The Federal Council of Mining (an institution that is composed of both the national and provincial governments) has established the contents for these reports during prospection, advanced exploration, and mining development. These mandate the social issues that should be addressed in the baseline studies, the project description, the impact assessment, and the management plan. The Mining Secretariat of the Province of Salta has also established the requirement of a Public Audience prior to the approval of an Environmental and Social Impact Report for mining exploitation. b) IFC. The World Bank’s Group Policy and Performance Standards regarding Social and Environmental Sustainability (IFC, 2006) is an essential reference for the design and application of a social management policy and plan. This document details a series of standards, requirements, and minimum criteria to be used for the development of investment projects, which will be used as a reference provided that they can be applied to the Project’s specific case. c) Equator Principles. These are principles adopted by a broad set of international financial institutions regarding requirements for social and environmental impact and performance of the projects that they finance. They are based on IFC Performance Standards.

20.4.3 Area of Influence

The area of influence of the Project lies entirely within Salta Province.

The direct area of influence during the exploration phase has been the municipality of Tolar Grande (the municipality has less than 150 residents). The town of Tolar Grande is approximately 35 km from the proposed Project area. If the Project is developed, the municipality of San Antonio de los Cobres (at least three hours away by car) would also receive significant economic benefit, as both a logistical base and source of labour, goods, and service providers, and would come to be considered part of the direct area of influence. The entire population of the Department is estimated to be less than 1,600 families, and its population density is less than one inhabitant per square kilometre.

The Project would use rail and road transport to bring supplies from other parts of Argentina and Chile. From the east, supplies would be brought in by rail or up the main highway from Salta past San Antonio de los Cobres and through the small settlement of Olacapato, where they would then travel west passing through the small settlement of Pocitos on the way to Tolar Grande. Supplies from Chile would primarily arrive by road via Sico Pass, then down to Olacapato and west to Tolar Grande through Pocitos; however, in some cases they might also come by rail. Alternative road options will be considered to shorten distances and reduce traffic impacts. The Project would seek to make significant use of rail transport within Argentina and to Chile, which should reduce traffic impacts.

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The indirect area of influence of the Project would come to include much of the rest of the Province of Salta. The province would receive significant economic benefits from the development and operation of the Project. The development of the Project would be the largest investment in the history of the province.

20.4.4 Social Context

The most relevant issues of the social context are listed below.

a) Economically disadvantaged area – The entire Departamento de los Andes is a sparsely populated area. Local employment comes primarily from work with the municipality or opportunities linked to local mining projects; in addition, some members of the population live off government subsidies. There is neither agricultural nor ranching activity in the region due to the challenging environmental conditions. The land in the area of the Project is unoccupied (it shows no signs of occupation in human history prior to the development of the railroad), would not support any type of agricultural or ranching activity, and is owned by the provincial government. In this context, the Project must work to promote the positive economic impacts that it would bring to both the municipality and department: i) Employment – The Project will need to identify potential project workers in the department, assess their skills and establish programs to involve them in both skilled and unskilled opportunities associated with the Project. It will also need to identify mechanisms to ensure that employment opportunities are broadly shared across the population. ii) Business development – The Project will need to assess the existing businesses within the department to identify ability and interest to potentially work as Project suppliers and service providers. It should also set up initiatives to promote local businesses’ involvement with Project development and operations. These initiatives should focus on building local administrative and technical capabilities to take advantage of Project opportunities as well as subsequent opportunities that might arise with economic growth in the department and province. iii) Taxes – Project development would generate very significant revenue for the local government through taxes. In order to ensure that the municipal authorities are positioned to make the best use of these funds, the Project should evaluate assisting in government planning and administrative capacity building in the years prior to the increase in revenue. b) Water and environmental concerns – The principal environmental concern in the municipality and department is water. Water is scarce and potable water is even more difficult to locate or utilize. Underground water is often salty or has other contaminants. Stakeholders will be concerned with whether or not the Project might negatively impact either the quality or quantity of their existing water sources or if the Project might be able to assist them in securing either more or better access to clean water. Only one potential water source for the Project is related to the sources used by Tolar Grande. However, Lumina Copper is committed to finding new sources of water for the Project in order to not have to tap into this community resource. Although the needs have been communicated by Lumina Copper, stakeholders may not be sufficiently aware of the Project’s potential sources or overall demand for water and, if the matter is not properly managed, incorrect perceptions could lead to unnecessary tension with the Project.

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c) Access – It usually takes about five hours to travel from San Antonio de los Cobres through Tolar Grande and onwards to the Project by bus. Part of the route between San Antonio de los Cobres and Tolar Grande can become impassable for several days during the year due to snow or flooding. The Project may need to evaluate alternative road routes that would enable Project development, but also improve year round access between Tolar Grande and San Antonio de los Cobres, unless other companies or the Province address this matter sooner. Any new options should prove beneficial to the local communities, by providing them with alternatives to the current road and its occasional challenges. d) Comunidad Kolla - The government recognized the Comunidad Kolla de Desierto as an indigenous community in the last decade. Its territory extends from Tolar Grande to the border with Chile and encompasses Taca Taca. The collective rights of the Comunidad Kolla del Desierto do not include actual ownership over land rights or other resources that they have not traditionally held or used (neither Taca Taca, nor the surrounding area shows any indication of recent or past occupation). The actual surface rights over most of the land in this area, other than any specific properties owned by individuals in the communities of Pocitos and Tolar Grande, belong to the Province of Salta. The exact number of community members is estimated to be around 50. The community members make up the majority of the adult residents of the town and participate in the daily life of the town and the town meetings. They do not speak a different language. The community members have participated, along with other town members, in ongoing Project consultation. The cacique is the elected head of the community. The Project maintains contact with both the current and previous cacique and also coordinates with him together with the mayor (Intendente). e) Consultation – The Project has maintained ongoing contact with project stakeholders throughout its exploration activities. Most of the regular contact has been in Tolar Grande, but it has also engaged stakeholders in San Antonio de los Cobres and Salta. Mining exploration is a familiar activity in the municipality and stakeholders are well aware of the Project’s activities. As the Project advances towards feasibility, mining is seen as an opportunity and local stakeholders have generally indicated positive expectations towards the potential development. That said, some of the community leaders have expressed concerns about the effects the Project would have on the socioeconomic and political structure of Tolar Grande.

20.4.5 Stakeholder Mapping

The Project has identified stakeholders and established engagement at various levels: Tolar Grande, San Antonio de los Cobres, Salta, and the national level. To date, most of the engagement has been focused on Tolar Grande. As the Project advances, Lumina Copper has identified the need to engage or continue to engage all stakeholders.

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21 Capital and Operating Costs

21.1 Capital Cost Estimate

Capital cost estimates were prepared for the Project PEA assuming construction of a mining and sulphide concentrating operation to produce copper and molybdenum concentrates. Based upon the mine plan developed for the Project, mining operations over a 28-year mine life would deliver 120,000 t/d of throughput to a mill and concentration plant comprised of two mill and flotation lines. Throughput of 180,000 t/d commencing in Year 8 is expected to be reached with the addition of a third line in the concentrator plant, allowing for increased production capabilities.

Table 21-1 summarizes the capital costs by major area. Initial capital costs are detailed further in Table 21-5. LOM sustaining capital costs are estimated at a total of $1.8 billion, including capital costs of $431 million to be expended in Years 6 and 7 with the anticipated concentrator plant expansion. Working capital requirements are estimated at $55 million. The estimates are expressed in first-quarter 2013 United States dollars.

Table 21-1: Mine Capital Cost Summary

Estimate Description ($000's) Mine 581,445 Preproduction Stripping 416,134 Plant and Processing 685,469 Infrastructure 215,340 Initial Capital Tailings Storage and Waste Rock Storage 139,540 Other 581,026 Contingency 386,513

Total Initial Capital (000's) $3,005,467

LOM Sustaining Capital Costs (excluding plant expansion) 1,375,463 Capital Costs for Plant Expansion and associated ancillary 431,234 Sustaining Capital costs

Total Sustaining Capital (000's) $1,806,698

Working Capital Total Working Capital (000's) $54,949

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21.2 Estimate Exclusions

The following items are not included in the capital estimate:

 Costs incurred prior to completion of a positive Feasibility Study (FS);  All Owner’s taxes, including any financial transaction tax, withholding tax, or value-added tax (VAT);  Reclamation costs, which are included in the financial analysis;  Future foreign currency exchange rate fluctuations;  Working capital and sustaining capital, which are included in the financial analysis;  Interest and financing costs;  Escalation beyond first-quarter 2013; and  Risk due to political upheaval, government policy changes, labor disputes, permitting delays, weather delays, or any other force majeure occurrences.

21.3 Basis of Estimate – Mining

WLRC developed mining capital costs by conducting FC analysis and a subsequent mine plan and production schedule. Mine production and concentrator feed schedules were then estimated from the phase resource tabulations. A three-year preproduction stripping period, allowing for personnel and equipment ramp-ups, will be needed to expose sufficient mineralized material for initial concentrator operations.

Mine equipment unit prices were derived from recent vendor quotations for a similarly sized project and include provisions for freight, insurance, and assembly. Costs were estimated for auxiliary equipment, including: explosives storage and handling, small excavating and transport equipment, a portable aggregate crushing and screening plant, electric power servicing trucks, all-terrain and crawler cranes, fuel/lube trucks, mechanic field trucks, tire handling equipment, forklifts, light plants, and assorted light vehicles. Allowances were made for shop equipment, pit water handling systems, a truck dispatch system, radio communications, office equipment and software, initial mine haul road construction and site preparation (for equipment erection), spare parts and supplies inventories, and contingency.

21.4 Basis of Estimate – Process and Infrastructure Direct Costs

21.4.1 Equipment and Materials

Ausenco estimated process and infrastructure direct costs using:

 Material take-offs (MTOs) based on preliminary layouts, process flow diagrams, and limited topographic information;  Budget quotations for major equipment;  Historical data; and,  Allowances based on similar projects.

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21.4.2 Construction Direct Labor

Estimated labor hours and rates have been benchmarked against recent projects in South America. A productivity factor of 2.5 has been applied to standard database hours to account for the remote location, expected skill level of working crews, high altitude, and windy weather applicable to this site.

The hourly labour rate includes direct labour, supervision of direct labour, mark-ups for non-working construction staff, transportation to and from the work site, small tools and equipment, indirect labour, applicable subsistence testing and material, safety, and hazardous awareness training.

A typical work week of 60 h has been assumed based on shifts of 10 h/d and 6 d/wk. The labour rates per each discipline for both the initial construction and plant expansion estimates are as follows.

Table 21-2: Labour Hourly Rates

Discipline Labour Hourly Rate ($US) Earthworks $48.00 Concrete $30.00 Structural, Mechanical, Piping $40.00 Electrical / Instrumentation $38.00

It has been assumed that the 60,000 t/d concentrator expansion construction will take place when the plant is operational, so brownfield productivity factors have been applied to all installation hours. These factors are based on experience with similar projects that consider the crowded work environment, shutdown periods, as well as safety and operations ongoing in the plant. These factors were applied to each discipline as follows.

Table 21-3: Brownfield Productivity Factors

Discipline Brownfield Productivity Factor Earthworks 1.15 Concrete 1.15 Structural, Mechanical, Piping 1.20 Electrical / Instrumentation 1.20

21.5 Basis of Estimate – Indirect Costs

Lump sum allowances or factors have been used to calculate indirect costs as is typical for a PEA. At this level, many of the resourcing and contract strategies are not defined, so reasonable and customary assumptions have been made based on experience on similar projects.

21.6 Owner’s Costs

Ausenco included a lump sum allowance of $3,000,000 in the initial capital cost estimate for equipment required at the concentrator. No allowance has been made for additional mobile equipment for the 60,000 t/d concentrator expansion.

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Lumina Copper and MTB developed the majority of the Owner’s costs based on factored actual project data, budgetary quotations, in-house historical data, published databases, and recent similar project information.

TFP provided capital cost estimates for the rail spur, new track, locomotives, wagons and rail maintenance facilities. Hugo Gil provided costs for the installation of the 345 kV power transmission line and associated infrastructure. Schlumberger provided costs for water treatment.

21.7 Other Costs – Freight, Duties & Taxes

21.7.1 Freight

Freight costs are included for bulk materials as well as for the segment of transportation of major equipment that the vendor has not supplied freight for in budget quotations. It is assumed that the bulk materials and majority of equipment can be supplied within Argentina.

A bulking factor was applied to inland freight for tanks and flotation cells due to the high volume to weight of the equipment. Below is a table showing the freight percentages applied.

Table 21-4: Freight Applied to Material/Equipment Cost

% Applied to Material / Freight Segment Equipment Cost Ocean Freight From Port in Americas (North & South America) 8.0% From Port Overseas (Europe, Asia, Australia, Africa) 12.0% Inland Freight From Port of Entry, Buenos Aires 2.0% From Closest Large City, Salta 1.0% Inland Freight Bulking Factor (Tanks, Flotation Cells) From Port of Entry, Buenos Aires 6.0% From Closest Large City, Salta 5.0%

21.7.2 Duties & Taxes

Duties and taxes are not included. Based on advice from TAGING Ingenieria Inteligente (TAGING), an Argentinian engineering firm engaged to provide estimating support, project equipment and materials will not be subject to duties.

21.8 Contingency

Contingency is an allowance to cover unforeseeable costs that may arise during the project execution, which reside within the scope-of-work, but cannot be explicitly defined or described at the time of the estimate, due to lack of information. It is assumed that contingency will be spent. However, it does not cover scope changes or project exclusions.

After an assessment by the Project team of Project confidence versus uncertainty by Project area, an overall average contingency of 15%, or $386.5 million, has been included in the initial capital

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cost for the Project. This contingency is based on the level of definition that was used to prepare the estimate. Sources for pricing included budgetary quotations, in-house historical data, published databases, factors, and estimators’ judgment.

21.9 Accuracy

This estimate has been developed to a level sufficient to assess/evaluate the Project concept, various development options, and the overall potential Project viability. After inclusion of the recommended contingency, the capital cost estimate is considered to have a level of accuracy in the range of minus twenty-five percent (-25%) plus thirty-five percent (+35%). This is based on the level of contingency applied, the confidence levels of the authors in their respective estimates, and an assessment comparing this estimate to standard accuracy levels on PEA estimates.

21.10 Initial Capital

The total estimated initial capital cost to design, procure, construct, and commission the facilities described in this report is $3.005 billion. The initial capital cost estimate is summarized in Table 21- 5.

Table 21-5: Summary of Initial Capital Costs

Estimate Subtotal Total Description ($000's) ($000's) ($000's) Mine 108,682 Crushing, Conveying and Storage 71,221 Grinding and Concentrator 504,537 Concentrate Thickening, Filtration, Storage, Handling 47,206

Reagents/Consumables 17,942 Direct Costs Storage/Distribution/Handling 1,151,673

Tailings Management 106,680

Ancillary Facilities 14,997 Site and Off-Site Utilities 217,829 Site and Off-Site Development 62,580 Contractor Indirects 35,286 Construction Construction Temporary Facilities 33,600 Indirect 130,718 Costs Construction Equipment 1,832 Construction Camp, including catering 60,000 EPCM Services 151,300 378,988 Geotechnical Facilities E&CQA 7,645 Contracted Indirect Third Party Inspections/Testing (QAQC)/Surveying 5,200 248,270 Costs Vendor Representatives/Commissioning Assistance 20,542 Initial Fills and Mine/Plant Equipment Spare Parts 63,582 Preproduction Mine Development 416,134 Owner's Primary Mine Equipment 441,460 977,196 1,057,462 Direct Costs Ancillary, Plant Mobile Equipment, Light Vehicles 47,390 Owners Camp 11,496

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Estimate Subtotal Total Description ($000's) ($000's) ($000's) Office/Engineering Equipment, Software, Furniture ,104 Plant & Mine Warehouse/Truck shop Equipment 1,000 Medical, Safety, Security, Communication 2,021 Railroad and Maintenance Facilities, Airstrip 55,592 Preproduction Employment & Training 26,961 Project & Construction Management 10,619 Operations Catering 10,435 Camp Power 5,161 Owner's ROW, Land Acquisition, Legal, Permits, Fees 6,912 Indirect 80,265 Costs Insurance 14,910 Corporate Travel & Services 1,267 Environmental 1,350 Medical, Security, Communication 1,250 Community Development 1,400 Freight, Duties & Taxes 30,831 Other 417,344 Contingency 386,513 Total Initial Capital Cost (000's) $3,005,467 $3,005,467

21.11 Operating Costs Costs have been estimated by operating areas of mining, processing, infrastructure maintenance, railroad operation, G&A, and mine reclamation and closure. For incorporation into the operating cost estimate, costs were reported under subheadings related to the functioning of each of the areas identified. Table 21-6 summarizes LOM operating cost by area, as well as average unit operating cost for the first seven years of production at 120,000 t/d, for Years 8-28 at 180,000 t/d and LOM.

Table 21-6: LOM Operating Cost Summary Average Unit Operating Cost Total Years 1-7 Years 8-28 Life of Mine LOM Description 120 Kt/d 180 Kt/d Cost $/T Process $/T Process $/T Process $000's Feed Feed Feed Mining 7,716,616 7.55 4.05 4.67 Processing 7,026,039 4.34 4.24 4.26 Infrastructure Maintenance 104,157 0.13 0.05 0.06 Railroad Operation 313,874 0.28 0.17 0.19 General and Administration 940,092 0.77 0.53 0.57 Mine Reclamation & Closure 30,518 ** ** 0.02 Total Operating Cost $ 16,131,296 $ 13.07 $ 9.04 $ 9.77 LOM = 28 years LOM K-Tonnes of Process Feed: 1,650,792 ** No Average Annual Cost indicated as cost is considered incurred after LOM in Year 28 and after.

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21.12 Basis of Estimate

The operating cost estimate addresses the operating cost assumptions used in the PEA for the proposed mining and sulfide concentrating operation starting as a 120,000 t/d operation, expanding to 180,000 t/d in Year 8.

The main operating cost assumptions used in the PEA are shown in Table 21-7. The operating cost estimate is based on a several sources including cost service data, vendor quotations, first principle calculations, and similar historical projects.

Table 21-7: Main Operating Cost Assumptions

Item Unit Cost Estimate Comment Diesel Fuel $/litre 1.00 Provided by MTB Electricity $/kWh 0.085 Provided by MTB $/h 8-20 Argentine Labour rates provided by Taging Labour % Benefits 70 Taging estimate for hourly labour

21.13 Mining Operating Costs

Mine operating costs are based on Owner mining and were developed by WLRC. Table 21-8 details the LOM operating cost summary for mining.

The Project mine operating cost estimates cover the following: pit operations (i.e., drilling, blasting, loading, and hauling); construction and maintenance of mine haul roads, sumps, and safety berms; placement of waste rock in the WRSFs; operating and maintenance labour; mine department supervision and technical services; crushing waste rock to supply aggregate for road surfacing and blasthole stemming; and other earthworks as may be required for day to day mining operations. The mine production schedule and equipment unit productivity estimates were used to calculate operating shifts and manpower requirements, which in turn were used to derive mine operating costs. Exploration costs are not included in the operating cost estimates.

Unit operating costs for major equipment incorporate labour, fuel and lubricant consumption, vendor estimates of maintenance and repair costs, and tire or undercarriage costs. These operating costs were adjusted for local labor rates and supply costs, while tracking recent experience for projects with similar fleets. The mining cost estimates were based on energy prices of $0.085 kWh for electricity, $1.00/l for diesel fuel, $0.80 kg for bulk ammonium nitrate prills, and $1.40 kg for bulk water-resistant emulsion.

Mine operating and maintenance labour rates range between $5.90 and $15.44 per hour. A fringe benefits burden of 70% was applied to the base labour costs; expatriate fringe burdens were capped at 40%. Overtime, paid at 1.5 times the base rate, was projected at 5% and 10% for craft operating and maintenance personnel, respectively.

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Table 21-8: LOM Operating Cost Summary - Mining

Average Unit Operating Cost Total Years 1-7 Years 8-28 Life of Mine LOM Description 120 Kt/d 180 Kt/d Cost $/T Process $000's $/T Process $/T Process Feed Feed Feed Drilling 517,976 0.55 0.26 0.31 Blasting 992,573 1.06 0.50 0.60 Loading 668,425 0.71 0.34 0.40 Hauling 3,983,454 3.73 2.13 2.41 Roads & Dumps 825,399 0.78 0.44 0.50 General Mine 312,907 0.31 0.16 0.19 General Maintenance 149,238 0.16 0.08 0.09 Supervision & Technical 297,141 0.27 0.16 0.18 Subtotal 7,747,112 7.55 4.08 4.69 Excavation of Salar (30,496) (0.007) (0.021) (0.018) Total Operating Cost $ 7,716,616 $ 7.55 $ 4.05 $ 4.67 LOM = 28 years LOM K-Tonnes of Process Feed: 1,650,792

21.14 Process Operating Costs

Ausenco developed the operating cost estimate associated with the mill and concentrator plant facilities. A summary of the average operating cost per tonne of feed treated for the Project is outlined in Table 21-9 and Figure 21-1. The costs have been separated into the key cost components. All costs have been based on estimates as of Q1 2013. The operating costs are considered to have accuracy of ± 30%, based on the assumptions listed in this section of the report.

Table 21-9: LOM Operating Cost Summary - Process

Average Unit Operating Cost Total Life of Mine Years 1-7 Description Years 8-28 LOM Cost 120 Kt/d 180 Kt/d $/T Process $/T Process $000's $/T Process Feed Feed Feed Labor 257,963 0.20 0.15 0.16 Power 3,014,782 1.75 1.84 1.83 Maintenance Materials 313,703 0.21 0.19 0.19 Reagents & Consumables 3,372,678 2.12 2.03 2.04 Miscellaneous 42,498 0.03 0.02 0.03 Concentrate Transportation 24,415 0.02 0.01 0.01 Total Operating Cost $ 7,026,039 $ 4.34 $ 4.24 $ 4.26

LOM = 28 years LOM K-Tonnes of Process Feed: 1,650,792

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0.8% 0.5% 4.6%

Labour

Power 40.5% 49.0% Maintenance Materials Reagents & Consumables Miscellaneous

Concentrate 4.8% Transportation

Figure 21-1: Percentage of Processing Costs by Component

The plant operating cost estimate has been developed from a number of sources. Cost determinations have been based on fixed and variable components relating to plant throughput and mineral characteristics. The sources of data are detailed in Table 21-10.

Table 21-10: Derivation of Plant Operating Estimate

Cost Category Source Of Cost Data Power Consumption from load estimate and power rate agreed by project team. Manning schedules developed by Ausenco from similar sized projects in the Labour region. Rates provided by Taging Ingeniera. Consumptions from Plenge test work and benchmarking data; unit prices from Reagents suppliers and Ausenco database. Consumptions predicted from test work and experience and unit prices from Consumables suppliers and Ausenco database. Maintenance Calculated as a percentage of direct capital costs. Materials

21.15 Infrastructure Maintenance

Table 21-11 shows a LOM operating cost for mine infrastructure maintenance. Ausenco provided costs for the mine access road maintenance, and rail and rail spur rehabilitation and maintenance were provided by TFP.

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Table 21-11: LOM Operating Cost Summary – Infrastructure Maintenance

Average Unit Operating Cost Total Years 1-7 Life of Mine Years 8-28 120 Kt/d LOM Description 180 Kt/d Cost $/T $/T Process $/T Process $000's Process Feed Feed Feed Mine Access Road Maintenance 26,419 0.023 0.015 0.016 Rail Rehabilitation & Maintenance 76,537 0.110 0.033 0.046 Rail Spur Rehab & Maintenance 1,201 0.001 0.001 0.001 Total Operating Cost $ 104,157 $ 0.13 $ 0.05 $ 0.06

LOM = 28 years LOM K-Tonnes of Process Feed: 1,650,792

21.16 Railroad Operation

Table 21-12 provides LOM costs for operating the railroad according to estimates by TFP. Labour rates for rail personnel and maintenance rates were provided by TFP and $1.00/l was used for diesel fuel for estimated consumption provided by TFP.

Table 21-12: LOM Operating Cost Summary – Railroad Operation

Average Unit Operating Cost Total Years 1-7 Years 8-28 Life of Mine LOM Description 120 Kt/d 180 Kt/d Cost $/T Process $000's $/T Process $/T Process Feed Feed Feed Labor 29,865 0.03 0.02 0.02 Maintenance 53,057 0.05 0.03 0.03 Diesel Fuel 230,953 0.21 0.13 0.14 Total Operating Cost $ 313,874 0.28 0.17 $ 0.19

LOM = 28 years LOM K-Tonnes of Process Feed: 1,650,792

21.17 General and Administrative

The G&A operating costs are described below and are shown in Table 21-13.

A preproduction staffing organization chart was developed to identify all G&A staff and labour positions. The final month of preproduction provided the fixed annual labour cost at start-up. Each staff and wage labour title was assigned a pay level provided by Taging for Argentine labour rates plus burden. Eight positions were estimated at expatriate rates, which are included for the duration of the life of mine: General Manager, Executive Mine Manager, Mill Manager, Mill Superintendent, Mechanical Superintendent, Chief Metallurgist, Technical Services Manager, and Maintenance Manager. The balance of G&A expenses were provided by Lumina Copper or estimated by MTB based on similar recently completed studies.

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Table 21-13: LOM Operating Cost Summary – G&A

Average Unit Operating Cost Total Years 1-7 Years 8-28 Life of Mine LOM Description 120 Kt/d 180 Kt/d Cost $/T Process $000's $/T Process $/T Process Feed Feed Feed

General and Administration 940,092 0.77 0.53 0.57 Expenses

Total Operating Cost $ 940,092 0.77 0.53 $ 0.57

LOM = 28 years LOM K-Tonnes of Process Feed: 1,650,792

21.18 Mine Reclamation

The Taca Taca reclamation and closure plan will focus on safety, stabilization of the land surfaces, post mine utilization of facilities and structures, and protection of the environment. Activities described in the reclamation and closure plans include structure dismantling, site grading, tailings covering, contingency efforts, and monitoring. Since the Project is located in an extreme arid, high altitude environment, revegetation is considered impractical and not conducive to the surrounding environment.

A cost per hectare for closure budgeting purposes is estimated to be $7,500. This estimate was provided by Lumina Copper’s environmental consultant and is based on historical experience at other mining properties as an all-in unit cost per hectare of disturbance, applied to the estimated total project disturbance area. This unit cost has been applied to a total disturbed area of 4,069 ha to obtain an estimated closure cost of $30.518 million. This estimate is used in the Project cash flow for final reclamation and closure and is shown in the final year of the financial model. Where feasible, concurrent reclamation and closure will be incorporated into operations in order to minimize/limit this number at the end of mine life.

Removal of equipment and structures is considered to be at no cost because the value of surplus/scrap materials will cover contractor’s cost.

21.19 C-1 Cash Costs (net of credits)

The C-1 cash costs were calculated using the Project economic model’s cash flow forecast values:

 Total operating costs

 Royalty costs including Provincial Mining Royalty and third party royalty

 Treatment costs, refining costs, and transportation costs (i.e., third party rail fee, port handling, and ocean freight)

 Revenue from gold and molybdenum

To calculate the cash cost per pound of copper, total expenses (operating, royalty, and TCs, RCs, and transportation) less total revenue from Au and Mo were divided by the number of pounds of

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copper to be sold over life of mine. Average life of mine cash cost was calculated to be $1.11/lb copper sold as shown in Table 21-14.

Table 21-14: Average LOM C-1 Cash Costs

Description LOM Amount Cumulative Amount Operating Cost $16,131,296,196 Royalties $1,529,016,493 TCs, RCs, Transportation $4,474,494,091 Subtotal Cash Cost w/o Credits $22,134,806,780 Gold, Molybdenum Credit ($5,467,067,887) Subtotal Cash Cost w/ Credits $16,667,738,893 Total Copper to be Sold 15,082,906,988 pounds Average LOM Cash Cost $1.11 per pound of copper

Annual cash costs per pound of copper were estimated using the same methodology. Total revenue from Au and Mo for each year was subtracted from total expenses (operating, royalty, and TCs, RCs, and transportation) from the same year. The result was then divided by the number of pounds of copper to be sold in the corresponding year. Estimated annual cash costs range from $0.79 in Year 10 to $1.46 in Year 21, discounting first and last years of mine life. Figure 21-2 provides a graph of the annual C-1 cash cost per pound of copper for each year in the life of mine.

Figure 21-2: C-1 Cash Costs Net of Credits

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22 Economic Analysis

22.1 General Criteria

MTB, under the supervision of the Qualified Person for this section, completed a PEA economic evaluation of the Project in Argentina for Lumina Copper. Major model inputs are provided below in Table 22-1.

Table 22-1: Economic Model Inputs

Description Values

Construction Period 30 months Preproduction Period 4 years Life of Mine (LOM) after Preproduction 28 years LOM Sulphide Mill Feed (Kt) 1,650,792 LOM Copper Concentrate (DMT) 21,615,739 LOM Molybdenum Concentrate (DMT) 227,884 Market Price Copper Price (LOM $/lb Cu) 2.75 Gold Price (LOM $/oz Au) 1,200.00 Molybdenum Price (LOM $/lb Mo) 12.00 Cost and Tax Criteria Estimate Basis 1st Qtr 2013 USD Inflation/Currency Fluctuation None Leverage 100% Equity Income Tax - Argentina 35% Argentine Retention Tax 10% Less: Puna Investment Credit -2.5% Net Retention Tax 7.5% (on Gross Revenue less: TCs, RCs, port handling, ocean freight, and

75% of rail use fees and railroad operation costs to be incurred in Chile)

Depreciation - Argentina Section 13 Infrastructure Construction & Equipment 60/20/20 Machinery, Equipment, and Vehicles 1/3 for 3 years

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Description Values

IVA - Value Added Tax (VAT) Payment/Recovery Excluded Royalties Salta Provincial Minemouth Royalty 3.0% (on Gross Revenue less: TCs, RCs, freight, processing, infrastructure maintenance, railroad operation, and G&A costs) Third Party Royalty on Modified NSR 1.5% (on Gross Revenue less: TCs, RCs, freight, railroad operation, Provincial Minemouth Royalty, and Net Retention Tax paid) Concentrate (Con) Transportation Charges Cu Con: Rail Use Fees, Port Handling, and Ocean Shipping ($/WMT) 79.67 Mo Con: Rail Use Fees and Port Handling ($/WMT) 15.67 Assuming containerization of Mo con product and considered delivered

22.2 Production Summary

At the foundation of the economic model, data was drawn from the mine production schedule developed by WLRC (refer to Table 16-6) and utilized by Ausenco to produce the process production schedule summarized in Table 22-2 below.

Table 22-2: Process Production Summary

Total Concentrate Recovered Metals in Concentrate Feed Produced Copper Con Molybdenum Con Year Processed Copper Molybdenum Copper Gold Molybdenum K-Tonnes Tonnes Tonnes Tonnes Lbs Ozs Tonnes Lbs 1 34,920 511,782 4,318 174,006 383,613,350 69,855 1,956 4,311,139 2 43,200 1,025,680 8,012 348,731 768,812,513 107,599 3,629 8,000,052 3 43,200 872,316 7,038 291,095 641,747,448 102,963 3,307 7,291,090 4 43,200 911,277 8,608 308,517 680,156,498 107,808 3,938 8,682,457 5 43,200 897,838 7,733 292,681 645,244,622 94,394 3,772 8,316,091 6 43,200 794,405 8,595 251,642 554,769,536 114,625 4,229 9,324,200 7 43,200 720,514 8,828 231,218 509,742,860 125,307 4,305 9,491,134 8 62,750 831,371 6,538 273,181 602,254,670 129,904 3,195 7,044,490

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Total Concentrate Recovered Metals in Concentrate Feed Produced Copper Con Molybdenum Con Year Processed Copper Molybdenum Copper Gold Molybdenum K-Tonnes Tonnes Tonnes Tonnes Lbs Ozs Tonnes Lbs 9 64,800 863,611 7,191 277,896 612,648,913 130,148 3,611 7,960,464 10 64,800 1,094,594 13,484 330,444 728,496,887 193,767 7,041 15,522,999 11 64,800 895,629 10,016 273,215 602,329,929 146,727 5,236 11,542,646 12 64,800 871,402 9,381 265,886 586,171,705 149,723 4,860 10,714,611 13 64,800 886,579 10,325 266,618 587,786,765 156,045 5,450 12,014,171 14 64,800 888,399 6,148 270,910 597,249,285 149,404 3,201 7,056,719 15 64,800 734,522 5,053 234,473 516,918,975 119,531 2,526 5,569,239 16 64,800 692,616 3,793 233,277 514,282,618 69,269 1,721 3,794,935 17 64,800 782,156 7,095 256,991 566,561,544 72,790 3,484 7,679,883 18 64,800 729,123 7,917 237,535 523,670,516 109,696 3,896 8,588,157 19 64,800 819,442 8,985 249,101 549,168,485 156,044 4,646 10,242,866 20 64,800 666,906 8,712 200,137 441,221,940 113,694 4,579 10,094,366 21 64,800 541,695 7,633 175,644 387,225,457 77,832 3,808 8,396,079 22 64,800 638,566 8,864 205,441 452,914,477 96,093 4,440 9,787,521 23 64,800 660,028 8,379 204,551 450,953,667 80,206 4,332 9,550,552 24 64,800 637,804 8,935 191,054 421,197,261 79,507 4,697 10,355,522 25 64,800 626,202 9,853 186,641 411,467,876 73,856 5,198 11,460,292 26 64,800 669,518 10,459 200,059 441,050,246 74,816 5,515 12,158,129 27 64,800 853,679 11,850 253,815 559,560,500 98,791 6,257 13,794,954 28 62,722 498,085 4,141 156,803 345,688,445 72,016 2,126 4,686,200

LOM 1,650,792 21,615,739 227,884 6,841,562 15,082,906,988 3,072,410 114,956 253,430,959

22.3 Gross Income from Mining

Market prices for copper, gold, and molybdenum have been selected after reviewing various recommendations as discussed in Item 19.0 and are considered to be conservative.

Terms of payment, minimum unit deductions, and discounts were based upon the H&H market study. Key inputs and the resulting gross revenue for LOM are shown in Table 22-3.

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Table 22-3: Calculation of Gross Revenue

Concentrate Recovered Market Min Deduct Revenue Payfor Payable Kt 000's Price or Discount $000's Cu 15,082,907 Lbs $2.75 /Lb 96.5% 40,026,264

Cu 3,072 $1,200 /Oz 100.0% 3,686,893 21,615.7 Au Ozs 0.0322 Ozs/T (835,232) 253,431 Lbs $12.00 /Lb 100.0% 3,041,172 Mo 227.9 Mo 14.00 % (425,764) Gross Revenue $000's $ 45,493,332

22.4 Net Smelter Revenue (NSR) Calculation

TCs and RCs as projected in the H&H market study were applied to the corresponding tonnes of concentrate or units of recovered metal. For purposes of calculating transportation, dry metric tonnes of concentrate were converted to wet metric tonnes assuming the concentrates contain 9% moisture. Three components of transportation are included in the economic analysis: rail use fee, port handling, and ocean freight. While the model incorporates Owner operation of rail freight, additional fees will be payable to owners of existing rail tracks to be utilized throughout Argentina and Chile. The estimated $10 per mile rail use fee was converted to $4.67/wmt.

Assuming a destination port of Mejillones, near Antofagasta, Chile, port charges are estimated at $22 wmt and $11 wmt of Cu and Mo concentrates, respectively. Cu concentrate travels to the port in bulk form and requires unloading to a facility for eventual loading onto shipping vessels for export to Asian markets. Since the Mo concentrate would be containerized at the mine site prior to loading on rail cars, lesser port charges would be incurred for offloading of the containers at the port and reloading to trucks. Mo concentrate is expected to be sold in Chile so there are no conveyance charges for ocean shipment.

Estimated by H&H, ocean freight for transport of Cu concentrate to Asian ports was included at $53/wmt. Components used in determining TCs, RCs, and freight are shown in Table 22-4. All treatment and refining charges and transportation costs were deducted from total gross revenue to obtain NSR.

Table 22-4: TCs, RCs, and Freight

Concentrate Treatment Con Port Ocean DMT Charges Recovered Metals Refining WMT Rail Use Handling Freight Metal 000's Charges Fee 000's /DMT 000's /WMT /WMT /WMT

Cu 15,082,907.0 $0.07 /Lb Lbs Cu 21,615.7 $70.00 23,561.2 $4.67 $22.00 $53.00

Au 3,072.4 $8.00 /Oz Ozs

Mo 227.9 Mo 253,431.0 248.4 $4.67 $11.00 Lbs

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22.5 Royalty Calculation

The Provincial Mining Royalty is a mineral tax imposed at the rate of 3.0% on the basis of mine head value. According to the Mine Investment Law, “mine head value” results from the first commercialization after considering the direct costs of exploitation, but not including capital asset depreciation. For purposes of this model, the Provincial Mining Royalty was calculated as 3.0% of NSR further reduced by other allowable deductions including processing, infrastructure maintenance, railroad operation, and G&A costs. NSR was defined in the previous section as gross revenue less TCs, RCs, and transportation expenses.

Additionally, a third party royalty is computed on a modified NSR, with 1.5% due on NSR less railroad operating expenses, Provincial Mining Royalties, and Net Retention Taxes paid. Royalties are considered paid in the corresponding annual period earned in accordance with the royalty terms. Deduction of total royalties from NSR yields gross income from mining.

22.6 Operating Margin

Operating cost estimates as previously described in Item 21.0 of this PEA served as input to the economic model. Total annual costs for each area (i.e., mining, processing, infrastructure maintenance, railroad operation, G&A, and mine reclamation and closure) are reported by year and then totaled for LOM operating cost. Annual and LOM unit operating costs per tonne of concentrator feed are calculated by dividing annual and LOM operating costs by corresponding tonnes of processed feed. Unit operating cost per lb Cu is found by substituting lbs Cu recovered in the denominator of the previous equation. Gross Income is determined by subtracting total annual operating costs from gross income from mining.

22.7 Retention Taxes

The Argentine Retention Tax is a gross sales revenue royalty treated as an export tax. Levied at the rate of 10% on concentrate value “at the border,” the retention tax is subject to partial reimbursement for mineral products produced in the Puna region of Argentina. Set forth by Resolution No. 762/93 of the Ministry of Economy, the investment credit was established to promote and develop mining in select isolated areas. Presently, the reimbursement is 2.5% and brings the retention tax to a net of 7.5%.

Basis for taxation is gross revenue less TCs, RCs, port handling, ocean freight, 75% of rail use fees, and 75% of railroad operating expenses. Twenty-five percent of rail transportation will occur in Argentina and 75% will occur in Chile, based upon actual kilometers to be travelled in each; hence the deduction of 75% of rail expense to establish “at the border” value of the concentrates. Subtraction of net retention taxes from gross income results in net profit before depreciation.

22.8 Depreciation and Income Tax

CASA provided a State of Accounts Relative to December 31, 2012 containing entries for the intangible assets shown in Table 22-5, which represent capitalized costs to date for the Project. Values were converted from Argentine pesos to US dollars at a rate of 5.04 ARS $: $1.

Argentine Income Tax Law (ITL) allows for deduction of expenses incurred in obtaining such assets in a manner proportionate to the exhaustion of the asset, i.e., unit of production. According to the procedure set by ITL, the total cost is divided by the number of units to be extracted and then multiplied by the number of units associated with a given period. For purposes of this model, the

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project-to-date (December 31, 2012) asset total of approximately $38.5 million was divided by LOM processed feed of 1,650,792 kt and then multiplied by tonnage attributable to each year of mine life.

In addition to the deduction of capitalized costs under ITL, Argentine Mining Investments Law (MIL) Section 12 permits the deduction of exploration and other expenses incurred in determining the feasibility of mining activities, resulting in double deduction for some expenses. In this analysis, approximately $37.9 million in capital expenditures were identified as exploration costs (as previously defined in the MIL) and therefore deductible over a five year period from the start of production.

Table 22-5: Deduction of Capitalized Costs to Date

Intangible Asset ARS $ US $ * Tax Treatment Amount Period

Taca Taca Mine 2,678,835 531,778 Group Investments in 191,023,813 37,920,285 Double Deduction $37,920,285 5-year Taca Taca Project

Total 193,702,648 38,452,062 Regular Deduction $38,452,062 U.O.P.

* Conv: 5.04 ARS = $1 US Total Deduction $76,372,347

Depreciation of other fixed assets is not addressed by ITL, except to state that acquisition costs must be apportioned according to the useful life of the asset. However, MIL Section 13 established an accelerated depreciation regime that may be elected by mining companies instead of the standard ITL allowances. All initial capital expenditures (exclusive of spare parts, consumables, and initial fills) were classified as belonging to one of two asset classes to be depreciated according to the methodology detailed in Table 22-6.

Table 22-6: MIL Section 13 Depreciation

Asset Depreciation Description Year 1 Year 2 Year 3 Investments in equipment and construction to 60% 20% 20% provide the required infrastructure

Investments in machinery, equipment, vehicles, 1/3 1/3 1/3 etc. not included above

Sustaining capital expenses (excluding spare parts, consumables, and initial fills) were categorized in the same way, as Article 13 extends to capital investments for the purpose of “expanding productive capabilities of existing mining operations as well as those that might be required during their functioning.” MIL guidelines also dictate that annual depreciation expense cannot exceed taxable income, but provisions are made for the carry-forward of excess depreciation, which was the case in the first three years of production.

After deduction of depreciation expense, Argentine income tax was calculated on net profit before taxes using the rate of 35%. Subtracting income taxes from net profit before taxes leaves net profit after taxes. Because depreciation is a non-cash expense, it is added back after determination of income tax liability for purposes of the cash flow estimate.

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22.9 Initial Capital Costs

Initial capital cost estimates as previously described in Item 21.0 served as input to the economic model. Of the total $3.005 billion estimated initial capital required, $63.582 million was set aside for spare parts, consumables, and initial fills, slated for expenditure in preproduction Year -1, but recaptured in Year 28 at end of mine life. The remaining capital costs were allocated to be expended over the four years preceding production: 5% in Year -4, 15% in Year -3, 30% in Year -2, and 50% in Year -1.

A down-payment of $28.187 million will be required on long-lead equipment in Year -5, but has been included in Year -4 expenditure. Initial capital costs also include $12.35 million required as down-payment on long-lead equipment to be purchased as sustaining capital and put into service during production Years 1 and 2.

22.10 Sustaining Capital Costs

Acquiring additional assets, increasing facility capacities, or replacing assets are considered sustaining capital costs over the life of the Project. Such expenditures fall into four main categories for the Project: mining, plant, geotechnical facilities, and Owner’s costs. Sustaining capital costs are summarized below in Table 22-7. In total, it was estimated that the Project would require $1.807 billion in sustaining capital over the LOM (including the estimated $431 million for the concentrator expansion).

Mining accounts for more than half of the total sustaining capital required with an estimated $953 million share of expenditures, most of which is accounted for by the purchase of primary mine equipment. Determining each piece of equipment’s useful life upon acquisition allows for its replacement capital cost to be scheduled in the last year of its useful life. Additional mining equipment will also be required to meet the demands of increased production capacity upon the addition of a third line to the concentrator plant.

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Table 22-7: Sustaining Capital Cost Summary

Estimate Subtotal Description ($000's) ($000's)

Primary Mine Equipment 910,380

Ancillary Mine Equipment 30,976

Plant Mobile Equipment & Light Vehicles 8,092 Mining 952,698 Plant & Mine Warehouse/Truck shop 500 Equipment

Pit Dewatering 2,750

Truck shop Expansion 7,471 Plant 438,705 Plant Expansion 431,234

Construction of Salar Embankment 24,146

Expansion of Jacking Header System 86,610

Tailings Booster Station 250

Upgrade Cyclone Station (production 120K- 443 180K t/d) Geotechnical Sanitary Landfill Expansion 1,096 310,585 Facilities Seepage Collection Pipeline Headers 2,382 (excavation, Geotextile, gravel, pipe)

Placement and compaction of sand in dam 140,566

TSF Spillway 1,950

Excavation of Salar 53,143

Mobile Equipment 16,365

Process mobile equipment 6,000

Concentrate Trucks 2,000

Geotec Mobile Equipment 27,090 Owner's Cost 104,709 Engineering/Office Equipment, Software, 500 Furniture

Mobile Radios 482

Railroad Equipment 52,272

Total Sustaining Capital Cost (S000's) $1,806,698

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The next largest demand on sustaining capital is realized in Years 6 and 7, when $431 million is required for the planned expansion of the mill and concentrator facilities to include a third production line. While the enlargement of the concentrator plant accounts for approximately 24% of sustaining capital over LOM, productive capabilities are increased by 50% for the remaining 21 years of mine life. Table 22-8 provides details of the concentrator expansion estimate, which is included in the total sustaining capital estimate above.

Table 22-8: Sustaining Capital for Concentrator Expansion

Estimate Subtotal Total Description ($000's) ($000's) ($000's)

Crushing, Conveying and Storage 4,865

Grinding and Concentrator 252,891 Direct Costs 258,374 Site and Off-Site Utilities 518

Site and Off-Site Development 100

Construction Temporary Facilities 10,300 Construction Construction Equipment 1,259 35,959 Indirect Costs Construction Camp, including catering 24,400

EPCM Services 38,800 87,080 Geotechnical Facilities E&CQA Contracted Third Party Inspections/Testing (QAQC)/Surveying 1,550 51,121 Indirect Costs Vendor Representatives/Commissioning Assistance 8,171

Initial Fills and Mine/Plant Equipment Spare Parts 2,600

Owner's Project & Construction Management 6,068 8,270 Indirect Costs Insurance 2,202

Freight, Duties & Taxes 12,910 Other 77,510 Contingency 64,600

Total Capital Cost for Concentrator Expansion ($000's) $431,234

22.11 Working Capital

Defined as the highest amount of funding needed during the initial operating period, working capital is used to cover expenses prior to the cumulative revenue exceeding the cumulative expenses, or the point at which the operation becomes self-sustaining in its cash flow.

The largest deficit of funds would occur in week seven of production, for an estimated amount of $54.95 million. This working capital cost was recorded in the cash flow model in Year 1, with recovery at the end of mine life in Year 28.

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22.12 Base Case Analysis

Results of this PEA estimate payback of initial capital investment to occur approximately 3.8 years after start of production. The base case financial model was developed from information described in this section. Based upon this information, the Project is estimated to have an after-tax IRR of 17.2%. NPV was calculated for the project using mid-period adjustment and various discount rates, the results of which are presented in Table 22-9. Assuming a discount rate of eight percent over an estimated mine life of 28 years, the after-tax NPV is estimated to be $2.087 billion.

Table 22-9: Net Present Value with Mid-Period Adjustment

Discount Rate 4% 6% 8% 10% 12%

NPV ($M) $4,610.1 $3,121.5 $2,087.3 $1,352.6 $820.5

22.13 Base Case Sensitivity Analysis

Table 22-10 reflects the sensitivities of IRR and NPV to changes in market price for copper, gold, and molybdenum; changes in capital costs; changes in operating costs expressed in units and on an annual basis; and changes in metallurgical recovery of copper, gold, and molybdenum. Changes for metal prices and operating and capital costs are expressed in 5% increments of negative and positive deviation from the Base Case values, while recovery is varied in 1% increments.

Table 22-10: Sensitivity Analysis of IRR and NPV

Base Case Variance -20% -15% -10% -5% Base +5% +10% +15% +20%

$/Lb Cu 2.20 2.34 2.48 2.61 2.75 2.89 3.03 3.16 3.30 IRR 10.8% 12.5% 14.1% 15.7% 17.2% 18.6% 20.0% 21.3% 22.6% NPV @ 8% ($M) 574.1 956.0 1,335.9 1,712.6 2,087.3 2,462.1 2,836.8 3,210.4 3,581.8 $/Oz Au 960.00 1,020.00 1,080.00 1,140.00 1,200.00 1,260.00 1,320.00 1,380.00 1,440.00 IRR 16.8% 16.9% 17.0% 17.1% 17.2% 17.3% 17.3% 17.4% 17.5% NPV @ 8% ($M) 1,984.9 2,010.5 2,036.1 2,061.7 2,087.3 2,112.9 2,138.5 2,164.1 2,189.7 $/Lb Mo 9.60 10.20 10.80 11.40 12.00 12.60 13.20 13.80 14.40 IRR 16.9% 16.9% 17.0% 17.1% 17.2% 17.2% 17.3% 17.4% 17.5% NPV @ 8% ($M) 2,000.2 2,022.0 2,043.8 2,065.5 2,087.3 2,109.1 2,130.9 2,152.6 2,174.4

Unit Operating Cost ($/T process feed) 8.31 8.79 9.28 9.77 10.26 10.75 11.24 11.73 LOM Operating Cost ($M) 13,711.6 14,518.2 15,324.7 16,131.3 16,937.9 17,744.4 18,551.0 19,357.6 IRR 18.9% 18.3% 17.8% 17.2% 16.6% 15.9% 15.3% 14.7% NPV @ 8% ($M) 2,567.8 2,407.7 2,247.5 2,087.3 1,927.1 1,767.0 1,605.9 1,444.6 Capital Cost ($M) 2,554.6 2,704.9 2,855.2 3,005.5 3,155.7 3,306.0 3,456.3 3,606.6 IRR 20.3% 19.2% 18.1% 17.2% 16.3% 15.5% 14.7% 14.0% NPV @ 8% ($M) 2,451.7 2,330.2 2,208.8 2,087.3 1,965.8 1,844.4 1,722.9 1,601.4

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Base Case Variance - 4% - 3% - 2% - 1% Base + 1% + 2% + 3% + 4% Copper Recovery 86.5% 87.4% 88.3% 89.2% 90.2% 91.1% 92.0% 92.9% 93.8% IRR 16.0% 16.3% 16.6% 16.9% 17.2% 17.4% 17.7% 18.0% 18.3% NPV @ 8% ($M) 1,795.4 1,868.4 1,941.4 2,014.3 2,087.3 2,160.3 2,233.3 2,306.2 2,379.2 Gold Recovery 61.2% 61.8% 62.4% 63.1% 63.7% 64.3% 65.0% 65.6% 66.2% IRR 17.1% 17.1% 17.1% 17.1% 17.2% 17.2% 17.2% 17.2% 17.3% NPV @ 8% ($M) 2,060.9 2,067.5 2,074.1 2,080.7 2,087.3 2,093.9 2,100.5 2,107.1 2,113.8 Molybdenum Recovery 54.4% 55.0% 55.5% 56.1% 56.7% 57.2% 57.8% 58.3% 58.9% IRR 17.1% 17.1% 17.1% 17.1% 17.2% 17.2% 17.2% 17.2% 17.2% NPV @ 8% ($M) 2,069.9 2,074.2 2,078.6 2,083.0 2,087.3 2,091.7 2,096.0 2,100.4 2,104.7

Graphical representations follow of the sensitivities of IRR and NPV to the incremental changes in metal prices in Figure 21-1 and Figure 21-2 and capital cost versus operating cost in Figure 21-3 and Figure 21-4, respectively. Also, IRR and NPV sensitivities are illustrated in Figure 21-5 and Figure 21-6 in relation to varying metallurgical recovery rates.

Figure 22-1: IRR Sensitivity Analysis – Metal Prices

Figure 22-2: NPV Sensitivity Analysis – Metal Prices

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Figure 22-3: IRR Sensitivity Analysis – Capital v Operating Costs

Figure 22-4: NPV Sensitivity Analysis – Capital v Operating Costs

Figure 22-5: IRR Sensitivity Analysis – Metallurgical Recovery

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Figure 22-6: NPV Sensitivity Analysis – Metallurgical Recovery

22.14 Economic Model

The cash flow model forecast immediately follows in Figure 22-7.

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Figure 22-7: Cash Flow Model Forecast

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23 Adjacent Properties

The Taca Taca Alto mineralized zone is located 4 km west of Taca Taca. It is hosted by a Paleozoic intrusive complex, which is intruded by granitic to rhyodacitic porphyries. The porphyritic intrusions associated with the mineralization were dated at 257 Ma (K/Ar, Rubinstein, 1997) and 260 Ma (Rio Tinto, 1999). This is much older than the mineralizing event at Taca Taca which is Oligocene (29 Ma). A 2.5 km by 3 km alteration zone is characterized by a potassic core overprinted by phyllic and advanced argillic alteration. Three mineralization stages have been recognized: early pyrite-bornite- chalcopyrite; pyrite-molybdenite related to the phyllic alteration phase; and, digenite-covellite related to supergene enrichment (Rubinstein et al, 2000).

The Qualified Persons have not verified the information concerning the Taca Taca Alto mineralized zone and this information is not necessarily indicative of the mineralization on the Taca Taca property.

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24 Other Relevant Data and Information

The Qualified Persons are unaware of any other data or information that would be relevant to this Technical Report which is not already contained in one of the existing sections of this Technical Report.

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25 Interpretations and Conclusions

This section presents the conclusions and recommendations of the Qualified Persons for the Project and this Technical Report.

25.1 Interpretations and Conclusions

Interpretations and conclusions of the Qualified Persons of the Taca Taca PEA are listed below:

 The results of the PEA indicate that the Taca Taca Project is a robust project at this stage of development demonstrating favourable economic potential that warrants further work toward the next stage of development, prefeasibility. The exploration program continues to demonstrate the potential for future growth of the resource.

 The sample preparation, security, and procedures followed by Lumina Copper are adequate to support a mineral resource estimate.

 Assay data provided by Lumina Copper were represented accurately and suitable for use in resource estimation.

 There are no environmental issues existing or anticipated that could materially impact the ability to develop the Taca Taca mine.

 There are no known factors related to metallurgical, environmental, permitting, legal, title, taxation, socio-economic, marketing, or political issues which could materially affect the mineral resource estimates.

 Safety factors and probabilities of failure (pf) estimated for the final pit slopes at Taca Taca are well within acceptable limits as defined by the state-of-practice rock mass strength considerations. The recommended overall pit slope templates are more conservative by rock mass strength because the designs are based on bench configurations controlled by structural fabric. Consequently, there is a future opportunity to optimize overall slopes by incorporating controlled blasting and/or single benching to steepen bench face angles as well as an opportunity to lessen the degree of pit slope depressurization.

 The metallurgical test work undertaken is reasonably extensive and suitable for this level of study. The comminution data are considered adequate for a conceptual milling circuit design. The design of the processing circuits is based on this test work data in conjunction with assumptions based on typical industry values.

 The Taca Taca mineralized material is of moderate competency and hardness, and amenable to grinding in a conventional SABC. The mineralogy is fine grained and test work indicates a requirement to re-grind to a fine particle size to achieve adequate liberation for flotation as is common within the industry.

 Recoveries over the LOM are expected to range from 88% to 92% for copper, 56% to 57% for molybdenum, and 61% to 65% for gold, which will be contained in the copper concentrate. LOM metal recoveries are expected to average approximately 90% for copper, 64% for gold, and 57% for molybdenum.

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 The cost to construct a 144 km transmission line has been included in the initial capital cost estimate. This power line will connect the Project to a 1,000 MW capacity line that was previously used to export power from the TermoAndes S.A. power plant located near the city of Salta. The Salta power plant is comprised of two 365 MW gas fired generating units that use gas sourced from the gas fields located in northern Salta Province. The plant currently has approximately 200 MW of excess generating capacity that is considered to be sufficient for the Project’s power requirements as defined in this PEA.

 A water supply and water balance analysis were completed for the Project. This analysis, supported by metallurgical test work using water with varying salinity levels, derived a processing make-up water flow sheet comprising a combination of high salinity water from the neighbouring salar, fresh water derived from wells, and desalinated brackish well water. Conceptual design, capital and operating costs for the construction and operation of a water treatment plant were completed for the PEA. Based on current studies, the processing water flow sheet shows that there are adequate sources of water available for the Project. Additional prospective areas have recently been identified near the Project that may be sources of additional fresh water which could further optimize the process water flow sheet.

 The Project has been designed to meet World Bank Guidelines for social and environmental management practices. Baseline studies completed to date have included surface and groundwater quality, acid rock drainage, meteorological, and social. Provisions have been made within the mine plan and operating costs to account for the environmental protection and rehabilitation of the Project once mining has been completed to meet the World Bank Guidelines.

 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 characterized as mineral reserves. There is no certainty that these inferred resources will ever be upgraded or that the results stated in this PEA will be realized. Mineral resources that are not mineral reserves have no demonstrated economic viability.

25.2 Risks and Opportunities

The following risks and opportunities associated with development of the Project have been identified by the Qualified Persons.

During the next phase of Project development, prefeasibility, a number of risks will be investigated further and possibly reduced or eliminated. Similarly, further investigation and evaluation of some opportunities may allow them to be incorporated in the Project.

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25.2.1 Risks  Long term depressed metals pricing, particularly for copper, together with the risks and uncertainties associated with metal price fluctuations.

 Political risks and uncertainties affecting legislation, regulatory requirements or general business climate, including for example, (i) potential changes in existing laws and potential imposition of more onerous laws in the future; (ii) potential for expropriation or nationalization of the Project; (iii) potential imposition of royalty or tax regimes adversely affecting the Project; or (iv) increased costs or difficulties associated with financing the Project.

 Shortage of skilled labor due to competing demands from the mining industry in general, and other mines in Argentina in particular.

 High inflation, substantial price escalation of Project equipment, bulk materials, and consumables, and maintenance of existing or implementation of additional monetary controls or restrictions on import by Argentina.

 Capital and operating cost escalations as Project plans and parameters change or are refined.

 Failure to obtain or maintain or a delay in obtaining necessary permits or approvals by government authorities.

 A structural geological model has not been developed for the pit. The current understanding of the Project structure is based on using average fault and joint orientations from the surface and downhole televiewer work. There is risk that unidentified structures exist which could negatively affect the pit slope stability presented in this Technical Report.

 The water treatment system will require 200 l/s of raw water feed with a TDS concentration of less than 50,000 mg/l. The approximate location of this raw water feed has been identified during the hydrogeologic investigation, but the specific quantity and quality is uncertain.

 The inability to construct WRSFs over the salar due to technical or permitting reasons would require alternative WRSF sites to be identified. The most likely alternative WRSF sites, identified to date, are to the southwest and west of the planned open pit. Using these alternative sites could materially increase waste rock haulage costs as these sites have higher dump crest elevations and longer horizontal distances.

 Should an alternative haulage system be employed (an opportunity listed below), flatter slope angles in the area of the alternative equipment may be required for engineering design.

 Diesel fuel is a significant component of the mine operating costs. Higher fuel prices could impact project returns given the stripping ratio, pit depth, and corresponding long haulage profiles.

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25.2.2 Opportunities

Substantial opportunities for improving the Project’s potential viability exist. They include:

 Higher metals pricing, particularly for copper, than used as a long term forecast in the financial model.

 Potential for expansion of mineral resources.

 Additional metallurgical optimization test work may result in higher metallurgical recoveries and/or better concentrate grades.

 More cost effective development with more detailed information.

 Increased field investigation and design may improve the design concept, construction methodology, and cost effectiveness of the salar excavation and brine containment.

 Additional investigation of rail transport options and negotiation with interested service providers (both rail and port) may lead to more cost effective transport of concentrates and project consumables.

 Preliminary evaluation has suggested that on-site production of an anode copper product, instead of copper concentrate, with saleable sulfuric acid byproduct, may improve potential project economic performance.

 There is some potential to expand mineral resources and, hence, the open pit in the northwest quadrant and on the east side further into the salar. Increasing the metal prices used to define mineralized material would also increase resources accessible by open pit, but at lower average head grades. Neither of these, however, would likely change the character of the Project.

 An alternative haulage system, such as in-pit crushing/conveying, could offer potential mine operating cost savings given the depth of the open pit. The north wall of Phase 2 would not be disturbed until about Year 18, offering a possible location for a semi-permanent conveyor line. A new conveyor could be integrated into the west wall of Phase 5 as a replacement, with a primary crusher foundation located in the south corner of the ultimate pit. This may eliminate about 300 m of lift for truck haulage of material.

 There is a future opportunity to optimize overall slopes by incorporating controlled blasting and/or single benching to steepen bench face angles as well as an opportunity to lessen the degree of pit slope depressurization.

 Identification of additional low TDS water sources could reduce water treatment plant requirements. Location of these sources closer to Taca Taca could reduce piping and pumping costs.

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26 Recommendations

Based on the results of the PEA, the Qualified Persons recommend that Lumina Copper complete a PFS to further define the Project in order to more accurately assess its technical and economic viability and to support permitting activities.

The tasks and estimate of the costs to complete the PFS are summarized below in Table 26-1.

Table 26-1: Estimated Cost for Recommended Future Work

Estimated Cost Task ($000)

2. Update resource model/estimate with additional exploration drilling $50 results, including QA/QC.

Subtotal Exploration and Resource Estimate $5,050

3. Complete Prefeasibility Study

a) Complete condemnation drilling for facility siting, including 5,000 m of $2,500 drilling ($500/m allowance; 21% IVA included).

b) Additional mine geotechnical data collection and engineering analysis, including 4,200 m of DD drilling at 5-7 locations ($500/m $2,700 allowance; 21% IVA included).

c) Revise mine design and production schedule based on updated resource model; upgrade accuracy of mining capital, operating, and $150 sustaining capital cost estimates to PFS level.

d) Complete additional hydrogeological field investigations, engineering analysis, and computer modeling in support of mine dewatering and $1,500 project water supply.

e) Complete further evaluation of new water feed to the water treatment $15 plant and upgrade capital and operating costs to PFS level.

f) Conduct additional metallurgical test work, primarily consisting of spatial, grade, and mineralogic variability testing of representative $750 samples from across the deposit, in numbers accepted by the industry as being statistically representative.

g) Complete process and infrastructure design and associated costs to $1,200 a PFS level.

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Estimated Cost Task ($000)

h) Complete field geotechnical and laboratory investigations; complete PFS level design and associated costs for geotechnical infrastructure $725 (TSF, WRSF, and roads).

i) Perform environmental baseline studies and permitting activities. $250

j) Complete an updated marketing study. $50

k) Complete an updated power supply study. $50

l) Update and upgrade logistics and transportation study to PFS level. $100

m) Perform study management and coordination, execution planning $450 and scheduling, Owner’s cost estimating, and economic evaluation.

Subtotal PFS Estimate $10,440

Total Estimated Cost for Recommended Future Work $15,490

Advancement of the PFS in accordance with the tasks listed above in Item 3 is not dependent on prior completion of the exploration and updated resource estimate shown in tasks 1 and 2 above.

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27 References

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(1999). Cerro del Cobre Project, Taca Taca Bajo. Rio Tinto Mining and Exploration.

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Plenge, C. (May 18, 2011). Taca Taca Project, Flotation Variability.

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28 Date and Signature Page

CERTIFICATE OF QUALIFIED PERSON

I, Robert Sim, P.Geo., do hereby certify that:

1. I am an independent consultant of: SIM Geological Inc. 6810 Cedarbrook Place Delta, British Columbia, Canada V4E 3C5 2. I graduated from Lakehead University with an Honours Bachelor of Science (Geology) in 1984. 3. I am a member, in good standing, of the Association of Professional Engineers and Geoscientists of British Columbia, License Number 24076. 4. I have practiced my profession continuously for 28 years and have been involved in mineral exploration, mine site geology and operations, mineral resource and reserve estimations and feasibility studies on numerous underground and open pit base metal and gold deposits in Canada, the United States, Central and South America, Europe, Asia, Africa and Australia. 5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. 6. I am responsible for the preparation of Item 14 and portions of Items 1, 7, 8, 9, 10, 25, 26, and 27 of the technical report titled “Taca Taca Copper/Gold/Molybdenum Project Preliminary Economic Assessment Report” dated May 24, 2013, with an effective date of April 9, 2013 (the “Technical Report”). 7. My most recent personal inspection of the Taca Taca property occurred on June 21 to 23, 2012 for a duration of 3 days. 8. I have had prior involvement with the property that is the subject of the Technical Report. I was a co-author of previous NI 43-101 technical reports titled “Taca Taca Technical Report, Puna (altiplano) Region of northwest Argentina”, effective date of October 9, 2008, “Taca Taca Porphyry Copper-Gold-Molybdenum Project, Argentina”, effective date of November 15, 2011, “Taca Taca Property, Porphyry Copper-Gold-Molybdenum Project, Argentina” effective date of April 11, 2012 “Taca Taca Property Porphyry Copper-Gold-Molybdenum Project, Argentina NI 43-101 Technical Report”, effective date of October 30, 2012. 9. As of as of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. 10. I am independent of Lumina Copper Corp. applying all of the tests in Section 1.5 of NI 43- 101. 11. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

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Dated this 24th day of May, 2013.

/signed and sealed by Robert Sim/

______

Robert Sim, P.Geo.

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CERTIFICATE OF QUALIFIED PERSON

I, Bruce M. Davis, PhD, FAusIMM, do hereby certify that:

1. I am an independent consultant of: BD Resource Consulting, Inc. 4253 Cheyenne Drive, Larkspur, Colorado, U.S.A. 80118 2. I graduated from the University of Wyoming with a Doctor of Philosophy degree in Statistics with an emphasis in Geostatistcs in 1978. I hold a Bachelor of Science degree in Mathematics and a Master of Science in Statistics from Brigham Young University. 3. I am a Fellow of the Australasian Institute of Mining and Metallurgy (Registration No. 211185). Further I am a member of the PDAC (#216302), the International Association for Mathematical Geology (#148), and the SME (#742830) (all in good standing). 4. I have practiced my profession as a geostatistician continuously for 33 years. I have training in the theory and practice of sampling particulate materials and have designed and executed mineral sampling programs since 1978. Similarly, by reason of my statistical training I have conducted assay quality control and data verification programs throughout my time as a practicing geostatistician. I have been involved in geostatistical studies, mineral resource and reserve estimations and feasibility studies on numerous underground and open pit base metal and gold deposits in Canada, the United States, Central and South America, Europe, Asia. 5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. 6. I am responsible for the preparation of Items 11 and 12, and portions of 1, 7, 8, 9, 10, 14, 25, 26, and 27 of the technical report titled “Taca Taca Copper/Gold/Molybdenum Project Preliminary Economic Assessment Report”, dated May 24, 2013, with an effective date of April 9, 2013, (the “Technical Report”). 7. My most recent personal inspection of the Taca Taca property occurred on June 21 to 23, 2012 for a duration of 3 days. 8. I have had prior involvement with the property that is the subject of the Technical Report. I provided geostatistical analysis in respect of and in some cases co-authored, prior NI 43-101 technical reports titled “Taca Taca Technical Report, Puna (altiplano) Region of northwest Argentina”, effective date of October 9, 2008, “Taca Taca Porphyry Copper-Gold- Molybdenum Project, Argentina”, effective date of November 15, 2011, “Taca Taca Property, Porphyry Copper-Gold-Molybdenum Project, Argentina” effective date of April 11, 2012 “Taca Taca Property Porphyry Copper-Gold-Molybdenum Project, Argentina NI 43-101 Technical Report”, effective date of October 30, 2012. 9. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. 10. I am independent of Lumina Copper Corp. applying all of the tests in Section 1.5 of NI 43 101.

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11. I have read NI 43-101 and Form 43-101F1 and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 24th day of May, 2013.

/signed by Bruce M. Davis/

______

Bruce M. Davis, PhD, FAusIMM

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CERTIFICATE OF QUALIFIED PERSON

I, William L. Rose, P.E., do hereby certify that:

1. I am the Principal Mining Engineer of: WLR Consulting, Inc. 9386 West Iowa Avenue Lakewood, Colorado 80232-6441, USA

2. I graduated from the Colorado School of Mines with a Bachelor of Science degree in Mining Engineering in 1977. 3. I am a Registered Professional Engineer in the State of Colorado (No. 19296), a Registered Professional Engineer in the State of Arizona (No. 15055) and a Registered Member of the Society for Mining, Metallurgy and Exploration, Inc. (No. 2762350RM) (all in good standing). 4. I have practiced my profession as a mining engineer continuously for 35 years since my graduation from college. I have been involved in open pit mine operations in both management and engineering positions, and have extensive experience in mine design and planning. I have conducted estimations of mineral resources and reserves, mine production schedules, equipment and workforce requirements, and capital and operating costs for numerous projects in North, Central and South America, Europe, Africa and Asia. 5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. 6. I am responsible for the preparation of Items 15 and 16, and portions of 1, 2, 21, 22, 25, 26, and 27 of the technical report titled “Taca Taca Copper/Gold/Molybdenum Project Preliminary Economic Assessment Report”, dated May 24, 2013, with an effective date of April 9, 2013, (the “Technical Report”). 7. My most recent personal inspection of the Taca Taca property occurred on January 20-21, 2011 for a duration of two days. 8. I have had no prior involvement with the property that is the subject of the Technical Report. 9. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. 10. I am independent of Lumina Copper Corp. applying all of the tests in Section 1.5 of NI 43- 101. 11. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 24th day of May, 2013.

/signed and sealed by William L. Rose/

______

William L. Rose, P.E.

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CERTIFICATE OF QUALIFIED PERSON

I, Kevin Scott, P. Eng., do hereby certify that:

1. I am Manager, Process and Studies for Ausenco Solutions Canada Inc. 855 Homer Street, Vancouver, BC V6B 2W2, Canada. 2. I am a graduate of University of British Columbia, Vancouver, Canada in 1989 with a Bachelor of Applied Science degree in Metals and Materials Engineering. 3. I am registered as a Professional Engineer in the Province of British Colombia (Licence # 25314) and the Province of Ontario (License # 90443342) (all in good standing). 4. I have worked as a Metallurgist continuously for a total of 23 years since my graduation from University. My relevant experience for the purpose of the Technical Report (defined below) includes: o Reviewing and preparing reports as a metallurgical consultant on a number of mining operations and projects for due diligence and financial monitoring requirements o Acting as a process engineer at three Canadian base metals mineral processing operations o Acting as a senior metallurgical engineer working for four multi-national engineering and construction companies on feasibility studies and in engineering design of mineral processing plants in Canada and South America o Acting as a senior process manager in charge of process design and engineering for a metallurgical processing plant in South America. 5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101) and certify that by reason of education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfil the requirements to be a “qualified person” for the purpose of this NI 43-101. 6. I am responsible for Items 1, 2, 3, 4, 5, 6, 13, 17, 18, 19, 22, 23, 24, and portions of 21, 25, 26, and 27 of the technical report titled “Taca Taca Copper/Gold/Molybdenum Project Preliminary Economic Assessment Report”, dated May 24, 2013, with an effective date of April 9, 2013, (the “Technical Report”). 7. I have visited the Taca Taca property on October 23 and 24, 2012. 8. I am independent of Lumina Copper Corp. applying the test set out in Section 1.5 of the NI 43-101. 9. I have had no prior involvement with the property that is the subject of the Technical Report. 10. I have read National Instrument 43-101 and Form 43-101F1, and confirm that the Technical Report has been prepared in compliance with that instrument and form. 11. To the best of my knowledge, information and belief, this technical report contains all the scientific and technical information that is required to be disclosed to make this Technical Report not misleading.

Dated this 24th day of May, 2013 /signed and sealed by Kevin C. Scott/ ______

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CERTIFICATE OF QUALIFIED PERSON

I, Scott C. Elfen, P.E. do hereby certify that:

1. I am the Global Lead Geotechnical and Civil Services of Ausenco Engineering Canada Inc., 855 Homer Street, Vancouver, BC V6B 2W2, Canada 2. I graduated from the University of California, Davis with a Bachelor of Science degree in Civil Engineering (Geotechnical) in 1991. 3. I am a Registered Civil Engineer in the State of California (No. C56527) by exam since 1996, and am also a member of the American Society of Civil Engineers (ASCE), Society for Mining, Metallurgy & Exploration (SME) that are all in good standing. 4. I have practiced my profession continuously for 19 years and have been involved in geotechnical, civil, hydrological, and environmental aspects for the development of mining projects; including feasibility studies on numerous underground and open pit base metal and precious metal deposits in North America, Central and South America, Africa and Australia. 5. I have read the definition of “qualified person” set out in National Instrument 43-101 (“NI 43- 101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101. 6. I am responsible for the preparation of Item 20, and portions of 1, 4, 17, 18, 21, 25, and 26 of the technical report titled “Taca Taca Copper/Gold/Molybdenum Project Preliminary Economic Assessment Report”, dated May 24, 2013, with an effective date of April 9, 2013, (the “Technical Report”). 7. My most recent personal inspection of the Taca Taca property occurred on January 21-22, 2011 for a duration of two days. 8. I have had no prior involvement with the property that is the subject of the Technical Report. 9. As of the effective date of the Technical Report, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading. 10. I am independent of Lumina Copper Corp. applying all of the tests in Section 1.5 of NI 43- 101. 11. I have read NI 43-101 and Form 43-101F1, and the Technical Report has been prepared in compliance with that instrument and form.

Dated this 24th day of May, 2013.

/signed and sealed by Scott C. Elfen/

______

Scott C. Elfen, P.E.

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