Cañariaco Project, Lambayeque Department, Peru, NI 43-101 ...

Cañariaco Project, Lambayeque Department, Peru, NI 43-101 Technical Report on Pre-feasibility Study Progress ReportPeruPeru

Prepared for:Candente Copper Corporation

Prepared by:David G. Thomas, P.Geo., MAusIMMJay Melnyk, P.Eng.Tony Lipiec, P.Eng.Alexandra Kozak, P.Eng.

Project No. 165270j

Effective Date: 18 January 2011

AMEC Americas Limited 111 Dunsmuir Street, Suite 400 Vancouver, B.C. V6B 5W3 Tel (604) 664-3030 Fax (604) 664-3057 www.amec.com

CERTIFICATE OF QUALIFIED PERSON

David G. Thomas, P. Geo. AMEC Americas Limited

111 Dunsmuir Street, Suite 400 Vancouver, B.C. V6B 5W3

Tel (604) 664-3030 Fax (604) 664-3057

[email protected] I, David G. Thomas, P.Geo., am employed as a Senior Geologist with AMEC Americas Ltd.

This certificate applies to the technical report entitled “Candente Copper Corporation, Cañariaco Project, Lambayeque Department, Peru, NI 43-101 Technical Report on Prefeasibility Study Progress Report” (the “Technical Report”), dated 18 January 2011.

I am a Professional Geologist registered in British Columbia (P.Geo NRL # 149114) and a member of the Australiasian Institute of Mining and Metallurgy (MAusIMM # 225250). I graduated in 1993 from Durham University, in the United Kingdom with a Bachelor of Science degree and in 1995 from Imperial College, University of London, in the United Kingdom with a Master of Science degree.

I have practiced my profession for over 15 years. In that time I have been directly involved in review of exploration, geological models, exploration data, sampling, sample preparation, quality assurance-quality control, databases, and mineral resource estimates for a variety of mineral deposits, including porphyry copper deposits.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43–101 Standards of Disclosure for Mineral Projects (NI 43–101).

I have visited the Canariaco Project between 7 and 11 June 2010.

I am responsible for Sections 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 14, 15, and 17, and those portions of the Summary, Conclusions and Recommendations that pertain to those sections of the Technical Report.

I am independent of Candente Copper Corporation as independence is described by Section 1.4 of NI 43–101.

I have previous involvement with the Project during 2010, and prepared the following technical report in 2010:

Thomas, D., and Lipiec, T., 2010: Candente Copper Corporation, Cañariaco Project, Lambayeque Department, Peru NI 43-101 Technical Report: technical report prepared by AMEC Americas Ltd for Candente Resource Corp., effective date 8 November 2010

I have read NI 43–101 and this report has been prepared in compliance with that Instrument.


As of the date of this certificate, 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.

“Signed and sealed”

David G. Thomas P.Geo

4 March 2011



Jay Melnyk, P.Eng. AMEC Americas Limited

111 Dunsmuir Street, Suite 400 Vancouver, B.C. V6B 5W3

Tel (604) 664-3030 Fax (604) 664-3057 [email protected]

I, Jay Melnyk, P.Eng, am employed as a Principal Mining Engineer with AMEC Americas Ltd.


I am a Professional Engineer in the province of British Columbia (P.Eng #25975). I graduated from the Montana Technical Institute with a Bachelor of Mining Engineering degree in 1988, and from the British Columbia Institute of Technology with a Diploma in Mining Technology in 1984.

I have practiced my profession for 22 years. I have been directly involved in open pit mining operations, and design of open pit mining operations in Indonesia, Canada, the United States, Chile, Peru and Mexico.


I have not visited the Canariaco Project.

I am responsible for Sections 19.1 to 19.4 of the Technical Report and those portions of the Summary, Conclusions and Recommendations that pertain to those sections.


I have been involved with the Canariaco Project since August 2010, during preparation of a Prefeasibility Study Progress Report and this Technical Report.




Jay Melnyk, P.Eng. Dated: 4 March 2011



Ignacy (Tony) Lipiec (P.Eng.) AMEC Americas Ltd.,

Suite 400, 111 Dunsmuir St Vancouver, BC., Canada

Tel: 604-664-3130; Fax: 604-664-3057

E-mail: [email protected] I, Ignacy (Tony) Lipiec (P.Eng.) am employed as a Principal Process Engineer with AMEC Americas Ltd.


I am a Professional Engineer in the province of British Columbia. I graduated from the University of British Columbia with a B.A.Sc. degree in Mining & Mineral Process Engineering, in 1985.

I have practiced my profession for 25 years, and have previously been involved with metallurgical design and process engineering for base metal and disseminated sulphide projects in North America and South America.


I have not visited the Cañariaco Project.

I am responsible for Section 16 and those portions of the Summary, Conclusions and Recommendations that pertain to that section of the Technical Report.


I have previous involvement with the Project during 2010, and prepared the following technical report in 2010:

Thomas, D., and Lipiec, T., 2010: Candente Copper Corporation, Cañariaco Project, Lambayeque Department, Peru NI 43-101 Technical Report: technical report prepared by AMEC Americas Ltd for Candente Resource Corp., effective date 8 November 2010





Ignacy (Tony) Lipiec (P.Eng.)

4 March 2011



Alexandra J. Kozak, P.Eng. AMEC Americas Limited

111 Dunsmuir Street, Suite 400 Vancouver, BC

Tel: (604) 664-4578 Fax: (604) 664-3057

[email protected] I, Alexandra J. Kozak, P.Eng., am employed as Manager, Process Engineering with AMEC Americas Limited.


I am a member of the Association of Professional Engineers and Geoscientists of British Columbia (APEGBC). I graduated from the University of Alberta with a Bachelor of Science degree in Mineral Process Engineering in 1985.

I have practiced my profession continuously since 1985 and have been involved in operations in Canada and Guyana and preparation of scoping, pre-feasibility, and feasibility level studies for gold, base metals and diamond properties in Canada, United States, Peru, Mexico, Mongolia, Ghana, and New Guinea. I am currently a Consulting Engineer and have been so since September 1996.


I have not visited the Canariaco Project.

I am responsible for Sections 1, 2, 3, 18, 19.5 to 19.10, 20, 21, 22, and 23 of the Technical Report.


I have been involved with the Canariaco Project since May 2010 as part of preparation of this Technical Report and as manager of the Prefeasibility Study Progress Report that the Technical Report is based on.

I have read NI 43–101 and this Technical Report has been prepared in compliance with that Instrument.


As of the date of this certificate, 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.


Alexandra J. Kozak, P.Eng.

Dated: 4 March 2011

IMPORTANT NOTICE

This report was prepared as a National Instrument 43-101 Technical Report for Candente Copper Corporation (CCC) by AMEC Americas Limited (AMEC). The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in AMEC’s services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended for use by CCC subject to the terms and conditions of its contract with AMEC. This contract permits CCC to file this report as a Technical Report with Canadian Securities Regulatory Authorities pursuant to National Instrument 43-101, Standards of Disclosure for Mineral Projects. Except for the purposes legislated under provincial securities law, any other uses of this report by any third party is at that party’s sole risk.

Candente Copper CorporationCañariaco Norte Deposit,

Lambayeque Department, PeruNI 43-101 Technical Report on Pre-feasibility Study Progress Report

Project No.: 165270 March 2011 TOC i

C O N T E N T S

1.0 SUMMARY ................................................................................................................................... 1-1 1.1 Principal Outcomes ......................................................................................................... 1-1 1.2 Terms of Reference ......................................................................................................... 1-2 1.3 Location and Access ....................................................................................................... 1-2 1.4 Mineral Tenure, Surface Rights, and Royalties ............................................................... 1-3 1.5 Permits and the Environment .......................................................................................... 1-3 1.6 Geology and Mineralization ............................................................................................. 1-4 1.7 History and Exploration ................................................................................................... 1-5 1.8 Drilling .............................................................................................................................. 1-5 1.9 Analyses and Quality Assurance/Quality Control ............................................................ 1-6 1.10 Data Verification .............................................................................................................. 1-7 1.11 Metallurgical Testwork ..................................................................................................... 1-7 1.12 Recoveries ....................................................................................................................... 1-8 1.13 Mineral Resources ........................................................................................................... 1-8 1.14 Mineral Resource Statement ........................................................................................... 1-9 1.15 Projected Mine Plan ...................................................................................................... 1-10 1.16 Proposed Process Route ............................................................................................... 1-11 1.17 Proposed Tailings Management .................................................................................... 1-11 1.18 Proposed Water Management ...................................................................................... 1-12 1.19 Proposed Workforce ...................................................................................................... 1-12 1.20 Infrastructure ................................................................................................................. 1-12 1.21 Projected Capital Costs ................................................................................................. 1-13 1.22 Projected Operating Costs ............................................................................................ 1-13 1.23 Financial Analysis .......................................................................................................... 1-14 1.24 Alternative Financial Cases ........................................................................................... 1-15 1.25 Exploration Potential ...................................................................................................... 1-16 1.26 Conclusions ................................................................................................................... 1-17 1.27 Recommendations ......................................................................................................... 1-17

2.0 INTRODUCTION .......................................................................................................................... 2-1 2.1 Terms of Reference ......................................................................................................... 2-1 2.2 Qualified Persons ............................................................................................................ 2-3 2.3 Effective Dates ................................................................................................................ 2-3 2.4 Previous Technical Reports............................................................................................. 2-4 2.5 References ...................................................................................................................... 2-5 2.6 Technical Report Sections and Required Items under Form 43-101F1 .......................... 2-5

3.0 RELIANCE ON OTHER EXPERTS .............................................................................................. 3-1 3.1 Mineral Tenure ................................................................................................................ 3-1 3.2 Surface Rights ................................................................................................................. 3-1 3.3 Royalties .......................................................................................................................... 3-2

4.0 PROPERTY DESCRIPTION AND LOCATION ............................................................................ 4-1 4.1 Property and Title in Peru ................................................................................................ 4-1

4.1.1 Mineral Tenure ................................................................................................... 4-1 4.1.2 Surface Rights .................................................................................................... 4-2 4.1.3 Royalties ............................................................................................................. 4-3



Project No.: 165270 March 2011 TOC ii

4.1.4 Environmental Regulations ................................................................................ 4-3 4.2 Mineral Tenure ................................................................................................................ 4-6

4.2.1 Tenure History .................................................................................................... 4-6 4.2.2 Current Tenure ................................................................................................... 4-6

4.3 Surface Rights ................................................................................................................. 4-1 4.4 Royalties .......................................................................................................................... 4-1 4.5 Permits ............................................................................................................................. 4-1 4.6 Environmental .................................................................................................................. 4-2 4.7 Socio-Economics ............................................................................................................. 4-3 4.8 Comment on Section 4 .................................................................................................... 4-4

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY ......................................................................................................................... 5-1 5.1 Access ............................................................................................................................. 5-1 5.2 Climate ............................................................................................................................. 5-1 5.3 Local Resources and Infrastructure ................................................................................ 5-1

5.3.1 Local Resources ................................................................................................. 5-1 5.3.2 Existing Infrastructure ......................................................................................... 5-2 5.3.3 Planned Infrastructure ........................................................................................ 5-2 5.3.4 Port ..................................................................................................................... 5-5

5.4 Physiography ................................................................................................................... 5-8 5.5 Comment on Section 5 .................................................................................................... 5-9

6.0 HISTORY ...................................................................................................................................... 6-1

7.0 GEOLOGICAL SETTING ............................................................................................................. 7-1 7.1 Cañariaco Norte Deposit Geology ................................................................................... 7-3

7.1.1 Lithologies .......................................................................................................... 7-4 7.1.2 Alteration ............................................................................................................ 7-6 7.1.3 Structures ......................................................................................................... 7-10

7.2 Prospects ....................................................................................................................... 7-11 7.2.1 Cañariaco Sur ................................................................................................... 7-12 7.2.2 Quebrada Verde ............................................................................................... 7-12

7.3 Comment on Section 7 .................................................................................................. 7-12

8.0 DEPOSIT TYPES ......................................................................................................................... 8-1 8.1 Comment on Section 8 .................................................................................................... 8-2

9.0 MINERALIZATION ....................................................................................................................... 9-1 9.1 Comment on Section 9 .................................................................................................... 9-2

10.0 EXPLORATION .......................................................................................................................... 10-1 10.1 Grids and Surveys ......................................................................................................... 10-1 10.2 Geological Mapping ....................................................................................................... 10-1 10.3 Geochemistry ................................................................................................................ 10-3 10.4 Pits and Trenches .......................................................................................................... 10-3 10.5 Geophysical Surveys ..................................................................................................... 10-3 10.6 Drilling ............................................................................................................................ 10-4 10.7 Bulk Density ................................................................................................................... 10-4 10.8 Geotechnical and Hydrogeology ................................................................................... 10-4 10.9 Other Studies ................................................................................................................. 10-5 10.10 Exploration Potential ...................................................................................................... 10-5 10.11 Comment on Section 10 ................................................................................................ 10-5



Project No.: 165270 March 2011 TOC iii

11.0 DRILLING ................................................................................................................................... 11-1 11.1 Drill Contractors ............................................................................................................. 11-1 11.2 Drill Methods .................................................................................................................. 11-1 11.3 Geological Logging ........................................................................................................ 11-4 11.4 Collar Surveys ............................................................................................................... 11-4 11.5 Down-hole Surveys ....................................................................................................... 11-4 11.6 Recovery........................................................................................................................ 11-5 11.7 Cañariaco Norte Drilling ................................................................................................ 11-5 11.8 Comment on Section 11 ................................................................................................ 11-6

12.0 SAMPLING METHOD AND APPROACH .................................................................................. 12-1 12.1 Geochemical Sampling .................................................................................................. 12-1 12.2 Pit and Trench Sampling ............................................................................................... 12-1 12.3 Core Sampling ............................................................................................................... 12-1 12.4 Quality Assurance and Quality Control ......................................................................... 12-1 12.5 Density/Specific Gravity ................................................................................................. 12-1 12.6 Comment on Section 12 ................................................................................................ 12-2

13.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY ........................................................ 13-1 13.1 Analytical Laboratories .................................................................................................. 13-1 13.2 Sample Preparation and Analysis ................................................................................. 13-2

13.2.1 CCC .................................................................................................................. 13-2 13.3 Quality Assurance/Quality Control Programs ................................................................ 13-3

13.3.1 CCC .................................................................................................................. 13-4 13.4 Databases ..................................................................................................................... 13-6 13.5 Sample Security ............................................................................................................ 13-6 13.6 Sample Storage ............................................................................................................. 13-6 13.7 Comment on Section 13 ................................................................................................ 13-7

14.0 DATA VERIFICATION ................................................................................................................ 14-1 14.1 Currie, 2004 ................................................................................................................... 14-1 14.2 MineFill, 2007 ................................................................................................................ 14-1 14.3 SRK, 2008 ..................................................................................................................... 14-1 14.4 AMEC, 2010 .................................................................................................................. 14-2 14.5 Comment on Section 14 ................................................................................................ 14-2

15.0 ADJACENT PROPERTIES ........................................................................................................ 15-1

16.0 MINERAL PROCESSING AND METALLURGICAL TESTING .................................................. 16-1 16.1 Metallurgical Testwork ................................................................................................... 16-1

16.1.1 Leach Tests ...................................................................................................... 16-1 16.1.2 Mineralogy ........................................................................................................ 16-2 16.1.3 Comminution Tests ........................................................................................... 16-3 16.1.4 Flotation Tests .................................................................................................. 16-3 16.1.5 Concentrate Generation ................................................................................... 16-6 16.1.6 Testwork Results .............................................................................................. 16-8

16.2 Recoveries ..................................................................................................................... 16-9 16.3 Proposed Process Route ............................................................................................... 16-9

16.3.1 Process Description .......................................................................................... 16-9 16.3.2 Crushing/Conveying ....................................................................................... 16-10 16.3.3 Coarse Ore Storage and Reclaim .................................................................. 16-10 16.3.4 Grinding .......................................................................................................... 16-14



Project No.: 165270 March 2011 TOC iv

16.3.5 Flotation .......................................................................................................... 16-15 16.3.6 Concentrate Dewatering ................................................................................. 16-16 16.3.7 Tailings Thickening and Disposal ................................................................... 16-16 16.3.8 Roasting and Stabilization .............................................................................. 16-17

16.4 Reagents and Grinding Media ..................................................................................... 16-20 16.4.1 Process Water ................................................................................................ 16-21

16.5 Process Control, Sampling, and Assay ....................................................................... 16-21 16.6 Concentrate Receiving and Loadout at Port Site ........................................................ 16-22 16.7 Proposed Tailings Management Facility ..................................................................... 16-22

16.7.1 Location .......................................................................................................... 16-22 16.7.2 Design Considerations ................................................................................... 16-23 16.7.3 Starter Embankment ....................................................................................... 16-24 16.7.4 TMF Embankment .......................................................................................... 16-24 16.7.5 Seepage Collection Impoundment ................................................................. 16-25 16.7.6 Construction Diversion Structures .................................................................. 16-25 16.7.7 Scorodite Management Facility ...................................................................... 16-26 16.7.8 Closure ........................................................................................................... 16-26

16.8 Comment on Section 16 .............................................................................................. 16-26

17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES ............................................. 17-1 17.1 Database ....................................................................................................................... 17-1 17.2 Topography ................................................................................................................... 17-1 17.3 Models ........................................................................................................................... 17-1 17.4 Composites .................................................................................................................... 17-3 17.5 Exploratory Data Analysis ............................................................................................. 17-3 17.6 Variography ................................................................................................................... 17-4 17.7 Grade Capping .............................................................................................................. 17-4 17.8 Estimation Methodology ................................................................................................ 17-5 17.9 Density ........................................................................................................................... 17-5 17.10 Model Validation ............................................................................................................ 17-6 17.11 Mineral Resource Classification .................................................................................... 17-6 17.12 Assessment of Reasonable Prospects of Economic Extraction .................................... 17-7 17.13 Mineral Resource Statement ......................................................................................... 17-9 17.14 Comment on Section 17 .............................................................................................. 17-11

18.0 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORT ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES .................................................................. 18-1

19.0 OTHER RELEVANT DATA AND INFORMATION ..................................................................... 19-1 19.1 Proposed Mining Operation ........................................................................................... 19-1

19.1.1 Dilution .............................................................................................................. 19-1 19.1.2 Pit Geotechnical Design Recommendations .................................................... 19-2 19.1.3 Pit Hydrogeology .............................................................................................. 19-2 19.1.4 Pit Optimization ................................................................................................ 19-4 19.1.5 Mine Production Schedule ............................................................................... 19-4

19.2 Waste Rock Management Facility ................................................................................. 19-8 19.3 Proposed Equipment ..................................................................................................... 19-9 19.4 Proposed Water Management .................................................................................... 19-11 19.5 Capital Costs ............................................................................................................... 19-13 19.6 Operating Costs ........................................................................................................... 19-14 19.7 Markets ........................................................................................................................ 19-16



Project No.: 165270 March 2011 TOC v

19.7.1 Treatment Charges ......................................................................................... 19-16 19.7.2 Concentrates .................................................................................................. 19-18 19.7.3 Acid ................................................................................................................. 19-18 19.7.4 Contracts and Agreements ............................................................................. 19-19

19.8 Taxation ....................................................................................................................... 19-19 19.9 Financial Analysis ........................................................................................................ 19-20

19.9.1 Basis of Base Case Financial Analysis .......................................................... 19-21 19.9.2 Results of Base Case Financial Analysis ....................................................... 19-22 19.9.3 Base Case Sensitivity Analysis ...................................................................... 19-22 19.9.4 Alternative Cases ........................................................................................... 19-25

19.10 Risks and Opportunities .............................................................................................. 19-26

20.0 INTERPRETATION AND CONCLUSIONS ................................................................................ 20-1

21.0 RECOMMENDATIONS .............................................................................................................. 21-1

22.0 REFERENCES ........................................................................................................................... 22-1

23.0 DATE AND SIGNATURE PAGE ................................................................................................ 23-1

T A B L E S

Table 1-1: Mineral Resource Statement for Cañariaco Norte at a 0.2% Cu Cut-off Grade (David Thomas P. Geo., Effective Date 8 November 2010) ........................................................... 1-10

Table 1-2: Summary of Capital Costs ................................................................................................... 1-13 Table 1-3: Average Operating Costs ..................................................................................................... 1-14 Table 1-4: Results of Financial Analysis for Base Case (base case is highlighted) ............................. 1-15 Table 1-5: Results of Financial Analysis for Alternative Cases ............................................................. 1-16 Table 2-1: QPs for the Report ................................................................................................................. 2-2 Table 2-2: Contents Page Headings in Relation to Form 43-101F1 Prescribed Item Headings ............ 2-6 Table 4-1: Summary of Environmental Requirements for Mining Exploration Programs ....................... 4-5 Table 4-2: Cañariaco Concessions held by Candente Copper Corp. (as at 26 November 2010) .......... 4-7 Table 7-1: Project Stratigraphic Column ................................................................................................. 7-4 Table 10-1: Exploration Summary Table ................................................................................................. 10-2 Table 11-1: Drill Summary Table............................................................................................................. 11-2 Table 11-2: Drill Hole Intercept Summary Table ..................................................................................... 11-6 Table 12-1: Bulk Density Values used in the 2010 Mineral Resource Estimate ..................................... 12-2 Table 16-1: Metallurgical Testwork Summary Table ............................................................................... 16-1 Table 16-2: Key Process Design Criteria .............................................................................................. 16-11 Table 17-1: Estimation Domains ............................................................................................................. 17-3 Table 17-2: Outlier Thresholds for Copper, Gold and Silver ................................................................... 17-5 Table 17-3: Optimization Parameters for Resource Pit Shell ................................................................. 17-8 Table 17-4: Copper Process Recoveries ................................................................................................ 17-8 Table 17-5: Marginal Cut-Off Calculation ................................................................................................ 17-8 Table 17-6: Mineral Resource Statement for Cañariaco Norte at a 0.2% Cu Cut-off Grade (David

Thomas P. Geo., Effective Date 8 November 2010) ......................................................... 17-10 Table 17-7: Mineral Resource Statement for Cañariaco Norte Showing Sensitivity to Various Cut-offs

(David Thomas P. Geo., Effective Date 8 November 2010). ............................................. 17-10



Project No.: 165270 March 2011 TOC vi

Table 19-1: Design Sector Slope Recommendations ............................................................................. 19-3 Table 19-2: LOM Production Schedule ................................................................................................... 19-6 Table 19-3: Mine Equipment Requirements .......................................................................................... 19-10 Table 19-4: Summary of Capital Costs ................................................................................................. 19-15 Table 19-5: LOM Mining Cost by Category ($000) ............................................................................... 19-16 Table 19-6: Average Annual Processing Costs .................................................................................... 19-17 Table 19-7: Summary of G&A Cost Estimate ........................................................................................ 19-17 Table 19-8: Base Case Metal Prices ..................................................................................................... 19-21 Table 19-9: Base Case Cashflow Analysis ........................................................................................... 19-23 Table 19-10: Results of Sensitivity Analysis for Base Case with NPV @ 8% .................................. 19-25 Table 19-11: Results of Financial Analysis for Alternative Cases .................................................... 19-26

F I G U R E S

Figure 2-1: Project Location Map ............................................................................................................. 2-2 Figure 4-1: Cañariaco Concessions held by Candente Copper Corporation ........................................... 4-8 Figure 5-1: Cañariaco Property Access Routes and Infrastructure .......................................................... 5-2 Figure 5-2: Planned Layout, Major Project Facilities ................................................................................ 5-3 Figure 5-3: Proposed Mine Site General Arrangement ............................................................................ 5-4 Figure 7-1: Regional Geology Map .......................................................................................................... 7-2 Figure 7-2: Regional Stratigraphic Column .............................................................................................. 7-3 Figure 7-3: Geological Map, Cañariaco Norte .......................................................................................... 7-7 Figure 7-4: Geological Drill Section 9,326,300N ...................................................................................... 7-8 Figure 7-5: Geological Drill Section 690,800E ......................................................................................... 7-9 Figure 7-6: Prospect Location Plan and Geological Map ....................................................................... 7-11 Figure 9-1: Example Vertical Section (9,326,300N) with 15 m Composites Coloured by Cu (%)

Ranges Overlapping Lithological Interpretation ..................................................................... 9-3 Figure 9-2: Example Vertical Section (9,326,300N) with 15 m Composites Coloured by Au (ppb)

Ranges ................................................................................................................................... 9-4 Figure 9-3: Example Vertical Section (9,326,300N) with 15 m Composites Coloured by Ag (g/t)

Ranges ................................................................................................................................... 9-5 Figure 11-1: Drill Hole Location Plan, Cañariaco Norte ........................................................................... 11-2 Figure 11-2: Drill Hole Location Plan, Cañariaco Sur and Quebrada Verde ............................................ 11-3 Figure 16-1: Process Overview – General Diagram ............................................................................... 16-13 Figure 16-2: Process Overview – Block Diagram ................................................................................... 16-13 Figure 16-3: Partial Roasting and Arsenic Removal Circuit ................................................................... 16-19 Figure 19-1: Pit Design Sectors ................................................................................................................ 19-3 Figure 19-2: Proposed Pit Phases (figure looks northeast) ..................................................................... 19-5 Figure 19-3: Material Movement by Period .............................................................................................. 19-7 Figure 19-4: Mill Feed Characteristics by Year ........................................................................................ 19-7 Figure 19-5: Proposed Layout, WRF ........................................................................................................ 19-8 Figure 19-6: Layout Plan, Proposed Water Management Structures .................................................... 19-12 Figure 19-7: Graph of Sensitivity Analysis Results for Base Case ........................................................ 19-25



Project No.: 165270 March 2011 Page 1-1

1.0 SUMMARY

AMEC Americas Limited (AMEC) was commissioned by Candente Copper Corporation (CCC) to prepare an independent Qualified Person’s Review and NI 43-101 Technical Report (the Report) on the Pre-Feasibility Study Progress Report which includes a Preliminary Assessment (PA) update for the wholly-owned Cañariaco Norte Copper Project (the Project) located in the Lambayeque Department, Peru. The Project contains the Cañariaco Norte copper deposit, and the Cañariaco Sur and Quebrada Verde prospects.

CCC will be using the Report in support of disclosures in the CCC press release dated 7 March 2011, entitled “Candente Copper Files NI 43-101 Technical Report and Updates Key Financial Results for the Cañariaco Norte Copper Project Pre-Feasibility Study Progress Report”, and the CCC press release dated 18 January 2011, entitled “Candente Copper Announces Positive Pre-Feasibility Progress Report for the Cañariaco Norte Copper Project”.

1.1 Principal Outcomes

• After-tax NPV of US$1,063.4 million for base case with US$2.25 per pound copper, US$1,015 per ounce gold, US$15.85 per ounce silver, and eight percent discount rate

• After-tax IRR of 18.8 percent for base case with US$2.25 per pound copper, US$1,015 per ounce gold, and US$15.85 per ounce silver

• Payback of preproduction capital in 2.9 years (on a pre-tax basis)

• Cash operating cost of US$0.99 per pound of copper including all on-site and off-site costs, toll treatment/refinery (TCRC) charges, net of by-product credits

• Average metal production of 262 million pounds (119,000 tonnes) copper per year, 37,000 ounces of gold per year, and 850,000 ounces of silver per year

• Average production of 295 million pounds (139,000 tonnes) copper per year for the first three years of production

• Pre-production capital cost of US$1.437 billion based on leased mining equipment and including contingency of 20 percent

• All in capital cost of US$1.565 billion based on leased mining equipment and including working capital, life-of-mine sustaining capital, and closure cost

• Processing rate of 95,000 tonnes per day using conventional crush/grind and flotation technology




• Waste to ore strip ratio of 0.98 to 1

• Average life-of-mine metal recoveries of 89.7 percent for copper, 55 percent for gold and 50 percent for silver

• Concentrate grades average approximately 30 percent copper, three grams per tonne gold and 73 grams per tonne silver

• 22 year mine life, with potential for extension by mining additional resources identified below proposed pit

• Cañariaco Norte is located at a moderate elevation with pit centroid and process plant at approximately 3,000 metres above sea level;

• Connection to national power grid; distance of approximately 57 kilometres

• New access road to major paved highway; distance of approximately 42 kilometres

• Significant potential for discovery of additional resources at nearby Cañariaco Sur and Quebrada Verde targets.

1.2 Terms of Reference

AMEC was requested to complete a Pre-Feasibility Study for CCC on the Cañariaco Norte Project. The study that is the subject of this Report is considered to be a Pre-feasibility Study Progress Report, because:

• Additional geotechnical drilling and rock quality assessment is required to complete the open pit slope design to a level consistent with pre-feasibility requirements; this drilling has been planned

• Geotechnical investigations are required in the area where above-ground structures are proposed as part of tailings management facility (TMF) design. Information from the planned geotechnical program may result in modifications to the design.

The mine plan is based on Measured and Indicated mineral resources only, with Inferred mineral resources considered as waste. The economic analyses based on these mineral resources, although modelled to a level consistent with a pre-feasibility study, are considered to be a PA under NI 43-101.

1.3 Location and Access

The Cañariaco Project is approximately 700 kilometres northwest of Lima, the capital of Peru, and approximately 110 kilometres to the northeast of the city of Chiclayo. Project access is either via the Pan-American Highway (700 kilometres, 11-hour trip),




or one of the thrice-daily commercial airline flights can be taken from Lima to Chiclayo. The route from Chiclayo to the Cañariaco Project is currently a 149 kilometre, six-hour trip along mostly unimproved roads via Incahuasi.

1.4 Mineral Tenure, Surface Rights, and Royalties

The Project currently consists of 16 Cañariaco concessions, covering a total area of 13,400 hectares and four Jehumarca claims, of 3,600 hectares, for a total Project area of 16,390 hectares. Mining legislation in Peru does not require location of concession boundaries on the ground. The boundary limits are defined by UTM coordinates. To maintain the property in good standing, annual property payments of $17,500 commenced in June 2002. AMEC was provided with information from legal advisors retained by CCC that all required property payments have been made, and the concessions are in good standing. Providing the annual property payments are made in a timely manner, the concessions will not expire.

The owner of the surface rights is the Comunidad Campesina San Juan Bautista de Cañaris (the Campesina). Under Peruvian Law 24656, CCC is required to have a valid agreement with the Campesina to undertake exploration. The previous agreement expired in December 2009, and CCC is in the process of negotiating a new, three-year, access agreement.

There are no royalties payable to any third parties on any of the concessions that currently comprise the Project.

It is expected that the Project will incur a three percent royalty payable to the Government of Peru once production commences. AMEC notes that if changes to the royalty legislation currently being considered by the Peruvian Government are made, then the royalty burden may increase.

1.5 Permits and the Environment

Exploration activities to date have been undertaken in accordance with the appropriate Peruvian regulations. A number of permits will be required to support project development and operation, and the main permit requirements have been identified to support any proposed construction and mining operations.

CCC is carrying out environmental baseline studies and monitoring under the direction of AMEC Earth and Environmental of Lima, Peru. Baseline data collected in 2007 indicated a diverse flora and fauna were present in the Project area, and identified several new flora and fauna species. AMEC notes that in many areas of Peru the




biodiversity assessments are not yet complete, and therefore it is not uncommon for detailed environmental impact investigations to identify new species. CCC has instigated two conservation projects that encompass the relocation of protected orchids found at the site and the identification of an area for conservation and environmental management.

A social and environmental impact assessment study framework has been prepared and is being fulfilled. Studies include determining the geochemical characteristics of the waste and mineralized rock, assessing the vicinity of the planned mine area for potential water supplies, and reviewing social and community conditions. Once these studies are completed to levels in accordance with the requirements of the Peruvian Government and meet the guidelines set out in the Equator Principles, then the process of obtaining all the required environmental, construction and mining permits will be undertaken.

At this stage of the Project, environmental liabilities are limited to those expected for an exploration-stage project, and include drill pads and access roads.

1.6 Geology and Mineralization

The Cañariaco Norte polyphase intrusive complex is approximately 1.7 kilometres in strike extent, 1.1 kilometres wide, and has been drill-tested to a depth of 770 metres. The deposit remains open at depth. It is a typical example of an Andean porphyry copper complex.

The country rocks hosting the intrusive complex comprise andesitic pyroclastic rocks and flows on the western, northern and southern sides of the intrusive complex where elevations are higher, and comprise dacite and rhyolite volcanic rocks on the eastern side of the intrusive complex where elevations are lower. Host rocks have been intruded along faults and zones of structural weaknesses by three main intrusive phases. The intrusions are cut by three late mineral breccias. Intrusive units comprise approximately 55–60 percent of the deposit, breccias approximately 30–35 percent and pre-mineral volcanic rocks approximately 5–10 percent. The deposit is capped by a leached zone with a variable thickness ranging from 0 metres to 120 metres thick, thick and averaging approximately 40 metres to 50 metres in thickness, where copper mineralization has been leached out.

The primary copper minerals are chalcopyrite, covellite, and chalcocite. Other copper sulphide minerals that may be present include enargite, bornite, and tennantite. Copper sulphide minerals are mainly associated with pyrite and other sulphide gangue minerals. Copper minerals and pyrite are disseminated, veined, and fracture-hosted, with copper grades directly related to the intensity of fracturing and the alteration type




and intensity. Higher grades are associated with potassic, phyllic, and argillic alteration and less commonly with propylitic and silicic alteration.

1.7 History and Exploration

Prior to CCC’s involvement, exploration was conducted by the Peruvian Servicio Nacional de Geología y Minería (Ingemmet), Placer Dome Exploration Inc. (Placer Dome), and Billiton Exploration and Mining Perú B.V. (Surcusal Peruana), from 1967 to 2000. Work completed by these companies included stream sediment sampling, geological mapping, rock chip and grab sampling, trenching and pitting, induced polarization (IP), resistivity, and ground magnetic surveys, petrographic studies, core drilling, mineral resource estimation, and very preliminary leach testwork.

Candente Copper Corp., through its wholly-owned Peruvian subsidiary Exploraciones Milenio S.A. (EMSA), acquired 100 percent ownership of the Project in February 2002. Since that date, CCC has completed geological mapping, prospecting, ground magnetic, resistivity and magnetic geophysical surveys, rock chip sampling, petrographic studies, bulk sampling for metallurgical testing, re-logging and re-sampling of historic drill core, core drilling, and mineral resource estimation.

The exploration programs discovered the Cañariaco Norte copper deposit, and the Cañariaco Sur and Quebrada Verde prospects.

A preliminary assessment (PA) was undertaken in 2006 on the Cañariaco Norte deposit, and was updated in 2008. The updated PA in this report supersedes the 2008 PA, and the 2008 PA is considered to be no longer relevant.

In November 2010, CCC released an updated mineral resource estimate; this estimate forms the basis for the 2011 PA update.

1.8 Drilling

A total of 263 core holes (75,672.31 m), including geotechnical, metallurgical, and hydrogeological drilling for all three targets, Cañariaco Norte, Cañariaco Sur, and Quebrada Verde, have been completed.

Core was logged for geological and geotechnical parameters. Drill collar locations were picked up by a surveyor, using a total station instrument. Down-hole surveys were performed using either a Pajari, Sperry Sun, or Reflex EZ-Shot instrument.

Drill core generated by Ingemmet and Placer Dome was halved by personnel from these operators; there is no information as to the typical sample intervals taken by




either company. The Surcusal Peruana and CCC drill core was halved and sampled on 2 m intervals by personnel from these operators.

A total of 9,424 bulk density readings were taken by CCC personnel on core, and an additional 550 specific gravity determinations were performed by ALS Chemex (Lima).

1.9 Analyses and Quality Assurance/Quality Control

Several primary assay laboratories have been used for routine analyses over the Project history. Ingemmet used the Plenge Laboratory in Lima and the Ingemmet internal laboratory, also located in Lima. Placer Dome utilized the SGS-XRAL (SGS) laboratory in Lima. For the Surcusal Peruana sampling, sample preparation was undertaken by ALS Chemex in Trujillo, and primary analysis by ALS Chemex in Lima.

Limited information is available on the sample preparation and analysis methods for Ingemmet or Placer Dome. Ingemmet samples were analyzed for copper and molybdenum, and more rarely gold and silver, using a colorimetric analytical method. Surcusal Peruana samples were 200 gram splits of a one kilogram, 200 mesh homogenized sample. A split from each sample pulp was assayed for gold (fire assay with atomic absorption finish, 10 parts per billion detection limit) and copper, lead, zinc, molybdenum and arsenic (multi-acid, total digest), with an atomic absorption (AA) finish for each element. SGS completed check assays on a split of one in 20 pulps using the same analytical procedures as the initial analysis performed by ALS Chemex.

Activation-Skyline Laboratories (Actlabs) in Lima, Peru performed all of the sample preparation and the majority of the analyses for the CCC programs. Inductively-coupled plasma (ICP) analyses were performed by the Ancaster, Canada, Actlabs laboratory. Some analyses for the re-analysis of pre-2008 core samples for gold and ICP were undertaken by ALS Chemex in Lima. ACME Laboratories (ACME), Lima were used as a check laboratory for pulp analyses.

Each CCC sample was subject to total copper and sequential copper leaching analysis, using a three-acid digest and AA finish. Gold used an aqua regia digest, with a fire assay and AA finish. Depending on the sample, a 36-element suite was analysed by Actlabs using inductively-coupled plasma optical emission spectrometry (ICP-OES), or a 33-element suite was analysed by ALS Chemex using an ICP atomic emission spectroscopy (AES) method.

There is limited currently-available information on any quality assurance/quality control (QA/QC) programs for the Ingemmet and Placer Dome work programs. Surcusal




Peruana used blanks, standard reference materials (SRMs) and check assays. The CCC QA/QC program used field, pulp and coarse reject duplicates, blanks, and SRMs.

1.10 Data Verification

All data in the field are recorded in written form in field books, log books, sample sheets, logging forms or shipping forms. All field data is hand-entered into Excel tables. Data from third parties such as laboratories or survey contractors are generally supplied in digital and printed form. All data are verified by CCC personnel.

A number of data verification programs and audits have been performed over the Project history, primarily in support of technical reports. No errors or omissions were noted during these reviews.

Drill data collected from the Ingemmet, Placer Dome, and Surcusal Peruana campaigns were re-logged by CCC personnel, and nine of the drill holes have been re-assayed. Based on the correlations between the historical grades and the CCC re-assay grades, all of the historical data were accepted into the final database.

Three pairs of twinned holes were drilled by CCC to verify grade uniformity at short distances. In general, similar average grades were noted over the same depth intervals.

1.11 Metallurgical Testwork

Initial metallurgical testwork was focused on leachable copper. However, as the Project has developed, and the mineralization was confirmed as primarily sulphide, conventional milling was determined to be a more appropriate process. Testwork has included Qemscan examination, comminution and variability comminution tests, tests on the effects of grind sizes, collectors and pH, sulphidization, cleaner flotation tests and locked cycle tests. Concentrates were also produced for proof-of-concept testing to show that As and Sb levels in concentrate could be reduced by using a roaster step, resulting in an enhanced concentrate which would have low or no penalty elements and would be broadly marketable.

The currently-envisaged process route uses conventional technology and proven equipment. The route will include crushing and conveying, rougher flotation, regrinding, cleaning, thickening, filtration, concentrate roasting, and agglomeration. Pyrite will be floated and removed prior to tailings discharge. Gas from the roaster step, which contains sulphur dioxide, will be treated to produce upgraded, saleable, sulphuric acid.




1.12 Recoveries

Recoveries used to support Mineral Resource estimation and plant design were variable for copper, based on the copper grade, ranging from 58.8 percent at 0.1 percent copper to 93.1 percent at 0.6 percent copper, 55 percent for gold and 50 percent for silver.

1.13 Mineral Resources

The estimation database comprises 225 core holes (67,485.06 m of core, of which 66,584.29 m have assay intervals) from the Ingemmet, Placer Dome, Surcusal Peruana, and CCC drill programs and has a cut-off date of for information to be included in the Report of 18 July, 2010.

Lithological and alteration interpretations were provided as sections by CCC; AMEC created bench polygons and models from these. Estimation domains were defined following evaluation of statistical distributions of lithological and alteration units. AMEC created a total of seven domains for copper and four domains for gold and silver. Domains for arsenic, which is a contaminant in the metallurgical process, were based on low- and high-grade populations considering a 250 ppm arsenic threshold as the upper limit for low-grade arsenic data.

AMEC regularized the assay intervals by compositing the drill hole data into six metre lengths using no geological or domain boundaries. Exploratory data analysis comprised basic statistical evaluation of the six metre composites for copper, gold, silver and arsenic. Sage2001 software was used to construct down-hole and directional correlograms for the estimation domains for copper, gold, silver and arsenic. Probability plots were evaluated to define grade outliers for copper, gold, and silver and by estimation domains. Copper, gold, and silver grades were capped in each domain; arsenic values were not capped.

The block model consists of regular blocks (15 metres by 15 metres by 15 metres) and no rotation is used. The block size was chosen such that geological contacts are reasonably well reflected and to support an open pit mining scenario.

AMEC estimated copper, gold, silver and arsenic grades by estimation domains using ordinary kriging (OK) interpolation for the majority of domains. Inverse distance weighting to the second power (ID2) was used to interpolate gold and silver in domains where variography was not considered sufficiently robust. Indicator kriging was also used to estimate the probabilities of low- and high-grade arsenic values within blocks. Density values were assigned to blocks based upon the lithological codes.




The Cañariaco Norte block model was validated to ensure appropriate honouring of the input data, including construction of a nearest neighbour (NN) model to validate the OK model. No biases or errors were noted with the model.

AMEC used the following criteria to pre-classify blocks into categories as:

• Measured mineral resources: require composites from a minimum of three drill holes within 75 metres radius from a block centroid, or samples from two drill holes with the closest sample within 25 metres of the block centroid.

• Indicated mineral resources: samples from a minimum of two drill holes within 110 metre distance of the block centroid.

Blocks that were not classified as Measured or Indicated categories, but had a composite within 135 metres from the block centroid were classified as Inferred. Remaining blocks were not classified. A semi-automated process was used to smooth the initial classification and avoid islands or isolated blocks of different categories.

Reasonable prospects of economic extraction were assessed by applying preliminary economic constraints within an open pit shell. Mining and process costs, as well as process recoveries were defined from on-going AMEC studies for the Project. AMEC defined a cut-off value of 0.2 percent copper for reporting mineral resources from these parameters.

1.14 Mineral Resource Statement

Mineral Resources for the Project were classified under the 2005 CIM Definition Standards for Mineral Resources and Mineral Reserves by application of a cut-off grade that incorporated process-related operating costs and recovery parameters, and constraint of the Mineral Resources to a pit shell based on commodity prices, recovery parameters and operating costs. Mineral Resources are tabulated in Table 1-1.

The Qualified Person for the Mineral Resource estimate is David Thomas, P.Geo. Mineral resources are reported at a cut-off grade based on the assumption of long-term metal prices as follows: copper price of US$2.50 per pound, a gold price of US$1,035 per ounce and a silver price of US$17.25 per ounce, and have an effective date of 8 November 2010. Mineral resources that are not mineral reserves do not have demonstrated economic viability.




Table 1-1: Mineral Resource Statement for Cañariaco Norte at a 0.2% Cu Cut-off Grade (David Thomas P. Geo., Effective Date 8 November 2010)

Grade Contained Metal

Category Tonnage

Mt Cu %

Au g/t

Ag g/t

Copper (Blb)

Gold (Moz)

Silver (Moz)

Measured 406.6 0.44 0.07 1.9 3.977 0.958 24.656 Indicated 596.5 0.38 0.06 1.6 4.964 1.081 30.403 Measured + Indicated 1,003.0 0.40 0.06 1.7 8.941 2.039 55.059 Inferred 293.3 0.33 0.05 1.5 2.165 0.448 13.547

Notes to accompany Mineral Resource Table: 1. Mineral resources that are not mineral reserves do not have demonstrated economic viability. 2. Mineral Resources base case is reported at a 0.2% Cu cut-off grade 3. Mineral Resources are reported as undiluted. 4. A Lerchs–Grossmann pit shell was used to constrain the Mineral Resources to assess reasonable prospects

of eventual economic extraction. 5. Mineral Resources are reported using assumed long-term prices as follows: copper price of US$2.50/lb,

gold price of US$1,035/oz and silver price of US$17.25/oz. 6. Rounding as required by reporting guidelines may result in apparent summation differences between

tonnes, grade and contained metal content. 7. Tonnage and grade measurements are in metric units. Contained gold ounces are reported as troy ounces,

contained copper pounds as imperial pounds.

1.15 Projected Mine Plan

The mine plan is based on diluted Measured and Indicated mineral resources contained in a Lerchs-Grossmann (LG) optimized pit shell generated at metal prices of $2.25 per pound copper, $930 per ounce gold, and $15 per ounce silver.

Large-scale conventional bulk open pit mining was selected for Cañariaco, featuring 311 millimetre diameter blast hole drills, 43 cubic metre rope shovels, and 290 tonne haul trucks working on 15 metre benches. A single pit with internal phases will provide process feed at a rate of 95,000 tonnes per day.

Diluted net smelter return (NSRD) values were calculated for cut-off application and block valuation during pit optimization. An average mining cost of $1.42 per tonne mined was estimated. A combined mill feed-based cost of US$3.57 per tonne milled was used for pit optimization: this also represents the marginal breakeven cut-off grade. The input parameters utilized for the mine plan development included a target mine life between 18 years and 22 years, a life-of-mine (LOM) head grade of no less than 0.40 percent copper, and a strip ratio of less than 1:1.

The mine plan results in a total of 728.2 million tonnes of mill feed and 713.5 million tonnes of waste (0.98:1 strip ratio) over a 22-year mine operating life. Average life-of-mine head grades to the process plant will be 0.40 percent copper, 0.067 grams per tonne gold and 1.71 grams per tonne silver. The mine grades during the first three years of production will be higher, with average feed grades of 0.48 percent copper, 0.086 grams per tonne gold and 2.14 grams per tonne silver.




A single elongated waste rock management facility (WRMF) will be developed adjacent to the pit.

1.16 Proposed Process Route

The process design is for a concentrator with a nominal processing capacity of 95,000 tonnes per day of mineralization from the open pit. The mineral processing and the roasting/acid plants are based on conventional technology and industry-proven equipment.

Run-of-mine (ROM) ore from the open pit will be crushed and conveyed to the concentrator where the ore will be ground to liberate the mineral values from the host rock and then separated by flotation. The bulk copper-silver-gold sulphide concentrate produced will be filtered and introduced into the roasting plant. There the concentrate will undergo a partial oxidative roast, which will remove the arsenic and part of the sulphur into the gaseous phase. The gas will be scrubbed to capture particulate matter and the arsenic into solution. This solution will report to a stabilization circuit that will remove any valuable copper and convert the arsenic into scorodite, which will be filtered and transferred to the tailings management facility (TMF).

After removal of the arsenic, the gas containing sulphur dioxide will be processed through a modular plant to produce saleable sulphuric acid.

The copper precipitate and solid calcine produced from the roaster and its cyclones will be agglomerated and stockpiled. Trucks will transport the combined concentrate to the port facilities, where it will be placed on ocean-going vessels for transport to overseas smelters.

1.17 Proposed Tailings Management

The TMF presented in this report is designed to manage tailings for mine operations at a processing rate of 95,000 t/d for 19 years as per the preliminary mine plan. Subsequent to completion of this facility design, an updated mine plan was developed that extended the mine life to 22 years. To accurately reflect this change, the TMF design will be updated during the next phase of project study. However, the current facility represents a reasonable assessment of the associated costs for the project based on the known conditions.

The main components of the TMF are as follows:

• Starter embankment.




• TMF embankment • Seepage collection (SC) impoundment • Construction diversion structures • Scorodite management facility.

Design criteria utilized are commensurate with appropriate governance literature and/or as appropriate, industry standards for this level of study. The facility is proposed to be constructed in the Quebrada Yerma valley, following a trade-off study on locations.

Water from the TMF will be reclaimed to the process plant during operations. The process plant will require 166 cubic metres per day of reclaim water from the TMF and 18 cubic metres per day of fresh water make-up to replace water losses from operations.

1.18 Proposed Water Management

The surface water management plan for the Cañariaco Norte project will preserve to the maximum extent possible the “non contact” status of surface waters. Waters that become affected by project activities, or “contact” waters, will be contained, and if necessary, treated, so that waters that are released to the environment always meet the applicable water quality regulatory guidelines. A system of impoundments, embankments, diversions, and spillways will be developed immediately as construction commences to handle run-off from construction related activities. This system will continue to be developed through construction and operations activities to ensure water discharged to the environment always meets the guidelines.

1.19 Proposed Workforce

The total workforce is estimated at 600 persons. The work schedule assumes production will operate 24 hours a day, seven days a week, 365 days a year.

1.20 Infrastructure

The mine site facilities are divided into three general areas: the mine, the plant site, which includes buildings and structures for repair and maintenance of mine and plant equipment repair and maintenance, and the camp area, which includes facilities for personnel accommodations, administration, and security. Other support facilities and services include site access, power supply and distribution, water supply, explosive storage and handling, communications systems and waste disposal facilities at the site.




Project infrastructure will also include facilities at the Marine Terminal Muchik (MTM), a planned development by Lumina Copper Corporation (Lumina) in northern Peru. Here, CCC will have its own dedicated concentrated receiving, storage, and reclaim facilities and will share general site services and ship berthing/loadout equipment with Lumina.

1.21 Projected Capital Costs

The capital cost estimate for the financial model base case assumes open pit mining by the owner with leased supply of mobile mining equipment including scheduled additions and replacements. All other project costs are the responsibility of the owner including process and infrastructure preproduction capital, life of mine sustaining capital, and closure costs. The estimate covers the direct field costs of executing the project, plus the Owner’s indirect costs associated with design, construction, and commissioning. All costs were expressed in third quarter (Q3) 2010 U.S. dollars. No allowance has been included for escalation, interest or financing fees, taxes or duties, or working capital during construction. The level of accuracy for the estimate is ±20 percent of estimated final costs, as per AACE Class 4 definition. Capital costs are summarized in Table 1-2, and total $1.599 billion.

Table 1-2: Summary of Capital Costs Area 2010 ($000) Pre-stripping 70,008 Mining Infrastructure & Equipment 101,942 Processing Plant & Acid Plant 381,277 On Site Infrastructures 133,444 Tailings Management 101,527 Project Access Road 38,795 Water Diversion and Reservoir 50,753 Port Site Facilities 26,887 Total Direct 904,633 Owner’s Cost 52,899 Indirects 243,076 Total Indirects 295,975 Contingency 236,522 Total Capital Cost 1,437,160 Escalation (excluded) — Working Capital (included in Financial Analysis) — Total 1,437,160

1.22 Projected Operating Costs

The operating costs for the Cañariaco Norte project are based on an Owner-operated mining fleet and process facility and have been prepared in third quarter 2010 U.S. dollars. The operating cost estimates were assembled by area and component, based on estimated staffing levels, consumables, and expenditures, according to the mine




plan and process design. Average operating costs over the LOM are shown in Table 1-3.

Table 1-3: Average Operating Costs

Area Unit US$ Unit US$/lb Cu

Mining $/t processed 2.74 $/lb Cu 0.360 Processing $/t processed 3.11 $/lb Cu 0.408 General & Administration $/t processed 0.43 $/lb Cu 0.056 Total Operating Costs $/t processed 6.28 $/lb Cu 0.824

1.23 Financial Analysis

The results of the economic analyses discussed in this section represent forward-looking information as defined under Canadian Provincial securities law. The results depend on inputs that are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here. The economic analysis is based on the evaluation of Measured and Indicated Mineral Resources. Mineral resources are not mineral reserves, and do not have demonstrated economic viability.

The Project has been valued using a discounted cash flow (DCF) approach. Net annual cash flows are discounted back to the date of valuation in end-of-year 2010 dollars and totalled to determine NPVs at the selected discount rates. The payback period is calculated as the time needed to recover the initial capital spent. All monetary amounts are presented in US dollars.

The pre-tax cumulative cash flow for the base case is $5,998 million with an IRR of 26.8 percent. The cash flow analysis shows that the Project will generate a positive cash flow in all years except Year -2 and Year -1 on a pre-tax basis. The annual positive cash flow results in a payback period of 2.9 years. At an eight percent discount rate, the pre-tax NPV of the project is $1,983 million.

The after-tax cumulative cash flow for the base case is $3,855 million with an IRR of 18.8 percent. The cash flow analysis shows that the project will generate a positive cash flow in all years except Year -2 and Year -1 on an after-tax basis. The annual positive cash flow results in a payback period of 4.1 years. At an eight percent discount rate, the NPV of the project is $1,063 million.

The financial analysis results for the base case are summarized in Table 1-4.




Table 1-4: Results of Financial Analysis for Base Case (base case is highlighted)

Summary of Cash Flow Pre-Tax After Tax

Cumulative net cash flow Undiscounted US$000 5,997,662 3,855,417 Net present value Discounted at 4% US$000 3,395,140 2,040,471 Discounted at 6% US$000 2,588,605 1,481,321 Discounted at 8% US$000 1,983,089 1,063,416 Discounted at 10% US$000 1,520,971 746,126 Discounted at 15% US$000 763,306 231,322 Internal rate of return US$000 26.8% 18.8% Payback period Years 2.9 4.1

Note: Base case copper prices are based on long-term reverting price curves whereby metal prices in the early years of operation are higher and gradually decrease to fixed long term prices after Year 10. This approach recognizes the industry-consensus view that future copper and precious metal prices will remain higher than historical price trends over the short- to medium-term, with reversion to lower long-term prices.

Sensitivity analysis was performed on the base case taking into account variations in the metal price, operating cost, foreign exchange, and mining cost. Analysis shows that the Cañariaco Project is most sensitive to changes in metal prices, as this directly affects the revenue stream.

The Project is less sensitive to capital expenditure and operating cost. The sensitivities of the two costs are very similar. Considering that the Project is priced in US dollars, the effects of exchange rate variation do not apply in the current model, although in reality some equipment, supplies, and services will be priced in Euros or the local Peruvian currency.

1.24 Alternative Financial Cases

Three alternative cases were evaluated. These used the parameters of the base case with the exception of the capital costs. The other three scenarios included:

• Full Owner-operated mine and process facilities

• Contractor mining

• Full lease of process and electrical equipment in addition to mining equipment.

The capital cost for the full Owner-operated and financed option is $1.599 billion for start-up and $206 million to sustain the operation. If the contract mining option were considered, then the cost could be reduced to $1.405 billion and $63 million for start-up and sustaining capital, respectively. The option involving full lease of mine and




major process and electrical equipment reduces the capital cost to $1.191 billion and $207 million for sustaining capital.

The financial evaluation results of the alternative cases are included as Table 1-5.

Table 1-5: Results of Financial Analysis for Alternative Cases

Item Unit Value at $2.25/lb Cu Value at $3/lb Cu Full Owner Operated – pre-tax

IRR % 25.8 35.8

CNCF* C$000,000 6,017 9,842

NPV 8% C$000,000 1,968 3,504

NPV 10% C$000,000 1,501 2,776

Payback Years 3.0 2.3 Contract Mining – pre-tax

IRR % 27.0 37.8

CNCF* C$000,000 5,754 9,579

NPV 8% C$000,000 1,919 3,456

NPV 10% C$000,000 1,476 2,751

Payback Years 2.8 2.1 Equipment Leasing – pre-tax

IRR % 29.9 42.0

CNCF* C$000,000 5,847 9,672

NPV 8% C$000,000 1,989 3,525

NPV 10% C$000,000 1,547 2,822

Payback Years 2.6 1.9 *Cumulative Net Cash Flow

1.25 Exploration Potential

The Cañariaco Norte deposit is open to depth and the potential exists for the mine life to be extended beyond the 22 years proposed in this report and/or the mining and processing rate to be increased. Both opportunities offer the potential to enhance the economics of the Cañariaco Norte Project.

The potential exists to discover and delineate additional resources at the Cañariaco Sur and Quebrada Verde targets which are located approximately 1.5 kilometres and 3.5 kilometres south of the Cañariaco Norte deposit. These targets are located within CCC's Cañariaco property and, should an economic deposit be delineated at either site, it is possible that development would utilize the proposed Cañariaco Norte facilities.




1.26 Conclusions

In the opinion of the QPs, the Project that is outlined in this Report has met its objectives in that mineralization has been identified that can support estimation of Mineral Resources. There is sufficient additional scientific and technical information to support an updated preliminary assessment (PA) study.

Work performed and reported in this Report was based on a pre-feasibility study commissioned by CCC, and is based on interim information from that study. Additional geotechnical drilling and rock quality assessment is required to complete the open pit slope design to a level consistent with generally accepted pre-feasibility requirements. The above ground structures in the tailings management facility were designed to a pre-feasibility level; however, geotechnical investigations in the TMF area have not been completed and as a result there is a risk that the TMF design may require revision.

1.27 Recommendations

The following work is recommended; no phase is contingent on another, and all can be conducted concurrently. The total cost of the recommendations ranges from US$2.21 million to US$2.41 million.

CCC should:

• Initiate insertion of fine pulp duplicate samples for precision analysis, as this is currently not part of the CCC sampling program. Such checks should be instituted in all future exploration programs. AMEC estimates that this would add approximately 5% to the total costs of the sample analytical programs

• Construct a relational database using a commercial database software package, for more detailed studies. Costs of such a program would depend on whether the database was constructed in-house or a third-party consultant was contracted to complete the exercise. Estimated cost: US$10,000

• Drill 11 geotechnical core holes (6,690 m) to provide additional information in the proposed pit area on rock mass properties, joint patterns, hydrogeology and geological modeling/resource estimation. Estimated cost: US$1.1 M

• Drill additional holes to increase the level of confidence of the lithological interpretation in areas outside the existing drill hole limits but constrained by the resource pit. The proposed geotechnical holes will cover part of the recommendation; however an additional 10 holes (6,000 m) is suggested. Estimated cost: US$900,000 to US$950,000




• Complete a full review of the alteration interpretation on vertical sections reconciled to bench plans. Construct a mineral zonation model for the next phase of study. Evaluate arsenic distribution using the interpreted units. Estimated cost: US$200,000 to US$350,000

• Institute a continuous program of specific gravity determinations from core samples preferably using the same laboratory and determination procedures as those used in the 2010 testing program. AMEC estimates that this would add approximately 2–5% to the total analytical program costs.

At the completion of this work, results should be reviewed and incorporated as appropriate into the ongoing pre-feasibility study or future detailed studies.




2.0 INTRODUCTION

AMEC Americas Limited (AMEC) was commissioned by Candente Copper Corporation (CCC) to prepare an independent Qualified Person’s Review and NI 43-101 Technical Report (the Report) on the pre-feasibility study progress report which includes a preliminary assessment (PA) update for the wholly-owned Cañariaco Norte Copper Project (the Project) located in the Lambayeque Department, Peru. The Project contains the Cañariaco Norte copper deposit, and the Cañariaco Sur and Quebrada Verde prospects. The Project location is shown in Figure 2-1.

CCC will be using the Report in support of disclosures in the CCC press release dated 7 March 2011, entitled “Candente Copper Files NI 43-101 Technical Report and Updates Key Financial Results for the Cañariaco Norte Copper Project Pre-Feasibility Study Progress Report”, and the CCC press release dated 18 January 2011, entitled “Candente Copper Announces Positive Pre-Feasibility Progress Report for the Cañariaco Norte Copper Project”.

All measurement units used in this Report are metric, and currency is expressed in US dollars (US$) unless stated otherwise. The Report uses Canadian English.

As at the effective date of the Report of 18 January 2011, the exchange rate was US$1 equal to approximately 2.78 Peruvian Nuevos Soles.

2.1 Terms of Reference

AMEC was requested to complete a Pre-Feasibility Study for CCC on the Cañariaco Norte Project. The study that is the subject of this Report is considered to be a Pre-feasibility Study Progress Report, because:

• Additional geotechnical drilling and rock quality assessment is required to complete the open pit slope design to a level consistent with pre-feasibility requirements; this drilling has been planned

• Geotechnical investigations are required in the area where above-ground structures are proposed as part of tailings management facility (TMF) design. Information from the planned geotechnical program may result in modifications to the design.

The mine plan is based on Measured and Indicated mineral resources only, with Inferred mineral resources considered as waste. The economic analyses based on these mineral resources, although modelled to a level consistent with a pre-feasibility study, are considered to be a PA under NI 43-101.




Figure 2-1: Project Location Map

Note: Figure courtesy Candente Copper Corp.

Table 2-1: QPs for the Report

Qualified Person Date of Site Visit Report Sections of Responsibility (or Shared Responsibility)

David Thomas 7 to 11 June 2010 Sections 4, 5, 6, 7, 8, 9, 10, 11, 12, 13, 14, 15, and 17, and those portions of the Summary, Conclusions and Recommendations that pertain to those sections

Jay Melnyk No site visit Sections 19.1 to 19.4 and those portions of the Summary, Conclusions and Recommendations that pertain to those sections

Tony Lipiec No site visit Section 16, and those portions of the Summary, Conclusions and Recommendations that pertain to that section

Alexandra Kozak No site visit Sections 1, 2, 3, 18, 19.5 to 19.10, 20, 21, 22, and 23




2.2 Qualified Persons

The following are the Qualified Persons (QPs) as defined in National Instrument 43-101, Standards of Disclosure for Mineral Projects, for portions of the Report:

• David Thomas, P.Geo., MAusIMM, Senior Geologist, AMEC Vancouver • Jay Melnyk, P.Eng., Principal Engineer, AMEC Vancouver • Tony Lipiec, P.Eng., Principal Process Engineer, AMEC Vancouver • Alexandra Kozak, P.Eng., Manager of Process Engineering, AMEC Vancouver.

Table 2-1 summarizes the site visit dates and the sections of responsibility for Report preparation. During the site visit, Mr Thomas visited the Project area, where he conducted an inspection of core and surface outcrops, viewed drill platforms and sample cutting and logging areas; discussed geology and mineralization interpretations with CCC’s staff and reviewed geological interpretations in sections and plans.

Gustavo Gonzaga, an AMEC senior geotechnical engineer inspected drill core stored at the Project core shack adjacent the Project site; Vincent Maddalozzo, a senior civil engineer with AMEC, visited the Project. Both AMEC staff members provided input to the QP authors for the Report on information from the site.

Information that supports the Report has been obtained from CCC, or external consultants, or has been prepared by the QPs. Reference documents are cited in the text as appropriate and summarized in Section 22 of this Report.

The QPs are not aware of any material scientific and technical changes to the information on the Project between the dates of the site visit and Report signature date.

2.3 Effective Dates

The Report has a number of effective dates, including:

• Effective date of the Mineral Resources: 8 November 2010

• Effective date of the progress report on the pre-feasibility study: 18 January 2011.

The effective date of the Report is taken to be the date of the interim report on the pre-feasibility study and is 18 January 2011.




2.4 Previous Technical Reports

CCC has previously filed the following technical reports on the Project:

• Thomas, D., and Lipiec, T., 2010: Candente Copper Corporation, Cañariaco Project, Lambayeque Department, Peru NI 43-101 Technical Report: technical report prepared by AMEC Americas Ltd for Candente Resource Corp., effective date 8 November 2010

• Bonson, C., Campbell, R., Bender, M., Doerksen, G., Johnston, A., Meyer, T., Nowak, M., Pilotto, D., Van Egmond, R., Critikos, P., Ostolaza, R., and Huanani, A., 2008: Revised Preliminary Economic Assessment Technical Report, Cañariaco Norte Project, Peru: technical report prepared by SRK Consulting Ltd for Candente Resource Corp., effective date 30 November 2008

• Bonson, C., Nowak, M., Doerksen, G., Johnston, A., and van Egmond, R., 2008: Technical Report Cañariaco Norte Project, Department of Lambayeque, Peru: technical report prepared by SRK Consulting Ltd for Candente Resource Corp., effective date 11 July 2008

• Stone, D.M.R., Godden, S., Van Egmond, R., and Tosney, J.R., 2007: Updated Technical Report and Preliminary Economic Assessment on the Cañariaco Norté Porphyry Copper Project Starter Pit Option: technical report prepared by Minefill Services Inc. for Candente Resource Corp., effective date 18 April 2007

• Collins, J., McCrea, J., and Rokosh, J., 2006: Cañariaco Copper Project, Peru, Preliminary Assessment and Economic Evaluation Report: technical report prepared by Merit Consultants for Candente Resource Corp., effective date 28 June 2006

• McCrea, J.A., 2006: Technical Report on the Cañariaco Copper Porphyry Prospect Department of Lambayeque Northwest Peru: technical report prepared for Candente Resource Corp., effective date 27 April 2006

• McCrea, J.A., 2005: Technical Report On The Cañariaco Copper Porphyry Prospect Department of Lambayeque Northwest Peru: technical report prepared for Candente Resource Corp., effective date 22 April 2005

• Currie, J.A., 2004: Technical Report On The Cañariaco Copper Porphyry Prospect Department Of Lambayeque Northern Perú Latitude 6° 05’ S Longitude 70° 17’ E, technical report prepared for Candente Resource Corp., August 2004

• Huanqui, F., Freeze, J.C., and Coder, J.M., 2002: Geological Report On The Cañariaco Copper-Gold Porphyry Prospect Department Of Lambayeque Northwest Perú NTS 13-E 70° 17’ E 6° 05’ S: technical report prepared for Candente Resource Corp., effective date 9 August 2002.




2.5 References

Reports and documents listed in the Reference sections of this Report were used to support preparation of the Report.

2.6 Technical Report Sections and Required Items under Form 43-101F1

Table 2-2 relates the sections as shown in the contents page of this Report to the prescribed Items Contents Page of Form 43-101F1. The only difference is that Item 25 “Additional Requirements for Technical Reports on Development Properties and Production Properties” is located following Item 19, “Mineral Resource and Mineral Reserve Estimates”.




Table 2-2: Contents Page Headings in Relation to Form 43-101F1 Prescribed Item Headings NI 43-101 Item Number Form 43-101F1 Heading

Report SectionNumber Report Section Heading

Item 1 Title Page Cover page of Report Item 2 Table of Contents Table of contents Item 3 Summary Section 1 Summary Item 4 Introduction Section 2 Introduction Item 5 Reliance on Other Experts Section 3 Reliance on Other Experts Item 6 Property Description and Location Section 4 Property Description and Location Item 7 Accessibility, Climate, Local Resources,

Infrastructure and Physiography Section 5 Accessibility, Climate, Local Resources,

Infrastructure and Physiography Item 8 History Section 6 History Item 9 Geological Setting Section 7 Geological Setting Item 10 Deposit Types Section 8 Deposit Types Item 11 Mineralization Section 9 Mineralization Item 12 Exploration Section 10 Exploration Item 13 Drilling Section 11 Drilling Item 14 Sampling Method and Approach Section 12 Sampling Method and Approach Item 15 Sample Preparation, Analyses and Security Section 13 Sample Preparation, Analyses and

Security Item 16 Data Verification Section 14 Data Verification Item 17 Adjacent Properties Section 15 Adjacent Properties Item 18 Mineral Processing and Metallurgical Testing Section 16 Mineral Processing and Metallurgical

Testing Item 19 Mineral Resource and Mineral Reserve

Estimates Section 17 Mineral Resource and Mineral Reserve

Estimates Item 20 Other Relevant Data and Information Section 19 Other Relevant Data and Information Item 21 Interpretation and Conclusions Section 20 Interpretation and Conclusions Item 22 Recommendations Section 21 Recommendations Item 23 References Section 22 References Item 24 Date and Signature Page Section 23 Date and Signature Page Item 25 Additional Requirements for Technical

Reports on Development Properties and Production Properties

Section 18 Additional Requirements for Technical Reports on Development Properties and Production Properties

Item 26 Illustrations Incorporated in Report under appropriate section number




3.0 RELIANCE ON OTHER EXPERTS

The QPs, as authors of this Report state that they are qualified persons for those areas as identified in the relevant “Certificate of Qualified Person” attached to this Report. The QPs have relied on, and believe there is a reasonable basis for this reliance, upon the following reports of other experts, which provided information regarding mineral rights, surface rights, and environmental status in sections of this Report as noted below.

3.1 Mineral Tenure

The QPs have not reviewed the mineral tenure, nor independently verified the legal status, ownership of the Project area, underlying property agreements or permits. AMEC has fully relied upon, and disclaims responsibility for, information derived from legal experts for this information through the following documents:

• Egusquiza, J., 2010: Legal Aspects Regarding the Status of Mining Concessions, Surface Rights of Cañariaco Copper Perú S.A. and a Brief Report of the Tax Regimen and Royalties: unpublished legal report prepared by Cañariaco Copper Perú S.A. legal counsel for AMEC Americas Ltd., 30 November, 2010.

This information is used in Section 4.2 of the Report.

• Muraro, T., 2010: Tenure: unpublished email from Cañariaco Copper Perú S.A. to AMEC Americas Ltd, 6 December, 2010.

This information is used in Section 4.2.1 of the Report.

3.2 Surface Rights

The QPs have fully relied upon information supplied by CCC’s staff and experts retained by CCC for information relating to the status of the current Surface Rights as follows:

• Egusquiza, J., 2010: “Legal Aspects Regarding the Status of Mining Concessions, Surface Rights of Cañariaco Copper Perú S.A. and a Brief Report of the Tax Regimen and Royalties: unpublished legal report prepared by Cañariaco Copper Perú S.A. legal counsel for AMEC Americas Ltd., 30 November, 2010.





3.3 Royalties

The AMEC QPs have fully relied upon information supplied by CCC staff and experts retained by CCC for information relating to the status of the current royalty provisions for the Project as follows:

• Egusquiza, J., 2010: “Legal Aspects Regarding the Status of Mining Concessions, Surface Rights of Cañariaco Copper Perú S.A. and a Brief Report of the Tax Regimen and Royalties: unpublished legal report prepared by Cañariaco Copper Perú S.A. legal counsel for AMEC Americas Ltd., 30 November, 2010





4.0 PROPERTY DESCRIPTION AND LOCATION

The Cañariaco Project is situated within the District of the Province of Ferreñafe, Department of Lambayeque, in northwestern Peru, and is approximately 700 km northwest of Lima, the capital of Peru, and approximately 110 km to the northeast of the city of Chiclayo. Project centroids are latitude 06° 05’ south and longitude 79° 17’ west.

4.1 Property and Title in Peru

Information in this sub-section has been compiled from the Mining Guide to Peru (Ministry of Energy and Mines – General Mining Bureau, 2006). The General Mining Law of Peru defines and regulates different categories of mining activities, prospecting, exploration, exploitation, and processing (D.S. No. 014-92-EM, 19921).

4.1.1 Mineral Tenure

The holder of a mining concession is entitled to all the protection available to holders of private property rights under the Peruvian Constitution, the Civil Code, and other applicable laws. A Peruvian mining concession is a property-related right; distinct and independent from the ownership of land on which it is located, even when both belong to the same person. The rights granted by a mining concession are defensible against third parties, are transferable and chargeable, and, in general, may be the subject of any transaction or contract.

To be enforceable, any and all transactions and contracts pertaining to a mining concession must be entered into a public deed and registered with the Public Mining Registry (Registro Público de Minería). Conversely, the holder of a mining concession must develop and operate his/her concession in a progressive manner, in compliance with applicable safety and environmental regulations, and with all necessary steps to avoid third-party damages. The concession holder must permit access to those mining authorities responsible for assessing that the concession holder is meeting all obligations.

Ownership of mineral claims is controlled by mining concessions that are established using Universal Transverse Mercator (UTM) coordinates to define areas of interest that are measured in hectares. Individual concessions must have a minimum size of 100 ha and a maximum size of 1,000 ha. New concessions must be orientated north-south; pre-existing concessions are based on the ‘punto de partido’ system and can be of any orientation. Mining titles are irrevocable and perpetual, as long as the




required annual maintenance fees (derecho vigencia) are up to date and fully paid to the Ministry of Energy and Mines (MEM) by 30 June each year following granting of a concession. The fees are paid in advance. The annual fee for metallic mineral concessions is $3/ha for each concession that is either actually acquired or pending (petitorio). The concession holder must reach a minimum level of commercial production of $100/ha in mineral sales within eight years of the concession being granted. If there is no production income after eight years, then the annual fee increases to $4/ha until the 14th year and to $10/ha thereafter.

Exploitation

The concession holder must sustain a minimum level of annual commercial production greater than $100/ha in gross sales before the end of the sixth year following the granting of the concession. If the concession has been put into production within the six-year period, then the annual rental fee remains the same up to the beginning of Year 9, after which it increases to $4/ha for Years 9 to 14. The annual rental rises to $10/ha for each year thereafter. If the concession has not been put into production within a six-year period, then the annual rental increases from the first semester of the seventh year to $9 ($3 for derecho vigencia, plus a $6 penalty) until the minimum production level is met.

If, by the start of the twelfth year from granting a concession, the minimum production level has not been achieved, then the annual rental increases to $23/ha ($3 for derecho vigencia, plus a $20 penalty). A concession holder can, however, be exonerated from paying penalties if he/she can demonstrate that at least ten times the penalty for the total concession was invested during the previous year. The investment must be documented and it must be accompanied by a copy of the relevant annual tax statement (declaración jurada de impuesto a la renta) and payment of the annual fees. A concession will terminate if the annual rental is not paid either for three years in total over the period that the concession is held, or for two consecutive years over the period the concession is held.

4.1.2 Surface Rights

It is a requirement of the Peruvian Government that any property developer either purchases the surface rights, or makes an appropriate agreement with the surface rights owner, for access to a property. In the case of the holders of mining concessions, they are protected under the Peruvian Constitution and Civil Code. Their concession rights do not, however, confer ownership of the land. Thus, the owner of a mining concession must deal with the registered land owner to obtain the

1 See http://www.minera.gob.pe/mineria/legislacion/data/D.S.N_014-92-LSM.doc




right of access to fulfill the production obligations inherent in the concession grant. All transactions and contracts pertaining to a mining concession must be registered with the Public Mining Registry in the event of subsequent disputes in law. The issue of land tenure continues to be of significance in Peru, not least because the national cadastral system for agricultural land ownership is not always accurate.

4.1.3 Royalties

Perú established a sliding scale for mining royalties in 2005. Under this regulation, all mining companies will be levied a royalty from 2018 once a mine is in production. Calculation of the royalty payable will be on a monthly basis, and will be based on the value of the concentrate sold (or its equivalent) using international metal prices as the base for establishing the value of contained metal. The sliding scale is applied to the value of ore concentrate or equivalent, as follows:

• First stage: 1.0% up to US$60 million in annual concentrate sales

• Second stage: 2.0% in excess of US$60 million and up to US$120 million in annual concentrate sales

• Third stage: 3.0% in excess of US$120 million in annual concentrate sales.

Legislation is currently being considered by the Peruvian Government, which may result in the following increases in the royalty rates:

• First stage: 2.0% up to US$60 million in annual concentrate sales

• Second stage: 4.0% in excess of US$60 million and up to US$120 million in annual concentrate sales

• Third stage: 6.0% in excess of US$120 million in annual concentrate sales

• Gold and silver: royalties of 10% and 5% respectively.

4.1.4 Environmental Regulations

The General Mining Law of Peru is the primary body of law with regard to environmental regulation of exploration and mining activities. The General Mining Law is administered by the Ministry of Energy and Mines (MEM). A detailed description of Peru’s environmental regulations is found on the MEM website2. Generally, the MEM requires exploration and mining companies to prepare an Environmental Impact Statement (DIA)– Category I, Environmental Impact Study

2 See http://www.minem.gob.pe




Semi Detailed (EIAsd)–Category II (Table 4-1), an Environmental Impact Assessment (EIA), a Program for Environmental Management and Adjustment (PAMA), and a mine closure plan. Mining companies are also subject to annual environmental audits of operations by the Organismo Supervisor de la Inversion en Energía y Minería (Osinergmin).

Under Peruvian regulations (D.S. 020-2008-EM y la R.M. 167-2008-MEM-DM) a DIA-Category I covers drilling of less than 20 drill holes within a 10 ha area. An EIAsd–Category II is applicable to mining and exploration programs with either more than 20 drill holes, exploration areas greater than 10 ha, or construction of more than 50 m of tunnels. Both classifications require development of public community involvement processes, which are administered under regulations D.S. 028-2008-EM and R.M. 304-2008-MEM-DM.

The MEM typically gives automatic approval of DIA–Category I studies, and turnaround is of the order of 10 days. An EIAsd–Category II study typically can take several months for approval, due to notification periods and public community participation processes. Drill approvals for both categories are predicated on the drill hole collar locations being within a 50 m radius of the sites proposed to MEM; failure to comply with this can lead to Osinergmin imposing a range of penalties, from fines to closure of operations.

A mining company that has completed its exploration stage work program must submit an EIA or a modified EIA either when applying for a new mining or processing concession, increasing the size of existing processing operations by more than 50%; or executing any other changes to an existing mining project that results in a >50% change in the mining rate or expected profit (DS 016-93-EM. Cap III, Art° 20), as follows:

• A new EIA must be developed when additional, previously un-mined areas are proposed to be added to an operation (DS 016-93-EM, D.S. 028-2008-EM and R.M. 304 -2008-MEM-DM, review articles 15 and 16), and must include preparation of an executive summary and scheduling of workshops and public community participation




Table 4-1: Summary of Environmental Requirements for Mining Exploration Programs

Classification Description Application Requirements Application

Fees Approval

Time

Category I – Environmental Impact Statement - DIA

Mineral exploration with less than 20 drill holes within a 10 ha area

Required information as shown in Art. 5 of Environmental Regulations for Mining Exploration

5% of Unit Tax Unit is ~US$1,000 Tax = US $50

10 days

Category II – Environmental Impact Study Semi Detailed -EIAsd

Mineral exploration with more than 20 drill holes, exploration areas greater than 10 ha, and/or construction of more than 50 m of tunnels

Prepare an Environmental Evaluation (EA) report as per Appendix 2 of Environmental Regulations for Mining Exploration

40% of Unit Tax Unit is ~US$1,000 Tax = US $400

30 days

• The EIA must incorporate planned expenditure on environmental programs at a rate that is no less than 1% of the value of annual production of the planned operation. The MEM must review and make a decision on the project within 120 days, including initial notification, and the initial stage of the public consultation process. The process of actual project approval may take 8–12 months. Within this period the applicant company must organize hearings and workshops to present project data and coordinate the dates and locations of such hearings with the MEM.

If the proposed project results in a <50% change in the mining rate or expected profit, the changes to the existing EIA may be accepted, subject to informational workshops and public hearings being held. The MEM must review and make a decision on the existing project EIA within 30 days, including initial notification, and the initial stage of the public consultation process. The approvals process may take 6–8 months.

A mining company must also prepare and submit a closure plan (Plan de Cierre) for each component of its operation. The closure plan must outline what measures will be taken to protect the environment over the short-, medium- and long-term, from solids, liquids and gases generated by the mining operation.

The General Mining Law of Peru has in place a system of sanctions or financial penalties that can be levied against a mining company that is not in compliance with the environmental regulations.




4.2 Mineral Tenure

4.2.1 Tenure History

The government of Peru auctioned the right to explore the Project under “Public Auction (Bid) PRI-51-2000, Private Investment Promotion – Mining Prospects”. The auction was held in Lima on 14 February 2001 and Candente through its wholly-owned Peruvian subsidiary Exploraciones Milenio S.A. (EMSA), was the sole bidder. Consideration was a one-time payment of $75,880.

On grant, the Project consisted of 21 “Cañariaco” claims totaling 16,560 ha. In addition, four claims, the “Jehumarca” claims, which are contiguous with the Cañariaco claims, were acquired in 2008, and have an area of 3,600 ha.

In 2009, CCC elected to discontinue the vigencia payments for the Cañariaco K, L, M, P, R and S concessions, totalling 4,170 ha. As payments have to be suspended for two consecutive years before they are forfeit, the Cañariaco K, L, M, P, R and S concessions did not expire until June 2010.

During 2009, CCC acquired the 400 ha Cañariaco W concession.

4.2.2 Current Tenure

The Project currently consists of 16 Cañariaco concessions, covering a total area of 13,400 ha and four Jehumarca claims, of 3,600 ha. Concessions are listed in Table 4-2 and shown in Figure 4-1. Tenure is held in the name of the wholly-owned CCC subsidiary, Cañariaco Copper Perú S.A.

Mining legislation in Peru does not require location of concession boundaries on the ground. The boundary limits are defined by UTM coordinates in the mineral concession title filed on record in the registrar of the Ministerio de Energía y Minas (Ingemmet) y Sistema Único Nacional de Registros Publicos (Sunarp).

To maintain the property in good standing, annual property payments of $17,500 commenced in June 2002. AMEC was advised by CCC that all required property payments have been made, and the concessions are in good standing. Providing the annual property payments are made in a timely manner, the concessions will not expire.




Table 4-2: Cañariaco Concessions held by Candente Copper Corp. (as at 26 November 2010)

Nº Name Code Hectare

Public Registry

Entry

Non-Production Penalty Payment

(Debt) 2010 (US$)

Validity Fee Paid Until **

1 Cañariaco A 150000004Y02 1,000 20002633 6,000.00 2010 2 Cañariaco B 15000005Y02 1,000 20002634 6,000.00 2010 3 Cañariaco C 15000006Y01 1,000 20002635 6,000.00 2010 4 Cañariaco D 15000007Y01 500 20002636 3,000.00 2010 5 Cañariaco E 01-00226-04 1,000 (662.5) # 11054442 2010 6 Cañariaco F1 01-00312-04 1,000 (789.5) # 11054445 2010 7 Cañariaco G 01-01790-05 1,000 (940) # 11054460 2010 8 Cañariaco H 01-03174-06 1,000 11071154 2010 9 Cañariaco I 01-03175-06 1,000 11071158 2010 10 Cañariaco J 01-03212-06 1,000 11071162 2010 11 Cañariaco N 01-03300-06 1,000 11071194 2010 12 Cañariaco Q 01-03309-06 400 11071198 2010 13 Cañariaco T 01-03312-06 800 11071208 2010 14 Cañariaco U 01-03313-06 400 11071210 2010 15 Cañariaco V 01-03314-06 900 11071218 2010 16 Cañariaco W ## 01-01827-09 400 2010

Nº Name Code Hectare Registration

Date Concession Granted Validity Fee Paid

Until ** 17 Jehuamarca 1 01-01551-05 900.00 01/06/05 07/11/05 2010 18 Jehuamarca 2 01-01552-05 900.00 01/06/05 09/11/05 2010 19 Jehuamarca 3 01-01553-05 900.00 01/06/05 07/11/05 2010 20 Jehuamarca 4 01-01554-05 900.00 01/06/05 11/16/05 2010

Notes: * By Supreme Decree Nº 054-2008-EM published in the Official Gazette “El Peruano” on October 10th 2008, a minimum

annual production requirement was established, equivalent to US$100.00 per year and hectare for metallic minerals or equivalent to US$ 50.00 per year and hectare for non-metallic minerals. This production requirement is applicable from the sixth year of grant of the mining concession. In the event of non-fulfilment of this minimum production obligation, the owner of a mining concession must pay, from the seventh year onward, a penalty of US$6.00 per year and hectare (in addition to the statutory validity fee payment) up to the year when the minimum annual production occurs. If the owner of mining concession continues without paying the penalty until the twelfth year, the penalty payment rises to US$ 20.00 per year and hectare (in addition to the statutory validity fee payment).

** The validity fee is currently US$3 per year and per hectare. Non-payment of the fee for two consecutive years results in cancellation of the mining concession

# The hectareage granted for Cañariaco F1, G and Cañariaco H are 1,000 hectares, but for purposes of establishing the validity fee, Ingemmet uses the hectarages shown in brackets

## Registration of the Cañariaco W title is in process


Lambayeque Department, PeruNI 43-101 Technical Report on Prefeasibility Study Progress Report


Figure 4-1: Cañariaco Concessions held by Candente Copper Corporation

Note: Grid squares on plan indicate scale and are 2 km x 2 km.




4.3 Surface Rights

The owner of the surface rights is the Comunidad Campesina San Juan Bautista de Cañaris. Under Peruvian Law 24656, CCC is required to have a valid agreement with the campesina to undertake exploration; typically this has been negotiated on an as-needs basis. The previous agreement expired in December 2009, and Candente is in the process of negotiating a new, three-year, access agreement.

4.4 Royalties

There are no royalties payable to any third parties on any of the concessions that currently comprise the Project.

The Project is likely to incur a royalty of 3%, payable to the Government of Peru, once production commences. This figure is based on the current royalty scheme, as the Project is likely to produce in excess of US$120 million in annual concentrate sales. The royalty burden may increase if the proposed royalties increase before the Peruvian Government are passed into legislation.

4.5 Permits

Exploration activities to date have been undertaken in accordance with the appropriate Peruvian regulations. A Class B environmental assessment (EA) was conducted by GEMA Consultants (GEMA) for CCC, as required by Peruvian law when trenching, drilling, or other major work is performed. GEMA also conducted a Class C EA and obtained a Class C permit from the Peruvian government for advanced drilling and exploration work at Cañariaco. Additional permits will be needed to support the planned geotechnical and infill drilling programs discussed in Section 21.

A number of permits will be required to support project development and operation. The main permit requirements identified to date for the start of construction and mining operations are summarized below.

• Environmental Impact Assessment (EIA), Studies and Permits • Water Authorization • Archaeological Evaluation • Mining Plan and Closure Plan • Mining Operation Certificate (COM) • Easement by Agreement • Procedure of Real Servitude • Construction Permit • Municipality Permits to Construct




• Labour Permits • Surface Water Use Licence • Ground Water Use Licence • Sanitary Authorization for Waste Water Treatment • Permits to Build Roads • Transportation deed for controlled substances and products • Beneficiation Concession (necessary to process the mineral) • Authorization to Start Operation.

CCC has committed to staying apprised of changes in Peruvian environmental and mining regulations, which are currently being updated. The legislation will need to be reviewed continuously to identify all the social and environmental studies, permits, and authorizations required, as well as the permitting process and the timeline itself so that it can match the project schedule with the licensing schedule.

4.6 Environmental

The site of the Cañariaco Norte Project is in Northern Peru, in an area environmentally similar to the surrounding region that extends into Ecuador and Columbia.

AMEC Peru, on behalf of CCC, has been carrying out environmental baseline studies and monitoring since 2007. Findings to date include:

• The Project site is subject to distinct wet and dry seasons

• Hydrological investigations show that naturally acidic waters are present in mineralized areas within the project footprint

• At present there are no Peruvian background guidelines for sediment levels in watercourses in this area. Upon comparison to Canadian guidelines, water samples from the area showed some dissolved metal levels above probable effect levels (PEL), with copper and zinc nearly 50% above Interim Guidelines (ISQGs). This is not unusual for naturally occurring mineralized areas where PELs are routinely exceeded

• Acidic to slightly acidic pH levels have been identified only at the confluence of the Oso and Norte watersheds

• Due to steep slopes and the high clay and fines contents, soils within the region are typically unstable

• AMEC assessed the baseline air quality in the project area on five separate occasions. The results indicate that air quality is typical of rural, non-industrialized areas, with no contaminants at concentrations of concern to human health.




The Project ecosystem consists of moors and cloud forests at higher elevations and cultivated land at lower elevations. No native fish species were found within the rivers and creeks at the Project site, but introduced trout were identified downstream in the main channel of the Río Cañariaco. Native fish species were found within the lower basin of both Río Cañariaco and Quebrada Yerma and in Río Huancabamba, approximately 40 km from the Project site.

The project area has been studied more intently than the surrounding region with respect to biodiversity indicators such as species richness (for orchids in particular, with more than 150 species recorded), level of endemism (for amphibians, orchids, insects, and small mammals), and the presence of endangered species. All of the protected and endangered species recorded or sited within the greater project footprint are also found in similar habitats within Peru or in neighbouring countries.

AMEC notes that in many areas of Peru biodiversity assessments are not yet complete and therefore it is not uncommon for detailed environmental impact investigations to identify new species. CCC has instigated two conservation projects that encompass the relocation of protected orchids found at the site and the identification of an area for conservation and environmental management.

Once the baseline studies are completed to levels in accordance with the requirements of the Peruvian Government and meet the guidelines set out in the Equator Principles, then a process of obtaining all the required environmental, construction and mining permits will be undertaken.

At this stage of the Project, environmental liabilities are limited to those expected for an exploration-stage project, and include drill pads and access roads.

4.7 Socio-Economics

CCC has received a certificate of confirmation from the Peruvian National Institute of Culture (INC) that no archaeological remains exist within the area of the proposed Cañariaco Norte open pit.

Approximately 2,000 ha of the likely Project footprint are located within lands held by the small community of San Juan Bautista de Cañaris, which was settled during the Spanish Colonial period. The remaining portion falls within lands held by the José Carlos Mariátegui hamlet. In addition to the settlements, there are isolated families, grazing pastures and a portion of the Cañaris forest within the Project area. Approximately 450 residential units inhabit the region, living on small farms where they engage in a subsistence way of life.




A baseline study will be required prior to Project development to provide data on existing residents, assets, and services within the Project area to assess potential relocation requirements and implications.

4.8 Comment on Section 4

In the opinion of the QPs, the information discussed in this section supports the declaration of Mineral Resources, based on the following:

• Information from legal experts support that the mining tenure held is valid and is sufficient to support declaration of Mineral Resources

• Surface rights are held by the Comunidad Campesina San Juan Bautista de Cañaris, and access is negotiated with the community as required. CCC is currently negotiating a three-year exploration access agreement with the community. AMEC notes that access agreements will be a prerequisite for the recommended drill programs in Section 21

• There are no royalties payable to any third party. On production, royalties are payable to the Peruvian Government, at a fixed scale. The fixed royalty scale is under review by the Government of Peru, and may increase

• Permits obtained by the company to date to undertake exploration activities are sufficient to ensure that activities are conducted within the regulatory framework required by the Peruvian Government

• Additional permits will be required for Project development; preliminary discussions have been held with the relevant statutory authorities. A number of permits will be required to support Project development and operation

• Development of the Cañariaco Norte Project would require resettlement of a small number of residents who inhabit the immediate Project area; such relocation will require additional studies and negotiations

• At the effective date of this report, environmental liabilities are limited to those expected for an exploration-stage project, and include drill pads and access roads

• The current state of knowledge on environmental and permit status for the Project supports the declaration of Mineral Resources.




5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1 Access

The Cañariaco property is accessible by road from Lima via the city of Chiclayo (Figure 5-1). Either the Pan-American Highway (700 km, 11-hour trip) or a commercial airline flight can be taken from Lima to Chiclayo.

The route from Chiclayo to the Cañariaco property is currently a 149 km, 6-hour trip along mostly unimproved roads via Incahuasi. The road is not suitable for heavy trucks, and there is no landing strip at the Project site.

5.2 Climate

Temperatures in the project area vary with altitude and range between approximately 3°C and 20°C. The region receives between 830 mm and 1,700 mm of rainfall each year. The rainy season extends from November to March.

The climate does not affect exploration and other ground-based operations, although helicopter-supported operations and road accesses within the Project area are more difficult during the rainy season. It is expected that future mining operations will be able to be conducted year-round.

5.3 Local Resources and Infrastructure

5.3.1 Local Resources

Peru has a long history of mining, with the result that mining professionals and machine operators are generally available in most population centres. CCC would be able to source Peruvian personal for virtually all mine administration and operation requirements.

Supplies and other normal services are available at Chiclayo. Specialist exploration services, such as drilling and geophysical and geochemical analyses, are usually contracted out of Lima.




Figure 5-1: Cañariaco Property Access Routes and Infrastructure

Note: Figure courtesy Candente Copper Corp. Deposits shown on plan other than Cañariaco

are held by third-parties and are not part of the Project that is the subject of this Report.

5.3.2 Existing Infrastructure

The Cañariaco Project has a fully established camp and catering facility for 100 persons. Two weather stations are also on site. Water supply is currently sourced from the Río Cañariaco, and electricity is supplied by an on-site generator system.

5.3.3 Planned Infrastructure

Figure 5-2 shows the layout of the main Project facilities; Figure 5-3

Access

Site access to support mining operations will require construction of a new access road. The Project site will be accessed from Chiclayo via the Pan-American Highway North, past Motupe to Olmos, and then east along the Corral Quemado paved highway to a point approximately 60 km east of the town of Olmos.




Figure 5-2: Planned Layout, Major Project Facilities




Figure 5-3: Proposed Mine Site General Arrangement

At this point, the planned all-weather access road to Cañariaco mine will branch off, and follow the Yerma and Cañariaco valleys south to the location of the proposed mine site.

Access to the plant site will be via a new 42 km single-lane (7 m wide) gravel road branching off the existing Corral Quemado Road along the Huancabamba Valley. The alignment follows a new bridge crossing over Río Huancabamba and then passes through challenging terrain, rising at a constant grade of 8% until it reaches the ridge top between the Yerma and Cañariaco valleys near the tailings management facility (TMF). The road then widens to 11 m to accommodate the surface-run tailings and reclaim pipelines running to and from the plant site. Continuation of the road to the process plant includes a single high point 1,200 m away from the plant.

On-site roads will be provided for vehicle access to facilities and remote structures. General access roads will be two-way and 8 m wide, service roads will be one-way and 5 m wide with pullouts, and mine haul roads will be one-or two-way as required and wide enough to accommodate haul trucks.




5.3.4 Port

The port site infrastructure for the project will be at the Marine Terminal Muchik (MTM), a planned development by Lumina Copper Corporation (Lumina) southeast of the town of Eten and the existing Consorcio Terminales Eten liquid bulk terminal in Northern Peru. Copper concentrate from the mine will be transported by truck to MTM at a rate of approximately 400,000 t/a. The loading berth at the port will be designed to accommodate up to 50,000 DWT loaded Handymax vessels.

The pre-feasibility study progress report assumes that the berthing facility, the trestle conveyor leading out to the berth, and the shiploader will be shared with Lumina. Lumina has stated to CCC that its development and operations strategy for the MTM facility anticipates shared usage with Cañariaco, and potentially other companies as well. Some service buildings and utilities throughout the port site, such as fire, process, and potable water supply and sanitary waste disposal, would also be shared, whereas concentrate receiving, storage, and reclaim facilities for the Cañariaco Norte product would be independently developed and operated by CCC.

Concentrate from the mine will be delivered to the port by truck, unloaded into a dump hopper, and conveyed to a storage building. The storage building has been sized to meet projected concentrate storage requirements and the anticipated schedule of vessel arrivals, including an allowance for vessel delays due poor weather or hurricane conditions en route to the port.

Concentrate reclaimed from storage will be conveyed initially to a sampling station, then to a transfer tower which feeds in to the shipping pipe conveyor, then to the berth conveyor, which in turn travels along the length of the berth to a towed tripper that feeds the travelling shiploader.

Power supply for the CCC facilities will be delivered to a substation at the concentrate storage and handling facility at 4.16 kV from the power auxiliary distribution system planned by Lumina. The substation will include incoming 4.16 kV switchgear and a step-down transformer.

As at the mine site, waste handling and disposal will be managed in an environmentally acceptable manner in accordance with legal requirements.

The marine berth will be constructed 3.2 km from shore at a water depth of -15.0 m, as required for the design 50,000 DWT Handymax vessel. Dredging to permit berth construction closer to shore was deemed uneconomical. The design and orientation of the berth are intended to minimize wave and swell action. A system of spread




moorings is intended to prevent ships from colliding with the berth during loading operations and to reduce loads on the mooring lines.

Vessel operations for export of CCC product will be scheduled in conjunction with Lumina to avoid berthing conflicts.

Accommodation

Accommodation facilities at the mine site will be constructed to house a construction crew of approximately 1,500 persons in a combination of single, double, and four-person occupancy rooms. The facilities will be converted to single- and double-occupancy rooms for the operations staff, projected at approximately 600 persons. Facilities will include areas for a kitchen, dining room, recreation, medical facility, office space for camp administration, and security for check-in/check-out.

Buildings

A two-storey main administration building will be constructed adjacent to the camp. The lower floor will be an open office arrangement and the second floor will have enclosed offices for senior staff. The building will include areas for reception, training, conferences, and storage. With adequate planning, it could be built early enough to accommodate the initial construction management and administration staff.

The mine truckshop will be the main services complex on site. The complex will contain maintenance shops for the mine mobile equipment fleet, warehouse space, first aid, a change room, dispatch area, lunchroom facilities, and offices for mine operations. It will also service all light- and medium-duty vehicles used on the site, including personnel buses and pickup trucks. A truck washdown pad, fuelling station, and ready-line for mine vehicles will be constructed nearby. A single-lane haul road will connect the truckshop to the pit area via the primary crusher station.

A mill maintenance shop adjacent to the process plant will service all stationary mill equipment such as pumps, instruments and motors.

Power Supply

Electrical power will be supplied to the Cañariaco Norte site by a 220 kilovolt overhead transmission line from the local utility substation at Carhuaquero, a distance of 57 kilometres from the mine site. The incoming transmission line will terminate at a new main site substation for the transformation of power from the transmission voltage level of 220 kilovolts to the site distribution/utilization level of 25 kilovolts. The




anticipated connected load for operation is 148 megawatts with an average load of 95 megawatts and power factor of 95 percent.

Power will be supplied to the various mine facilities through radial feeders originating at the main substation and routed on site through cable tray in pipe racks, either installed on overhead powerlines, direct buried, or in duct banks. Step-down transformers will provide equipment utilization voltages from the site distribution voltage. All process electrical and control rooms will be modular units constructed off site with all electrical controls and instrumentation equipment installed, wired, and completely tested before shipment to site.

Emergency power will be produced by standby diesel generating units sized to provide power to mine and process equipment and the permanent camp in the event of a utility power failure. The standby power plant will consist of three generator sets and be rated for a nominal 4.5 MW. In addition, an uninterruptible power supply (UPS) system will provide backup power to critical control systems, and emergency battery power packs will supply backup power to the fire alarm system and emergency egress lighting fixtures.

Water Supply

Fresh water for site services will be obtained from a fresh water reservoir constructed as part of the Río Cañariaco diversion around the waste rock management facility (WRMF). Fresh water will be used for process make-up water, fire water, and to supply the potable water treatment plant (PWTP). The reservoir will also provide start-up water for the process plant until sufficient reclaim water is available in the TMF to furnish continual supply, estimated to be two months after start-up. Normally, most process water is derived from thickener overflow, reclaim water pumped from the TMF, and other reclaim sources. Regular fresh water makeup is required for approximately 5% of total process flows, where the water has to be of higher quality.

The PWTP will be designed to provide sufficient potable water for the construction workforce. The water will be pumped to the camp and administration buildings and flow by gravity to the other site facilities. Both fresh and potable water will be trucked to the bulk emulsion plant as required.

There is expected to be limited impact on down-stream water users.

Should discharge of water to the environment be required, the water management system will ensure that discharge waters meet Peruvian regulations. CCC has incorporated design parameters for site water management that will minimize the amount of “contact water” generated by Project construction and operation.




Communications

The site communications system will cover the complete voice, data, fax, Internet/email and radio communications requirements for the process and camp facilities. The system will be designed to meet all regulatory requirements needed for operating approvals. The mine communications system will be independent of the rest of the system.

Fuel

Diesel fuel will be stored in four 500 m3 tanks, providing more than two weeks of supply the event of fuel delivery disruption or power failure. The fuel will be used to run the construction / emergency generators and will be piped to the re-fuelling station near the truckshop.

Waste

Solid waste from the plant site will be handled and disposed of in an environmentally acceptable manner in accordance with legal requirements and best industry practices. As applicable, waste will either be incinerated, disposed of in a landfill, recycled, or removed off site to special handling facilities.

A sewage treatment plant (STP), sized for the complete construction workforce, will be built on site. The plant will be self-contained in an enclosure housing all mechanical and electrical equipment and will be designed to produce effluent that complies with Peruvian regulations for effluent quality. Sewage will be collected from the different sources through a network of pipelines, holding tanks, lift stations, and a pumper truck, where necessary. Treated effluent will be discharged directly to the environment during construction and to the TMF during operations.

5.4 Physiography

Elevations in the Project area range between 2,200 and 3,600 masl, as the Project is situated on the eastern side of the continental divide. The topography varies from steep incised valleys at lower elevations to open grassy highlands, locally known as jalcas or paramus, at upper elevations. Within the area that mineral resources have been estimated, the topography is steep with fingers of sub-tropical forests extending up the small valleys. These forests transition into open grasslands and broad valleys as the elevation increases above 3,000 masl to the north of the main mineralized zone.

The property crosses several climate zones, and temperature differences between the lower forested valleys and the upper grassy ridges can be up to 10°C.




Vegetation at higher elevations consists mainly of ichu and other types of natural grasses used for livestock grazing. Localized agriculture plots for subsistence farming are maintained within the forested valleys. The community of Cañaris, north of the Project area within temperate highlands, is agriculturally-based. The main crops are potatoes, maize, and a variety of vegetables and herbs.


In the opinion of the QPs:

• There is sufficient suitable land available within the concessions for any future tailings management facility, mine waste rock facility, and installations such as a process plant and related mine infrastructure.

• A review of the existing and likely power and water sources, manpower availability, and transport options indicate that there are reasonable expectations that sufficient labour and infrastructure is available to support declaration of Mineral Resources.




6.0 HISTORY

Copper was first discovered at Cañariaco Norte in the period 1967–1970 during a regional stream sediment sampling program carried out by the Peruvian Servicio Nacional de Geología y Minería (Ingemmet), in co-operation with a British Geological Survey team. This work identified a copper anomaly, A-2. Subsequently from 1971 to 1974, Ingemmet undertook a detailed geochemical study of the Cañariaco River valley, which delineated four centres of alteration and copper mineralization. The geochemistry was supported by geological mapping, rock chip and soil sampling, induced polarization (IP) and resistivity ground geophysical surveys, and core drilling of five drill holes (1,500 m). Thin section and petrographic studies were also performed. A mineral resource estimate was undertaken at the conclusion of this work.

Placer Dome Exploration Inc. (Placer Dome) optioned the area from the Peruvian Government in 1994. Work completed included geological mapping, rock chip sampling, trenching (2,200 m) and pitting (80 test pits), petrographic studies, re-interpretation of the available Ingemmet IP and resistivity data, a very preliminary estimate of tonnage and grade, and three core holes (874 m). This work identified a porphyry copper system related to breccias and a Tertiary intrusive complex emplaced in a volcanic basement of andesite to dacite tuffs and ignimbrites. However, in 1997, the option was not exercised and reverted to the Government.

During 1999, Billiton Exploration and Mining Perú B.V. (Surcusal Peruana) took up an option from the Peruvian Government. The work program comprised geological mapping, soil and rock chip sampling, IP and ground magnetic geophysical surveys, and eight core holes (957.7 m). Petrographic samples were collected and studied and leach testing was carried out on some of the drill core. A resource estimate was completed in 1999. The option was dropped in 2000, with the property returning to the Peruvian Government.

The property was subsequently put up for auction. Candente Copper Corp., through its wholly-owned Peruvian subsidiary Cañariaco Copper Peru S.A. (CCP), formerly Exploraciones Milenio S.A, (EMSA), acquired 100% ownership of the Project in February 2002. Since that date, CCC has completed geological mapping, prospecting, ground magnetic geophysics, rock chip sampling, petrographic studies, bulk sampling for metallurgical testing and re-logging of existing drill core. A total of 257 core holes (74,058 m) were drilled to the end of 2008. Mineral resources were first estimated on behalf of CCC in 2004, and were updated in 2006 and 2008.

A preliminary assessment (PA) was undertaken in 2006. The study envisaged conventional open pit mining of the Cañariaco Norte deposit, with mineralization




crushed, then placed on a heap leach pad. Copper recovery was planned through a solvent extraction/electrowin (SX–EW) plant to produce copper cathode. Mining would commence at 30,000 t/d and expand to 60,000 t/d. The financial analysis, using the costs and assumptions at the time, indicated a positive financial outcome.

In 2008, a revised PA (the 2008 PA) was undertaken, because additional leach testing had indicated that heap leaching recoveries were erratic, and a more robust metallurgical process was required. The production rate from the proposed open pit was 75,000 t/d, feeding a conventional process plant consisting of semi-autogenous grind (SAG) and ball mills and a flotation circuit to produce copper concentrate. The financial analysis, using the costs and assumptions at the time, indicated a positive financial outcome. The mineral resource update and the PA study discussed in this Report supersede the resource estimate used as the basis of the 2008 PA and the 2008 PA results should be considered historical.

In 2010, CCC commissioned a mineral resource update. The updated mineral resources, information on the progress of the pre-feasibility study on the Project, and the PA-level economic analysis on those Mineral Resources forms the subject of the remainder of this Report.




7.0 GEOLOGICAL SETTING

A regional geological plan for the Project area is shown in Figure 7-1.

Basement rocks comprise pelitic schists of the Precambrian to Early Paleozoic Olmos Complex that are unconformably overlain by Late Triassic-Early Jurassic La Leche Formation marine sediments that have minor intercalated volcanic units. An early to Late Jurassic volcano-sedimentary sequence, the Oyotún Formation, overlies the earlier units. Following regional uplift, erosion, and subsequent subsidence, the lower portion of the Goyllarisquizga Group, a regionally extensive quartz arenite, was unconformably deposited on the earlier lithologies. In turn, the arenite was overlain by Early Cretaceous to mid-Late Cretaceous marls, shales, and limestone.

In the Early Tertiary, volcanic units of the Llama and Porculla Formations of the Calipuy Group were erupted, and were followed by uplift and erosion. Renewed volcanic activity commenced with the eruption of volcanic rocks of the Huambos Formation during the Late Tertiary. Porphyry stocks, breccias and dikes that are also Late Tertiary in age intrude the Cretaceous and Tertiary units (Figure 7-2). Age dates using K/Ar and Re–Os on intrusive rocks, breccias and alteration minerals in the Cañariaco Norte deposit returned dates that range in age from 15.8 Ma to 17.9 Ma.

A number of circular features/intrusions have been identified district-wide (refer to Figure 7-1). A single circular feature that measures 8 km x 10 km encompasses all three mineralized centres in the Project, including Cañariaco Norte, Cañariaco Sur and Quebrada Verde. This feature is centred on, or close to major fault intersections suggesting that the emplacement of the intrusive complex has been localized by fault intersections.

Faults that have been identified at a regional scale consist of two, parallel, long-range, northwest–southeast-trending district-scale faults. One bisects the Project area, and the second fault is approximately 7 km to the northeast. The northwest–southeast faults have a probable conjugate northeast–southwest fault set, which tend to be of medium-range. These faults appear to control the location and development of the intrusive–breccia complexes and related mineralization and alteration in the Project area.




Figure 7-1: Regional Geology Map

Note: Figure courtesy Candente Copper Corp. Tenure outlines shown on the plan are superseded by the tenure outlines in Figure 4-1. Grid squares on the plan are 4 km x 4 km




Figure 7-2: Regional Stratigraphic Column


7.1 Cañariaco Norte Deposit Geology

The Cañariaco Norte intrusive–breccia is hosted within a multiphase intrusive–breccia complex that is approximately 1.7 km in strike extent and 1.1 km wide. The deposit has been drill-tested to a depth of approximately 770 m and remains open at depth. The majority of the copper-gold mineralization is hosted within the intrusive and breccia units, but locally extends for variable distances into the enclosing volcanic units. Intrusive units comprise approximately 55–60% of the deposit, breccias approximately 30–35%, and pre-mineral volcanic rocks approximately 5–10%.

The intrusive units are nested and collectively roughly oval in shape with older intrusive rocks being cut by successively younger intrusive bodies. In general, the intrusive units are north–south-trending, steeply-dipping bodies. The breccia units cut the intrusive units, are oval to circular in shape and are steeply plunging. The dykes generally strike northwest–southeast with a steep southwesterly dip. The shape and positioning of the intrusive, breccia and dyke units was largely controlled by northwest–southeast and northeast–southwest-trending faults.




7.1.1 Lithologies

Table 7-1 presents the stratigraphy for the Project area.

Table 7-1: Project Stratigraphic Column Unit Rock Code Rock Type Description Breccias IBxP Breccia (youngest) Polymictic breccia TBx Breccia (middle) Tourmaline breccia IBxH Breccia (oldest) Hydrothermal breccia Intrusives FP Dykes Feldspar porphyry dykes CQP Intrusive (youngest) Coarse quartz porphyry BFP Intrusive (middle) Biotite feldspar porphyry CQFP Intrusive (oldest) Crowded quartz feldspar porphyry VC Calipuy Volcanic Pre-mineral andesitic, dacitic and rhyolitic volcanics

The oldest rocks, ascribed to the Calipuy Group, are a series of dacite tuffs with lesser, bedded, rhyolite tuffs overlain by andesite porphyry flows and pyroclastic rocks. Andesite pyroclastic rocks and flows dominate on the western, northern and southern sides of the intrusive complex where elevations are higher. The eastern side of the intrusive complex is lower in elevation, and thus the dacite and rhyolite volcanic rocks that underlie the andesite volcanic rocks are exposed adjacent to the intrusive complex.

Three major intrusive generations have been identified. The oldest intrusive unit is a crowded quartz–feldspar porphyry (CQFP), which consists of 1–3% quartz eyes, 35% feldspar phenocrysts, and 3–5% hornblende–biotite. Grain sizes of the constituents range 1–2 mm. The unit is interpreted as dioritic in composition.

The middle intrusive unit is a biotite–feldspar porphyry (BFP) that is interpreted to be granodioritic in composition, with 3–5% quartz eyes, 15–20% feldspar phenocrysts, 2-5% euhederal biotite, and traces of hornblende. Grain sizes typically range from 3-5 mm.

The youngest intrusive unit, a coarse quartz porphyry (CQP), consists of 5–10% quartz eyes (grain size range from 3–5 mm), 15–20% euhedral feldspar crystals (<3 mm) and traces of biotite (1–2 mm). The CQP is interpreted to be of quartz monzonitic composition.

The youngest intrusive stage is a set of feldspar porphyry dykes/breccias (FP) that range in thickness from 2 m to 30 m, and have a northwest–southeast strike, with a steep, southwesterly dip. The dykes have a fine-grained ground mass with 10–20% feldspar phenocrysts (3–10 mm) and 5–10% hornblende phenocrysts (2–8 mm). The dykes commonly display cooling contacts. Where the dykes intersect the breccias, dyke fragments occur as large (>10 m), rotated, and weakly-fractured blocks within the




breccias. The dykes have only been weakly altered, and contain minor copper mineralization where they have been brecciated by the hydrothermal breccia unit.

Three breccia bodies that post-date the intrusive rocks are recognized. The oldest unit, hydrothermal breccia (IBxH), consists of matrix-supported angular to sub-angular biotite–feldspar porphyry and crowded quartz–feldspar porphyry fragments that display little or no evidence of transport. Fragment sizes within the central portion of the breccia are generally 1–5 cm in a fine-grained matrix. Near the southern margin of the breccia, there is a high component of feldspar porphyry dyke fragments up to tens of meters in size. The unit is not well mineralized, with copper grades related to the inclusion of mineralized porphyry fragments. The margins of the hydrothermal breccia can show crackle brecciation.

The middle tourmaline breccia (TBx) has a fine-grained matrix consisting of quartz and tourmaline. The unit is extensive, and was emplaced along the northeastern margin of the intrusive bodies. Breccia fragments are angular to sub-angular, 1–10 cm in size, and include clasts of the biotite–feldspar porphyry, crowded quartz–feldspar porphyry and the hydrothermal breccia. The breccia is not mineralized, apart from copper grades related to the inclusion of mineralized porphyry fragments.

The youngest breccia unit, polymictic breccia (IBxP), is a late-stage breccia with an erratic shape, cross-cutting all earlier units. Sub-rounded to rounded clast fragments include vein quartz, all three intrusive units, and the two earlier breccia phases. Fragments range from 0.5–10 cm with the breccia margins often grading into a crackle breccia. The unit shows multiple breccia pulses, the last of which is a fluidized micro-breccia with rounded fragments that are typically <3 mm in size, crosscutting all other pulses. Copper mineralization occurs in both the matrix and the fragments.

Mineralization occurs primarily as disseminations, and in fractures, sulphide and quartz veins, faults and breccias. Fracture density is the single most important factor influencing copper grades and alteration intensity, although breccias and faults can locally be important.

Copper mineralization was introduced as a series of events closely following the emplacement of each of the intrusive units, and the polymictic breccia unit. Initially, copper mineralization comprised chalcopyrite, minor bornite, introduced following emplacement of each of the crowded quartz–feldspar porphyry and biotite–feldspar porphyry units, with the greatest amount introduced following emplacement of the biotite–feldspar porphyry unit. This event was followed by introduction of chalcocite, covellite, minor tennantite–digenite, following emplacement of the coarse quartz porphyry unit. The mineralizing process terminated with enargite, chalcocite and




covellite, minor tennantite–digenite, introduced concurrently with emplacement of the polymictic breccia unit.

Near surface, the deposit has been intensely weathered, resulting in the formation of a leached cap that contains less than 0.05% Cu, trace pyrite and tenorite and variable concentrations of limonite, goethite, jarosite, and hematite. The leached cap varies significantly in thickness, ranging from less than a metre to as much as 120 m, averaging approximately 40–50 m in thickness. The water table is at, or near, surface.

A geological plan for the deposit is presented in Figure 7-3. Figure 7-4 and Figure 7-5 are drill sections through the deposit.

7.1.2 Alteration

The alteration forms distinct concentric zones with a central potassic alteration, central to intermediate, overlapping, and partly overprinting phyllic, argillic and advanced argillic alteration and fringing propylitic and minor silicic alteration. Alteration intensity is directly related to the intensity of fracturing in the hosting lithologies and brecciation in the polymictic breccia unit. Locally, alteration distribution and intensity is controlled by northeast- and northwest-trending faults.

Phyllic and argillic alteration dominate in the upper 50 to 300 m in the southern half of the deposit and the upper 100 m to locally 150 m in the northern half of the deposit (excluding overburden and the leached cap). Most of the area of the northern and southern halves of the deposit at depth under the layer of phyllic and argillic alteration is dominated by potassic alteration.




Figure 7-3: Geological Map, Cañariaco Norte





Figure 7-4: Geological Drill Section 9,326,300N





Figure 7-5: Geological Drill Section 690,800E





Alteration type descriptions are as follows:

• Potassic alteration consists of variable proportions of secondary potassic feldspar and biotite, chlorite and magnetite. Associated with deposition of chalcopyrite (bornite).

• Propylitic alteration comprises illite, chlorite, epidote and smectite. Silicic alteration comprises very fine grained pervasive silica. Associated with deposition of chalcopyrite and minor bornite. Found mainly at deposit margin, but can locally occur within the deposit. Can extend for significant distances outside of the intrusive–breccia complex into the bordering volcanic rocks.

• Phyllic alteration includes sericite and quartz with quartz veinlets and stockworks (0.5 cm to 1.5 cm thick). Associated with deposition of chalcocite, covellite and minor tennantite and digenite. Commonly intermixed with argillic alteration. Affected parts of the crowded quartz–feldspar porphyry and biotite–feldspar porphyry units and all of the coarse quartz porphyry and polymictic breccia units. Late-stage phyllic alteration has extended along northeast and northwest-trending faults outwards from the polymictic breccia unit, and affected units distal from the polymictic breccia unit

• Argillic alteration consists of kaolinite and illite. Associated with deposition of chalcocite, covellite and minor tennantite and digenite. Affected parts of the crowded quartz–feldspar porphyry and biotite–feldspar porphyry units and all of the coarse quartz porphyry and polymictic breccia units.

• Advanced argillic alteration comprises alunite, kaolinite, pyrophyllite and dickite. Associated with deposition of enargite, chalcocite and covellite (tennantite and digenite). Primarily associated with the polymictic breccia unit, but also has variably affected the bordering coarse quartz porphyry and biotite–feldspar porphyry units. Has extended along northeast and northwest-trending faults outwards from the polymictic breccia unit, and affected units distal from the polymictic breccia unit.

• Silicic alteration is found mainly at the deposit margin, but can locally occur within the deposit. Can extend for significant distances outside of the intrusive–breccia complex into the bordering volcanic rocks.

7.1.3 Structures

The northwest–southeast- and the northeast–southwest-trending fault sets appear to control the emplacement of the intrusive phases at Cañariaco Norte, as well as the distribution of copper mineralization and alteration. In addition, the faults form structural corridors for late-stage polymictic breccias and associated alteration and




mineralization. The smaller set of northwest–southeast-oriented faults control dyke emplacement.

7.2 Prospects

Prospect locations are shown in Figure 7-6.

Figure 7-6: Prospect Location Plan and Geological Map





7.2.1 Cañariaco Sur

The Cañariaco Sur deposit is located 1.3 km southwest of Cañariaco Norte, and is situated on a coincident copper–gold–molybdenum anomaly with approximate dimensions of 1,200 m x 800 m. CCC has interpreted the prospect as a porphyry-copper centre dominated by potassic alteration with chalcopyrite–bornite–molybdenite mineralization and anomalous Cu–Au–Mo grades.

The host rocks to the mineralized intrusions are Calipuy Group volcanic rocks. Two main intrusions were noted from geological mapping, an early 750 m x 500 m quartz monzonite stock, and a later diorite stock with an area of approximately 100 m x 50 m. A third intrusive phase may also exist (Sillitoe, 1999).

7.2.2 Quebrada Verde

The Quebrada Verde prospect is 3.2 km south of Cañariaco Norte and 1.5 km south of Cañariaco Sur. It consists of a 1 km x 750 m diorite porphyry stock that intruded Calipuy Group volcanic rocks (Wilson, 1999). A satellite stock of the same diorite porphyry measuring 400 m x 400 m is located 200 m to the northeast. An east–west striking, post-mineralization granodiorite dike, which has dimensions of approximately 1 km x 100 m, intruded the volcanic rocks and the larger diorite porphyry stock.

Soil and drill core assay data suggest a concentric mineralization zonation with a wide, low-grade copper ‘ring’ hosting elevated gold values along the inner edge and elevated molybdenum values on the outer edge.


In the opinion of the QPs, knowledge of the deposit settings, lithologies, and structural and alteration controls on mineralization within the Cañariaco Norte deposit is sufficient to support Mineral Resource estimation.

The near-by Cañariaco Sur and Quebrada Verde prospects are at an earlier stage of exploration, and the lithologies, structural, and alteration controls on mineralization are currently insufficiently understood to support estimation of Mineral Resources.




8.0 DEPOSIT TYPES

The Cañariaco Norte deposit is considered to be an example of a primary porphyry-copper system. The following discussion of the typical nature of porphyry-copper deposits is sourced from Sillitoe, (2010), Berger et al., (2008), and Sinclair (2006).

Porphyry-copper systems commonly define linear belts, some many hundreds of kilometres long, as well as occurring less commonly in apparent isolation. The systems are closely related to underlying composite plutons, at paleo-depths of 5 km to 15 km, which represent the supply chambers for the magmas and fluids that formed the vertically elongate (>3 km) stocks or dike swarms and associated mineralization.

Commonly, several discrete stocks are emplaced, resulting in either clusters or structurally controlled alignments of porphyry-copper systems. The rheology and composition of the host rocks may strongly influence the size, grade, and type of mineralization generated in porphyry-copper systems. Individual systems have life spans of circa 100,000 years to several million years, whereas deposit clusters or alignments, as well as entire belts, may remain active for 10 million years or longer.

Deposits are typically semicircular to elliptical in plan view. In cross section, ore-grade material in a deposit typically has the shape of an inverted cone with the altered, but low-grade, interior of the cone referred to as the “barren” core. In some systems, the barren core may be a late-stage intrusion.

The alteration and mineralization in porphyry-copper systems are zoned outward from the stocks or dike swarms, which typically comprise several generations of intermediate to felsic porphyry intrusions.

Porphyry-copper mineralization occurs in a distinctive sequence of quartz-bearing veinlets as well as in disseminated forms in the altered rock between them. Magmatic-hydrothermal breccias may form during porphyry intrusion, with some breccias containing high-grade mineralization because of their intrinsic permeability. In contrast, most phreatomagmatic breccias, constituting maar–diatreme systems, are poorly mineralized at both the porphyry copper and lithocap levels, mainly because many such phreatomagmatic breccias formed late in the evolution of systems.

Copper mineral assemblages are a function of the chemical composition of the fluid phase and the pressure and temperature conditions affecting the fluid. In primary, unoxidized or non-supergene-enriched ores, the most common ore–sulphide assemblage is chalcopyrite ± bornite, with pyrite and minor amounts of molybdenite. In supergene-enriched ores, a typical assemblage can comprise chalcocite + covellite ± bornite, whereas, in oxide ores, a typical assemblage could include malachite +




azurite + cuprite + chrysocolla, with minor amounts of minerals such as carbonates, sulphates, phosphates, and silicates. Typically, the principal copper sulphides consist of millimetre-scale grains, but may be as large as 1–2 cm in diameter and, rarely, pegmatitic (larger than 2 cm).

Alteration zones in porphyry-copper deposits are typically classified on the basis of mineral assemblages and consist of potassic, propylitic, phyllic and argillic zones. In silicate-rich rocks, the most common alteration minerals are K-feldspar, biotite, muscovite (sericite), albite, anhydrite, chlorite, calcite, epidote, and kaolinite. In silicate-rich rocks that have been altered to advanced argillic assemblages, the most common minerals are quartz, alunite, pyrophyllite, dickite, diaspore, and zunyite. In carbonate rocks, the most common minerals are garnet, pyroxene, epidote, quartz, actinolite, chlorite, biotite, calcite, dolomite, K-feldspar, and wollastonite. Other alteration minerals commonly found in porphyry-copper deposits are tourmaline, andalusite, and actinolite.


In the opinion of the QPs, the Cañariaco Norte deposit is considered to be an example of a porphyry system based on the following:

• Multiple emplacements of successive intrusive phases and a variety of breccias

• Copper-bearing igneous rocks are intrusive into host volcanic and sedimentary rocks

• Mineralization is spatially, temporally, and genetically associated with the intrusive–breccia activity and hydrothermal alteration of the intrusive and breccia bodies

• Large zones of veining and stockwork mineralization, together with minor disseminated and replacement mineralization occur throughout large areas of intrusive–breccia and hydrothermally-altered rock

• Hydrothermal alteration is extensive and zoned, which is common to porphyry copper deposits. The alteration assemblages are consistent with the physico-chemical conditions of a porphyry environment

• Mineralization is focused in well-developed quartz–sulphide stockworks; veins, crackle and breccia zones are also present

• Tenor of copper and gold grades

• Large tonnage.




9.0 MINERALIZATION

A discussion of the mineralization styles and related depth, width (thickness), orientation and continuity is presented for the Cañariaco Norte deposit in Section 7. The discussion in this section of the Report relates to the mineralization type, character, and mineralogy of the deposit.

Although all of the units listed in Table 7-1 host mineralization, the average grades vary. In general, the highest copper grades (>0.5% Cu) occur in the biotite–feldspar porphyry and polymictic breccia units. Intermediate copper grades (0.35%–0.5% Cu) are most common in the hydrothermal breccia and the coarse quartz porphyry and the lowest grades (0.2%–0.35% Cu) are found in the crowded quartz feldspar porphyry, tourmaline breccia and volcanic host rocks. The feldspar porphyry dykes typically do not contain copper grades unless the dykes have been fractured, brecciated, or crackle-brecciated near the south margin of the hydrothermal breccia, where they have been weakly mineralized (average copper grade ranges of 0.1%–0.2% Cu).

A Qemscan examination of core samples selected for metallurgical testwork in 2008, and performed by SGS Metallurgical Laboratories in Santiago, Chile, indicated the predominant minerals and mineral associations present in the Cañariaco Norte samples are:

• Copper mineral species: dominantly chalcopyrite, covellite, chalcocite and enargite, with minor bornite, tennantite and digenite and trace tetrahedrite. Copper sulphide minerals are mainly associated with pyrite and other sulphide gangue minerals

• Non-metallic gangue minerals: quartz and micas/biotite/chlorite/clay group

• Sulphide gangue mineral: pyrite. Pyrite is present as liberated pyrite or in association with other sulphide minerals.

Copper mineralization paragenesis comprised the following stages:

• Initially chalcopyrite (bornite) + pyrite • Intermediate, chalcocite and covellite (tennantite, digenite) ± pyrite • Lastly enargite, chalcocite and covellite (tennantite, digenite) + pyrite.

Copper minerals and pyrite are disseminated, veined, and fracture-hosted, with copper grades directly related to the intensity of fracturing and alteration type and intensity. Higher grades are associated with potassic, phyllic, and argillic alteration and less commonly with propylitic and silicic alteration.




CCC staff interpret that chalcopyrite (bornite) was deposited initially with potassic and propylitic alteration. As the system evolved chalcocite and covellite (tennanite, digenite) developed in association with phyllic–argillic alteration, and overprinted and replaced chalcopyrite (bornite) and potassic alteration in the central and upper parts of the deposit. Enargite, chalcocite and covellite (tennantite, digenite), associated with advanced argillic and phyllic alteration was the last to form. These minerals are typically restricted in distribution to the polymictic breccia unit and the adjacent coarse quartz porphyry and biotite–feldspar porphyry unit, where they occur as disseminations, clots, veins, and fracture coatings cutting earlier-formed chalcopyrite, chalcocite, covellite, potassic, phyllic, and argillic alteration.

Chalcopyrite (bornite), chalcocite, covellite, enargite (tennantite, digenite) are primarily hypogene in origin, with only limited development of supergene chalcocite–covellite. In the opinion of CCC staff, between the water table being at or near the present topographic surface, and an active erosive environment, any supergene chalcocite/covellite that formed is being rapidly removed. Locally thin (<30 m) discontinuous layers of supergene chalcocite–covellite occur immediately under the leach cap.

Pyrite is common in all alteration types, averaging approximately 5% throughout the deposit as disseminations, veins, and fracture coatings. Magnetite is less common and primarily associated with chalcopyrite (bornite) in potassic alteration. Tourmaline is restricted to the matrix of the tourmaline breccia unit.

Gold and silver values are anomalous throughout the deposit; however, higher gold grades only occur with higher copper grades. Gold grades range, on average, between 0.04 g/t Au and 0.11 g/t Au and silver grades average 1.3 g/t Ag–2.5 g/t Ag throughout the different rock types in the deposit (Bonson et al., 2008).

Molybdenum (Mo) grades are low, averaging <40 ppm and are slightly higher on the margins and at depth in the deposit (Bonson et al., 2008). Molybdenum grades are unlikely to be economically recoverable based on current information.

Figure 9-1, Figure 9-2 and Figure 9-3 display cross-section through the Cañariaco Norte deposit showing copper, gold, and silver grades, respectively.


In the opinion of the QPs, the mineralization style and setting of the deposit is sufficiently well understood to support Mineral Resource estimation.




Figure 9-1: Example Vertical Section (9,326,300N) with 15 m Composites Coloured by Cu (%) Ranges Overlapping Lithological Interpretation

Note: grid squares are 200 m x 200 m




Figure 9-2: Example Vertical Section (9,326,300N) with 15 m Composites Coloured by Au (ppb) Ranges





Figure 9-3: Example Vertical Section (9,326,300N) with 15 m Composites Coloured by Ag (g/t) Ranges





10.0 EXPLORATION

Exploration activities such as geological mapping and sampling have been performed by CCC and predecessor companies, Ingemmet, Placer Dome, and Surcusal Peruana. Contractors were used for activities such as geophysical surveys.

Exploration activities on the Project have included regional and detailed mapping, stream sediment, grab, rock, and soil sampling, trenching and pitting, core drilling, ground geophysical surveys, mineralization characterization studies and metallurgical testing of samples. Petrographic studies and density measurements on the different lithologies have also been carried out.

A summary of the work programs completed to the Report effective date are summarized in Table 10-1.

10.1 Grids and Surveys

The Project uses WGS-84 UTM (zone 17). Data collected to June 2006 used PSAD-56 (zone 17). All survey work on the Project prior to the re-establishment of survey control points in 2004 and a datum switch in June 2006 was re-done.

Topographic data used to constrain Mineral Resources were based on aerial photograph coverage provided by Horizons South America S.A.C. Aerial Mapping Services.

10.2 Geological Mapping

Regional and detailed geological mapping was completed by Ingemmet, Placer Dome, and CCC in a number of phases. Map scales varied from regional (1:25,000) to prospect scale (1:1,000). Map results were used to identify lithologies, areas of quartz veining, alteration, silicification and sulphide outcrop that warranted additional work. Interpretation of air photos were used to identify areas that required geological mapping and sampling. During the mapping program, structural measurements were collected from faults, veins, and fractures to provide additional structural detail for geological interpretations.




Table 10-1: Exploration Summary Table Year Operator Work Undertaken 1967–1970 Ingemmet Regional stream sediment sampling. Identified anomalous copper in drainages surrounding the

Cañariaco Norte deposit 1971–1974 Ingemmet Geological mapping at 1:25,000 scale defined 9 km2 of alteration and mineralization, with four

distinct centres identified. Limited soil and rock chip sampling; numbers of samples unknown Nine trenches; locations unknown. A total of 23 rock chip samples taken from the trenches and analysed for Cu and Mo. Infill geological mapping at 1:5,000 scale; this detailed mapping was hampered by dense vegetation and steep topography IP and resistivity ground geophysics over Cañariaco Norte Five vertical core drill holes completed at Cañariaco Norte to depths of 300 m for 1,500 m total drilling. A total of 579 core samples were collected along 3 m or 1.5 m lengths and analysed for Cu and Mo. A total of 66 composite geological samples were analysed for total copper (TCu), Mo, Ag, and Au. Thin section and petrographic studies on selected drill core samples Mineral resource estimate

1994–1997 Placer Dome

Geological mapping at 1:1,000 scale at Cañariaco Norte; preliminary geological mapping, scale not known, at Cañariaco Sur Soil and rock chip sampling over an area of 1.3 km by 1.5 km at Cañariaco Norte. A total of 715 rock chip samples assayed for Au, Cu, Mo, (and some for Ag and As). Reconnaissance rock chip and grab sampling at Cañariaco Sur Trenching (2,200 m) and pitting (80 test pits) at Cañariaco Norte and Cañariaco Sur Re-interpretation of Ingemmet IP and resistivity data; the re-interpretation noted some correlation between copper grades and chargeability and/or resistivity. It was also noted that some of the strongest IP targets had not been drilled by Ingemmet Three core holes (874 m) at Cañariaco Norte Petrographic studies on selected drill core samples Mineral resource estimate

1999–2000 Surcusal Peruana

Geological mapping; scale unknown Soil and rock chip sampling. Sampled outcrops in the streams south of the drilled area at Cañariaco Norte. Low-level Cu anomalies (<500 ppm) were collected from phyllically-altered quartz porphyry at Cañariaco Norte. The northern side of the quartz porphyry generated additional surface copper anomalies (500–2000 ppm Cu) were found with outcrops of basaltic andesite. In the outcrops around the drill holes in the northern part of the quartz porphyry, gold graded in the 100–200 ppm Au range. In the southern half of the system, the grades were generally in the 40–100 ppb Au range. Soil samples were only taken at Cañariaco Sur and Quebrada Verde, and outlined co-incident Cu, Au and Mo anomalies over a 750 m diameter area with smaller anomalies of copper and gold adjacent to this target. Soil sample line spacing was typically 200 m with 100 m infill where initial results were favourable. Sample spacing was 50 m along the lines. IP and ground magnetic geophysical surveys on four lines at Cañariaco Norte. For the IP survey, a dipole–dipole array, with electrode spacing of 100 m, was used. The results were not useful because only a very weak current was received at the potential electrode. The ground magnetic data used the same lines and spacings; no domains of magnetite destruction could be defined. Eight core holes (957.7 m) drilled at Cañariaco Norte, Cañariaco Sur, and Quebrada Verde Petrographic studies on selected drill core samples Leach testing on on 12 samples from two intervals of core; results indicated that the mineralization was potentially bio-heap leachable. Mineral resource estimate

2002 to present

CCC Geological mapping, including 1:2,000 scale at Cañariaco Norte and Cañariaco Sur. Preliminary mapping of the Cañariaco Sur and Quebrada Verde zones was also carried out to assess soil geochemistry anomalies outside of the main mineralized zones. Zones of silicification and quartz vein stockworks occur peripheral to the main circular feature at Cañariaco Sur. Structural measurements. A total of 311 structural measurements including faults, veins, fractures etc. were collected Prospecting, rock chip and grab sampling. A total of 148 rock samples were collected in 2002–2004 from surface for gold, total copper and 35-element inductively-coupled plasma (ICP) analyses. In addition, a number of PIMA and petrology samples were collected to define favourable alteration including alunite and sericite and to define the various igneous phases. A




Year Operator Work Undertaken total of 392 soil geochemistry samples and 355 rock samples were collected over the geophysical grid at Cañariaco Sur and Quebrada Verde in 2008. Soils with elevated levels of copper of up to 5,720 ppm (0.57%) and gold up to 330 ppb covering over an area of 0.9 km x 1.4 km and potassic alteration mapped over a length of 2.3 km at Cañariaco Sur. Anomalous levels of copper of up to 2,200 ppm (0.22%) and up to 497 ppb gold in soils cover an area of approximately 0.7 km x 0.9 km at Quebrada Verde. Re-logging of existing drill core 20 line km of IP/resistivity and ground magnetic geophysics; outlined areas of magnetite destruction and delineated the various alteration zones and helped confirm and/or identify old/new fault structures at Cañariaco Norte. A coincident magnetic high; IP/resistivity low and chargeability high identified at Cañariaco Sur. A large IP chargeability high centered on a resistivity anomaly and covering an area of 0.8 km by 2.0 km identified at Quebrada Verde. Petrographic studies on selected drill core samples Four bulk samples for leach testing collected in 2004. 250 core holes (72,116,1 m) at Cañariaco Norte, Cañariaco Sur and Quebrada Verde

10.3 Geochemistry

Stream sediment, soil and rock chip sampling were used to evaluate the mineralization potential and generate targets for core drilling at Cañariaco Norte, Cañariaco Sur, and Quebrada Verde. Sampling was performed by Ingemmet, Placer Dome, Surcusal Peruana, and CCC.

Stream sediment sampling identified elevated copper and gold values. Rock chip sampling outlined copper–gold–silver–molybdenum mineralization associated with quartz porphyry outcrops at Cañariaco Norte, Cañariaco Sur, and Quebrada Verde. Soil sampling provided the initial drill target areas for Cañariaco Norte and a follow-up exploration target at Quebrada Verde (Figure 10-1).

PIMA alteration studies undertaken by CCC were used to define favourable alteration styles for mineralization, including alunite and sericite, and to define the various igneous phases.

10.4 Pits and Trenches

Nine trenches were excavated by Ingemmet, and a total of 23 rock chip samples taken. Placer Dome completed trenching (2,200 m) and pitting (80 test pits) at Cañariaco Norte and Cañariaco Sur. Test data are superceded by drill data.

10.5 Geophysical Surveys

Geophysical surveys, comprising IP, resistivity, and ground magnetic methods, were completed at Cañariaco Norte, Cañariaco Sur, and Quebrada Verde by Ingemmet, Surcusal Peruana, and CCC.




Results of the surveys prior to the CCC work were considered inconclusive, due to wide line spacings. CCC completed 20 line km of combined IP/resistivity and ground magnetic surveys. The surveys at Cañariaco Norte outlined areas of magnetite destruction and addition, thus delineating the various alteration zones and helping to confirm and/or identify old/new fault structures. A coincident magnetic high, IP/resistivity low and chargeability high was identified at Cañariaco Sur. A large IP chargeability high, centered on a resistivity anomaly and covering an area of 0.8 km by 2.0 km, was identified at Quebrada Verde.

10.6 Drilling

Drilling completed on the Project is discussed in Section 11 of the Report.

10.7 Bulk Density

Bulk density data collected to date on the Project are discussed in Section 12 of this Report.

10.8 Geotechnical and Hydrogeology

Geotechnical drilling has been undertaken to test the orientations of the feldspar porphyry dikes and to test the orientation of the polymictic breccia within the crowded quartz feldspar porphyry. All drill holes after drill hole C06-045 have been logged for standard geotechnical parameters. Following a visit by SRK personnel in late 2008, geotechnical logging parameters were revised. All drill holes after C07-189 were logged using the revised protocols.

CCC has also carried out geotechnical mapping of trench exposures (veins, joints and faults), and strength testing using a point load tester.

Some of the core drill holes had piezometer pipe installed and 12 of those holes had packer tests completed in them, as part of a hydrogeological data acquisition program overseen by Ground Water International. Two holes were drilled specifically for hydrogeological testing on the ridge to the west of the area that has estimated mineral resources.

A geotechnical drilling program to provide additional information on rock mass properties, joint patterns and hydrogeology within the proposed pit area is planned. AMEC notes that information from this program could result in changes to the slope angle of the pit wall as a result of more detailed geotechnical information, and such




changes would affect the pit shell used to constrain the mineral resources in Section 17. CCC expects the drilling will be undertaken in 2011.

10.9 Other Studies

Petrography and mineralogical studies in support of mineralogical and geological interpretations have been completed by Ingemmet, Placer Dome, Surcusal Peruana, and CCC.

10.10 Exploration Potential

Cañariaco Norte is part of an extensive porphyry complex that comprises also Cañariaco Sur and a third target, Quebrada Verde. The three deposits cover a minimum length of 5 km and an average width of 2 km. The porphyry complex is covered entirely by the Cañariaco property. Geological mapping, geophysics and geochemical sampling by CCC, and by Surcusal Peruana in 1999 (including limited drilling), indicate a strong potential for additional porphyry copper–gold mineralization within the complex.

Locally, mineralization at Cañariaco Norte is still open at depth and at the central–west portion of the deposit.


The exploration programs completed to date are appropriate to the style of the deposits and prospects within the Project. The petrographic research work supports the genetic and affinity interpretations.




11.0 DRILLING

Drilling on the Project consists of 263 core holes (75,665.1 m), including geotechnical and hydrogeological drilling. A drill summary table by operator is included in Table 11-1; this table does not include the geotechnical or hydrogeological drilling. Drill hole locations for the Cañariaco Norte deposit are shown in Figure 11-1. Figure 11-2 shows the locations of the three drill holes completed at Cañariaco Sur and Quebrada Verde.

11.1 Drill Contractors

The Ingemmet program drill contractor was Craelius Tarratest Peruana S.A. who utilized three Long Year machines (two model 34s and one model 38). No information is available for the contractors for the Placer Dome and Surcusal Peruana drill programs.

The CCC drill programs initially used Pac Rim Drilling SA, a Peruvian subsidiary of Kluane Drilling Corp. (Kluane), who provided standard man-portable Hydrocore rigs. Kluane was acquired by Energold Drilling Corp. in September 2007, and the subsidiary was renamed Energold Drilling Peru S.A.C. (Energold); from September 2007, Energold has been the Project drill contractor.

11.2 Drill Methods

The Ingemmet program used NCD core size to 50 m, NXWL (60.7 mm core diameter) to 170 m and BXWL (48.4 mm core diameter) thereafter. Both 10 ft (3 m) and 5 ft (1.5 m) core barrels were used.

No information is available on the core size for the Placer Dome drilling; however, the Surcusal Peruana drill program used BQ diameter (36.5 mm).

The CCC programs initially used NTW (56 mm) diameter core. During the 2005 drill program, ground conditions and depth required a reduction to a BTW core diameter (42 mm) in drill holes 05-014 at 302.35 m and 05-019 at 268 m). In 2006 and 2007, all drill holes less than 325 m in depth were drilled with NTW diameter core; holes deeper than 325 m needed to be reduced to BTW at around 250 m to 300 m depth, to allow the machine to drill to the planned depth. The 2008 drill program employed HQ (63.5 mm diameter) or NQ (47.6 mm diameter) core sizes.




Table 11-1: Drill Summary Table Core Year Project Operator Deposit or Prospect Number Holes Metres 1971–1974 Ingemmet Cañariaco Norte 5 1,500.0 1994–1997 Placer Dome Cañariaco Norte 3 853.91 1999–2000 Surcusal Peruana Cañariaco Norte 3 555.7 Cañariaco Sur 3 473.0 Quebrada Verde 1 100.0 2004 Candente Cañariaco Norte 12 2,647.38 2005 Cañariaco Norte 24 7,388.87 2006 Cañariaco Norte 46 15,983.55 2007 Cañariaco Norte 123 31,357.2 2008 Cañariaco Norte 41 13,772.1 Cañariaco Sur 2 1,040.6 Total 263 75,672.31

Figure 11-1: Drill Hole Location Plan, Cañariaco Norte




Figure 11-2: Drill Hole Location Plan, Cañariaco Sur and Quebrada Verde

Note: Figure from Huangui et al., 2002. Billiton = Surcusal Peruana




For the CCC programs, core was transferred to wooden core boxes. Rock quality descriptions (RQD) measurements were performed at the drill site, then boxes were lidded, and transported by porters hired by CCC to the Project core processing facility where it was photographed, logged for geologic and geotechnical information.

No information is available on the core transportation used in the earlier drill programs by Ingemmet, Placer Dome, or Surcusal Peruana.

11.3 Geological Logging

During detailed core mark-up at the logging facility, CCC geologists perform first-pass geological logging of the core. This initial log identifies lithological boundaries, major structures, leaching depth and broad mineralization intervals. In tandem with the logging, the geologist notes where major sample breaks should be placed at lithological boundaries.

A second phase of detailed core logging is carried out immediately after cutting on the preserved split core at CCC’s core storage facility in Chiclayo, where each core is logged for lithology, structure, veining, alteration and mineralization. All logged descriptions are indexed to 2 m sample intervals, so that the mineralization, veining, structure, lithology and alteration affecting any sample are known.

Core from all other legacy drill campaigns has been re-logged by CCC.

11.4 Collar Surveys

CCC drill collar locations were picked up by a surveyor using a total station instrument. All legacy drill collars were picked up by survey in 2006, and tied into the Project grid.

11.5 Down-hole Surveys

For the 2004 drill program, down-hole surveys were undertaken using a Pajari analog survey tool with a timed locking compass and inclinometer. Readings were taken at the drill machine to confirm true vertical holes. Any spurious readings were discarded and a second survey was done at the same depth. The 2005 program used the same instrumentation and procedures as the 2004 program. However, a number of the 2005 program drill holes do not have down hole survey data due to a malfunction in the instrument.

All the drill holes from the 2006 and 2007 drill programs were downhole surveyed with a Sperry Sun single shot downhole survey tool supplied by the drilling company.




The 2008 program drill holes were down-hole surveyed for dip and azimuth with a Reflex EZ-Shot digital down-hole survey tool, at 50 m intervals.

Selected drill cores from the 2006 and 2008 drill programs were oriented for geotechnical purposes using an A.C.E. tool.

11.6 Recovery

Drill core recovery data from the legacy drill campaigns are not available. CCC staff have recorded core lengths and calculated core recoveries and RQD at the sites of the drill holes.

AMEC reviewed the core recovery data from drill holes C07-164 through to C08-244 and found a trend of decreasing copper grade with decreasing core recovery. The copper grades decrease from 0.3% to 0.2% at a core recovery of 80%, however only 2% of the data are affected. There is a low risk of a very low number of assays having a negatively-biased copper grade due to low core recoveries.

11.7 Cañariaco Norte Drilling

The central portion of the deposit has been delineated at a nominal, but not regular spacing, due to relief restrictions. In this area, the drill grid spacing is 75 m by 75 m with most of the core holes drilled vertically. On the deposit margins, drill spacing widens to as much as 100 m by 100 m.

Drill hole depths range from 14 m to 766 m, averaging 290 m. Drill hole orientations are typically vertical, but have also been drilled at angles ranging from 45˚ to 75˚.

The relationship between true widths, drill intercepts, lithologies and copper grades for drill hole intervals in selected drill holes are shown on the cross-sections included as Figure 9-1, Figure 9-2 and Figure 9-3 in Section 9.

Example drill intercepts for the deposits are summarized in Table 11-2, and are illustrative of nature of the mineralization. The example drill holes contain leached cap and sulphide intersections and areas of higher-grade in lower-grade intervals.




Table 11-2: Drill Hole Intercept Summary Table

Deposit Hole ID From (m)

To (m)

Drill Intercept Interval

(m) Copper Grade

(Cu%) Gold Grade

(g/t Au)

Silver Grade

(g/t Ag) Cañariaco Norte C07-200 76 300.25 224.25 0.79 0.140 2.58 Cañariaco Norte C05-020 12 357.95 345.95 0.49 0.045 2.46 Cañariaco Norte C08-213 27.25 46.95 19.7 0.96 0.142 2.19 Cañariaco Norte C08-213 52.4 98.0 45.6 1.01 0.145 3.01 Cañariaco Norte C06-074 52.4 98.0 237.0 0.605 0.111 1.99 Cañariaco Norte C07-144 70.6 271.0 200.4 0.812 0.129 2.48 Cañariaco Norte C06-076 34.9 363.0 328.1 0.911 0.169 2.20

Note: Depending on the dip of the drill hole, and the dip of the mineralization, drill intercept widths are typically greater than true widths.


In the opinion of the QP, the quantity and quality of the lithological, geotechnical, collar and downhole survey data collected in the exploration and infill drill programs are sufficient to support Mineral Resource estimation as follows:

• Core logging meets industry standards for gold, silver, and copper exploration

• Collar surveys have been performed using industry-standard instrumentation

• Down-hole surveys were performed using industry-standard instrumentation

• Recovery data from core drill programs are acceptable to allow reliable sample data for mineral resource estimation.

• Geotechnical logging of drill core meets industry standards for planned open pit operations.

• Depending on the dip of the drill hole, and the dip of the mineralization, drill intercept widths are typically greater than true widths.

• Drill orientations are generally appropriate for the mineralization style, and have been drilled at orientations that are optimal for the orientation of mineralization for the bulk of the deposit area.

• Drill orientations are shown in the example cross-sections (Figures 9-1 to 9-3), and can be seen to appropriately test the mineralization. The sections display typical drill hole orientations for the deposits, show summary assay values using colour ranges for assay intervals that include areas of non-mineralized and very low grade mineralization, and outline areas where higher-grade intercepts can be identified within lower-grade sections. The sections confirm that sampling is representative of the gold, silver, and copper grades in the deposits, reflecting areas of higher and lower grades.




• Drill hole intercepts as summarized in Table 11-2 appropriately reflect the nature of the gold, silver, and copper mineralization.

• No factors were identified with the data collection from the drill programs that could affect Mineral Resource estimation.




12.0 SAMPLING METHOD AND APPROACH

12.1 Geochemical Sampling

There is no information available on the sampling methods for the Ingemmet, Placer Dome or Surcusal Peruana geochemical sampling programs. Geochemical samples taken by CCC were typically 2–5 kg.

12.2 Pit and Trench Sampling

There is no information available on the sampling methods for the Ingemmet and Placer Dome geochemical sampling programs.

12.3 Core Sampling

Drill core generated by Ingemmet and Placer Dome was halved; there is no information as to the typical sample intervals. The Surcusal Peruana drill core was halved and sampled on 2 m intervals.

CCC drill core is halved using a circular rock saw. Samples are 2 m in length unless a geological contact is present within the sample interval. In those instances, the sample interval is terminated at the contact. The subsequent sample interval terminates at the next meter depth mark that allows a 1.50 m minimum sample length.

12.4 Quality Assurance and Quality Control

The quality assurance and quality control (QA/QC) programs for the Project are discussed in Section 13.

12.5 Density/Specific Gravity

A total of 9,424 bulk density readings were taken by CCC personnel during core logging using a weight in air and weight in water. Samples were taken every 10 m to 20 m to include all rock and alteration types. Data were recorded for drill hole ID, depth, rock type, alteration, weight in air and weight submerged in water.

As a part of the update to the 2008 mineral resource estimate, 550 bulk density determinations were collected by CCC personnel from drill core intervals and submitted to ALS Chemex (Lima) for density determinations. The determinations were performed using a wax-coated immersion technique (specialty assay procedure




OA-GRA08a). AMEC assigned constant specific gravity values to each lithological rock type (Table 12-1).

Table 12-1: Bulk Density Values used in the 2010 Mineral Resource Estimate

Bulk Density Oxide

(g/cm3) Sulphide (g/cm3)

VC 2.13 2.51 CQFP 2.33 2.50 BFP 2.19 2.48 CQP 2.17 2.50 FP 2.03 2.56 IBXH 2.17 2.50 TBX 2.17 2.45 IBXP 2.17 2.50


A description of the geology and mineralization of the deposit, which includes lithologies, geological controls and widths of mineralized zones, is included in Section 7 and Section 9.

A description of the sampling methods, location, type, nature, and spacing of samples collected on the Project is included in Section 10 and Section 12.

A description of the drilling programs, including sampling and recovery factors, are included in Section 11 and Section 12. All collection, splitting, and bagging of core samples were carried out by company personnel, with the company and personnel varying depending on the date of the drill program. No factors were identified with the drilling programs that could affect the reliability of the sample data used for Mineral Resource estimation.

Figures 11-1 and 11-2 in Section 11, which show drill hole collar locations, indicate that the sizes of the sampled areas are representative of the distribution and orientation of the mineralization.

Figures 9-1 to 9-3 in Section 9 show approximate drill hole collar traces in relation to the orientation of the mineralization. The figures also show drill hole assay intervals include areas of non-mineralized and very low grade mineralization, and confirm that sampling is representative of the copper, gold, and silver grades in the deposit, reflecting areas of higher and lower grades.




Data validation of the drilling and sampling program is discussed in Section 14, and includes review of database audit results.

In the opinion of the QPs, the sampling methods are acceptable, meet industry-standard practice, and are adequate for Mineral Resource estimation purposes, based on the following:

• Data are collected following industry standard sampling protocols;

• Sample collection and handling of core was undertaken in accordance with industry standard practices, with procedures to limit potential sample losses and sampling biases;

• Sample intervals in core, comprising average 2 m intervals, are considered to be adequately representative of the mineralization. Not all drill material may be sampled depending on location and alteration.

• Bulk density determination procedures are consistent with industry-standard procedures;

• There are sufficient acceptable bulk density determinations to support the bulk density values utilized in tonnage interpolations.




13.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY

From Project inception to date, Project staff employed by CCC and predecessor companies were responsible for the following:

• Sample collection • Core splitting • Preparation of samples for analytical laboratory submission • Sample storage • Sample security.

13.1 Analytical Laboratories

Several primary assay laboratories have been used for routine analyses over the Project history.

Ingemmet used the Plenge Laboratory in Lima and the Ingemmet internal laboratory, also located in Lima. No information is available as to accreditation of the laboratories at the time sampling was performed.

Placer Dome utilized the SGS-XRAL (SGS) laboratory in Lima, which was independent of Placer Dome. No information is available as to accreditation of the laboratory at the time sampling was performed.

For the Surcusal Peruana sampling, sample preparation was undertaken by ALS Chemex in Trujillo, and primary analysis by ALS Chemex in Lima. The umpire laboratory was SGS. Both laboratories were independent of Surcusal Peruana. No information is available as to accreditation of the laboratories at the time sampling was performed.

Activation-Skyline Laboratories (Actlabs) in Lima, Peru performed all of the sample preparation and the majority of the analyses for the CCC programs. Actlabs is independent of CCC, and holds ISO:9000 accreditation for the Peruvian laboratory. Inductively-coupled plasma (ICP) analyses were performed by the Ancaster, Canada, Actlabs laboratory, which has Standards Council of Canada (SCC) accreditation for International Standards Organization (ISO)17025.

Some analyses for the re-analysis of pre-2008 core samples for gold and ICP were undertaken by ALS Chemex in Lima. ALS Chemex is independent of CCC, and holds ISO:9000 accreditation for the Peruvian laboratory.




ACME Laboratories (ACME), Lima were used as a check laboratory for pulp analyses. ACME is independent of CCC, and holds ISO:9000 accreditation for the Peruvian laboratory.

13.2 Sample Preparation and Analysis

Limited information is available on the sample preparation and analysis methods for Ingemmet or Placer Dome. Ingemmet samples were analyzed for Cu and Mo, and more rarely Au and Ag, using a colorimetric analytical method.

Surcusal Peruana samples were 200g splits of a 1 kg, 200 mesh homogenized sample. A split from each sample pulp was assayed for Au (fire assay with atomic absorption finish, 10 pbb detection limit) and Cu, Pb, Zn, Mo and As (multi-acid, total digest), with an atomic absorption (AA) finish for each element. SGS completed check assays on a split of one in 20 pulps using the same analytical procedures as the initial analysis performed by ALS Chemex.

13.2.1 CCC

Sample Preparation

Sample preparation undertaken on the CCC samples comprised drying, then crushing using a jaw crusher to >70% less than 10 mesh. The sample is thoroughly blended using a riffle splitter. A sub-split is taken, which is pulverized to >95% less than 150 mesh, and this pulp is submitted for analysis.

Copper

Each sample was subject to total copper and sequential copper leaching analysis which returns results for acid-soluble, cyanide-soluble and residual copper grades. Total copper analysis was performed using a three-acid digest and AA finish (laboratory method ME-3 or three-acid digestion).

The sequential leach analysis consisted of:

• An initial leach step, where samples were dissolved in sulphuric acid, and the copper grade determined by AA to give the acid-soluble copper value.

• A secondary leach step, where samples were dissolved in sodium cyanide and the copper grade determined by AA to give the cyanide-soluble copper value.




• A third leach step, where the samples were dissolved using a three-acid digest and the copper grade determined by AA finish to give the residual copper value.

For any given sample, copper grades were obtained by adding the three parts of the sequential copper analysis. Results are compared to the copper grade reported in the total copper analysis, and if the sum of the sequential leach grades has a >0.03% difference to the total copper grade, the analysis is repeated for both parts of the process.

Gold

The sample for gold analysis is taken from the remaining pulps after copper analysis. The ±250 g pulps are homogenized and a 30 g split is weighed out for fire assay fusion, cupelled to obtain a bead, and digested with aqua regia, followed by an AA finish, with a detection limit of 5 ppb Au (Actlabs code EF1).

Only a portion of the pre-2008 drilling originally had gold assays. These were analysed using a fire assay (FA) methodology with an atomic absorption finish (FA-AA) on a 30 g sample. The remainder of samples were not systematically analysed for gold until a large analysis in-fill campaign was conducted in March 2008. Due to the large volume of samples, analyses were split between Actlabs and ALS Chemex. The FA-AA method on a 30 g sample was used by both laboratories.

For the 2008 drill campaign, the FA-AA method on a 30 g sample has been used, with all analyses performed by Actlabs.

Inductively-coupled Plasma Multi-element

The analytical method utilized at Actlabs was a 36-element inductively-coupled plasma optical emission spectrometry (ICP-OES) method following aqua regia digestion (laboratory code 1E3). ALS Chemex performed a 33-element ICP atomic emission spectroscopy (AES) method after four-acid digestion (laboratory code ME-ICP61).

13.3 Quality Assurance/Quality Control Programs

There is no information on any quality assurance/quality control (QA/QC) programs for Ingemmet and Placer Dome.

Surcusal Peruana used blanks (crushed quartz every 20 samples), standard reference materials (SRM: one in every 20 samples) and check assays (one in every 20 samples). Surcusal Peruana’s protocol was that a blank sample was not to be




submitted adjacent to an SRM sample, but could not be any more than 10 samples from an SRM. Chain-of-custody and sample preparation protocols were also part of Surcusal Peruana’s QA/QC program at Cañariaco.

13.3.1 CCC

Duplicates

No field duplicates were utilized in the 2004 drilling program, but were included from drill hole C05-013 of the 2005 drill program. Field duplicates typically comprised quarter drill core.

Coarse reject duplicates were run on drill holes C04-007 and C05-023.

Pulp duplicates were taken at Actlabs every 10th sample.

CCC has consistently sent 5% of pulps prepared and analyzed by Actlabs to independent laboratories for check assays on total copper analyses. Drill holes C04-001 to C06-082 had 5% of the samples sent to ACME Laboratories, Lima for analysis by a four-acid digestion of a 0.25 g pulp split. Drill holes C07-083 to C07-202 had 5% of the samples sent to SGS laboratories, Lima for analysis using the same analytical procedures as the original assay work (three-acid digestion on a 0.25 g split). All independent checks included standards, blanks and duplicates. Pulp duplicates were also made during all independent laboratory checks at ACME and SGS.

During 2010, CCC sent a suite of 530 pulp reject samples to the SGS laboratory in Lima for check analyses on silver ICP analyses. The samples were analysed by a four-acid digest on a 0.3 g pulp split (method ICP40B). The samples were randomly selected from drill holes throughout the holes drilled by CCC. All independent checks included standards, blanks and duplicates.

Blanks

Field blank material is obtained from a barren outcrop of volcanic host rock situated outside of the alteration halo at Cañariaco. Blanks are submitted at a frequency of one in 30 samples, so that each laboratory batch of 80 will always have at least two blanks. Field blank fail limits are set at 0.05% copper or five times the detection limit of 0.01%.




Standard Reference Materials

CCC has used six commercially available SRMs, purchased from CDN Laboratories in Vancouver, Canada. SRMs are submitted at a frequency of one in 30 samples, so that each laboratory batch of 80 will always have at least two SRMs. An additional two SRMs were purchased from Geostats Pty. in Australia specifically for use in the ICP check assays conducted during 2010. The selected SRMs have low-grade silver contents which are close to the average silver grade of the Cañariaco Norte deposit.

During 2006, SGS was retained to prepare two SRMs from unweathered outcrops from the Cañariaco property, for sequential leach analysis.

Where the copper value of the SRM was outside the acceptable value specified for the standard by CDN Laboratories (two standard deviations from the mean), they were failed. Re-analysis of failed copper standards was not carried out at time of drilling. This was mainly due to the fact that most batches had two, or sometimes three, standards in them. If one standard failed but the other did not, the batch was considered to be acceptable, particularly if the other QA/QC samples (blanks and duplicates) did not fail.

Re-sampling

An extensive re-sampling program of historical pre-2004 drill hole data has been conducted by CCC. Nine out of eleven historical holes were re-sampled. Only holes C73-001 and C73-005 holes were not re-sampled. C73-001 was twinned by C04-001 and C73-005 did not have significant copper grades.

Sections of the old core were quartered and sent to Actlabs, Lima, for processing. Due to the state of the legacy drill core after several moves, composite samples of up to 20 m were made within mineralization types and between fixed blocks to ensure proper metreage measurements.

A comparison between the original and the CCC assays shows an acceptable correlation between the datasets. Based on this, all of the historical data were added to the final database, and the CCC assays from the re-sampling program were excluded due to the composite lengths.

Twin Holes

Three pairs of twinned holes were drilled by CCC to verify grade uniformity at short distances. Two of the twin sets (holes C04-007 with C07-104 and C04-023 with C07-106) show similar average grades over the same depth intervals. The third twin




set (C04-005 with C07-146) shows a significantly higher average grades in the original drill hole from 2004 than in its twin drilled in 2007. This may indicate a vertically oriented mineralized vein set intercepted in one drill hole and not the other.

13.4 Databases

All date in the field is recorded in written form in field books, log books, sample sheets, logging forms or shipping forms. Various phases of record keeping are repeated in the subsequent step to confirm recorded values or numbers.

All field data is entered into Excel tables either in the Cañariaco camp or at the CCC Lima office. Errors in data entry picked up during the verification stage can be confirmed and corrected from filed data.

Data from third parties such as laboratories or survey contractors are generally supplied in digital and printed form. These records are printed out and kept in binders for reference during data verification.

13.5 Sample Security

Sample security at the Cañariaco Project during the CCC drilling programs relied upon the remote nature of the site, the fact that the samples were always attended or locked at the sample dispatch facility. Sample collection and transportation have always been undertaken by company or laboratory personnel using company vehicles.

Drill samples were picked up at site by a freight firm using a dedicated vehicle, and transported to the sample preparation facility. Pulps were transported by laboratory personnel to the appropriate analytical facility.

Chain of custody procedures consisted of filling out sample submittal forms that were sent to the laboratory with sample shipments to make certain that all samples were received by the laboratory.

13.6 Sample Storage

Prior to the construction of an access road into camp, all core was stored on racks within secure storage facilities. In October 2007, all existing core was moved to a secure core storage facility located in Chiclayo, close to CCC’s regional offices and thereafter core boxes were transported directly to the new core storage facility for detailed logging and permanent storage.





The QPs are of the opinion that the quality of the gold, copper, and silver analytical data are sufficiently reliable (also see discussion in Section 14) to support Mineral Resource estimation and that sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry standards as follows:

• Geochemical sampling covered sufficient area and was adequately spaced to generate first-order geochemical anomalies, and thus was representative of first-pass exploration sampling

• Drill sampling has been adequately spaced to first define, then infill, gold and copper anomalies to produce prospect-scale and deposit-scale drill data. Drill hole spacing varies with depth. Drill hole spacing in the core of the deposit is approximately 50 m. Drill hole spacing increases with depth as the number of holes decrease and holes deviate apart, and is more widely-spaced on the edges of the deposit

• Sample preparation for samples that support Mineral Resource estimation has followed a similar procedure since 2004. The preparation procedure is in line with industry-standard methods for copper–gold–silver deposits

• Core drill programs were analysed by independent laboratories using industry-standard methods for copper, gold, and silver analysis

• There is limited information available on the QA/QC employed for the earlier drill programs; however, there are twin drill holes that confirm the grades and lithologies, and the core from the drill programs has been re-assayed, so that the data can be accepted for use in Mineral Resource estimation.

• Typically, CCC drill programs included insertion of blank, duplicate and SRM samples. The QA/QC program results do not indicate any problems with the analytical programs, therefore the copper, gold, and silver analyses from the core drilling are suitable for inclusion in Mineral Resource estimation

• Data that were collected were subject to validation, using in-built program triggers that automatically checked data on upload to the database

• Verification is performed on all digitally-collected data on upload to the main database, and includes checks on surveys, collar co-ordinates, lithology data, and assay data. The checks are appropriate, and consistent with industry standards

• Sample security has relied upon the fact that the samples were always attended or locked in the on-site sample preparation facility.




• Chain-of-custody procedures consist of filling out sample submittal forms that are sent to the laboratory with sample shipments to make certain that all samples are received by the laboratory

• Current sample storage procedures and storage areas are consistent with industry standards.




14.0 DATA VERIFICATION

A number of data verification programs and audits have been performed over the Project history, primarily in support of technical reports.

14.1 Currie, 2004

During a site visit in support of completion of a technical report, Currie (2004) traversed the entire Cañariaco Norte zone, examining numerous outcrops. Four samples were taken from outcrop material and trenches. Analyses indicated that the assays from the outcrops were consistent with the grades seen in drill holes.

14.2 MineFill, 2007

MineFill performed detailed data verification for all available data from 1973 to 2006; only verified assay information was used in the estimation of mineral resources. Assay values were deemed verified when the original signed assay certificate or photocopy was present and the database reflected the assay certificate values accordingly. The sample numbers and assay values on the certificates were called out by an individual, as another individual located the corresponding sample numbers within the database and verified the assay values. The data were marked as verified, corrected or unverified, accordingly. Of the five holes drilled by Ingemmet in 1973, only three were verified. The remaining 88 core holes in the 2007 database were completely verified and corrected for use in mineral resource estimates.

14.3 SRK, 2008

SRK undertook the following checks:

• Detailed verification of assays using signed assay certificates • Assay verification from electronic laboratory files • Verification of down-hole survey data • Verification of drill hole positions in field • Comparison of Cu grades from re-sampling of historical drill holes to original data • Verification of Cu assay data from twinned drill holes • Comparison of Cu assays from different analytical procedures • Comparison of Cu assays from vertical versus inclined holes • Collection and independent analysis of check assay samples.

No errors or omissions were noted by SRK with the reviewed data that could impact mineral resource estimation.




SRK also selected a suite of 21 core samples for independent analysis at ALS Chemex, Vancouver by four-acid digest and ICP-AES. Samples were chosen on the basis of their Cu grade (determined by CCC), lithology, mineralization type (presence of leachable Cu) and age of drill hole, in an effort to reflect the variability in the deposit and the consistency of analytical results over time. Samples comprised bags of pieces of quartered core samples taken over a 2 m interval in an attempt to ensure that the sampled interval co-incided with a CCC sample interval. However, the very fractured, rubbly nature of the recovery of several intervals meant that the samples reflect a sample of gravel-sized core fragments, rather than split core.

SRK’s comparisons of results with the original CCC assays yielded relatively good agreement on the whole. Graphic evaluation of the data indicated that approximately 65% and 90% of paired data fell below the 10% and 20% absolute relative difference lines. Significantly, there was no strong bias within the data, with points falling above and below the parity line. The data did not, in SRK’s opinion show an obvious correlation between the relative difference and lithology, mineralization, grade or time of assay.

SRK also concluded that the reasonable correlation between the results from SRK checks using a four-acid digestion technique with ICP finish and those obtained by CCC using a three-acid leach and AA finish further confirmed that the three-acid leach technique was suitable for the Cañariaco mineralization.

14.4 AMEC, 2010

AMEC reviewed 1,930 copper, gold, silver and molybdenum assays, or 5.6% of the analytical data in the database as verification on the data quality. No errors were noted and the analytical data were considered suitable to support mineral resource estimation.


The process of data verification for the Project has been performed by external consultancies, primarily in support of technical reports.

The QPs, who rely upon this work, have reviewed the appropriate reports, and are of the opinion that the data verification programs undertaken on the data collected from the Project adequately support the geological interpretations, the analytical and database quality, and therefore support the use of the data in Mineral Resource estimation:




• No significant sample biases were identified from the QA/QC programs undertaken

• Twin and infill drilling in areas where drill core data from predecessor companies was available indicated that the legacy data were sufficiently in accordance with the data generated by the check programs that the original historic assay values could be used

• Sample data collected adequately reflect deposit dimensions, true widths of mineralization, and the style of the deposit

• External reviews of the database have been undertaken in support of technical reports, producing independent assessments of the database quality. No significant problems with the database, sampling protocols, flowsheets, check analysis program, or data storage were noted

• Drill data are typically verified prior to Mineral Resource estimation by running a software program check.




15.0 ADJACENT PROPERTIES

There are no adjacent properties at a similar stage of development as the Project.




16.0 MINERAL PROCESSING AND METALLURGICAL TESTING

16.1 Metallurgical Testwork

Over the project history, a number of metallurgical testwork campaigns have been undertaken. These are summarized in Table 16-1.

Table 16-1: Metallurgical Testwork Summary Table

Year Laboratory Testwork Performed 1999 Geomet S.A., Santiago, Chile Sequential leach tests 2004 Kappes, Cassiday &

Associates, Reno, Nevada Column leach tests

2006 SGS Metallurgical Laboratory, Santiago, Chile

Leachability tests


QEMSCAN examination Comminution and variability comminution tests Effects of grind sizes Effects of collectors and pH Effect of sulphidization Cleaner flotation tests Locked cycle tests


Effects of grind sizes Effects of collector, frother type and dosage, and pulp pH Effects of cleaning Flotation tests Concentrate generation Mineralogical analysis

2010 Outotec, Antofagasta, Chile Tailings thickening tests 2010 Outotec, Sweden Proof-of-concept roasting testwork

Initial testwork was focused on leaching; however, additional drill data indicated that the mineralization was primarily sulphidic and flotation/comminution tests were subsequently performed.

16.1.1 Leach Tests

In 1999, Surcusal Peruana sent 12 samples from two intervals of core to Geomet S.A. in Santiago, Chile for diagnostic leach tests. Four different test phases, using four different solvents were trialed. Huanqui et al (2002) concluded that the potential leachable copper content of the finely-ground (minus 200 mesh) samples averaged 86.03% copper.




In 2004, Kappes, Cassiday & Associates (KCA) of Reno, Nevada, performed column leach tests on two bulk core samples. The oxide column test sample returned 92% recovery after 93 days. The more sulphidic sample returned a recovery of 52% after 93 days.

During 2006, a total of 32 core samples were sent to SGS Metallurgical Laboratory in Santiago, Chile for testwork to assess the leachability of copper from high cyanide-soluble copper areas in the deposit. Results were similar to those obtained by KCA, and confirmed that not all the soluble copper minerals in the mineralized material were completely leachable. Results were variable and indicated potential recoveries of 60% to 70% of the contained leachable copper.

16.1.2 Mineralogy

A Qemscan examination of metallurgical samples during 2010 testwork indicated the predominant minerals and mineral associations at Cañariaco Norte samples are:

• Copper mineral species: chalcopyrite, covellite, chalcocite, bornite, enargite, tennantite, and tetrahedrite; primarily associated with pyrite and other sulphide gangue minerals. Pyrite present as liberated pyrite or in association with other sulphide minerals.

• Non-metallic gangue minerals: quartz, micas, biotite, chlorite, and clay group minerals.

Particle mineralogical analysis of the three 2010 rougher concentrate samples indicated:

• Chalcopyrite is the main ore mineral with values between 6.06 and 9.30 wt%. The main gangue in the samples is quartz, pyrite, and sericite/muscovite. Other gangue includes plagioclase/albite, chlorite, clays, Ti oxides and biotite. The fine fraction (-400 mesh) has low amounts of pyrites and is primarily phyllosilicates.

• In the high-chalcopyrite and low-arsenic rougher concentrates, the copper occurs mainly as chalcopyrite (70.7 and 71.7% respectively). The remaining copper is from bornite, enargite, covellite, chalcocite/digenite and trace tetrahedrite.

• In the high arsenic secondary copper mineral rougher concentrate, the copper occurs as chalcopyrite (49.1%), bornite (19.4%), enargite (12.3%), covellite (12.9%), chalcocite/digenite and trace tetrahedrite.

• Copper sulphide liberation increases with the decrease in particle size.

• Copper sulphides are typically associated with pyrite and silicates/phyllosilicates.




16.1.3 Comminution Tests

Comminution tests were performed in 2008 at the SGS Metallurgical Laboratory (SGS) on four specific copper mineralization composite samples to emphasize chalcopyrite-rich, arsenic-rich, and secondary copper mineral-rich materials. In addition there was a comminution variability campaign on selected spatial samples (48 in total) and the testwork on these involved the following test types:

• Crushing work index (CWI), • SAG mill comminution (JKSMC) • Drop weight tests (JKDWT) for ten samples • Specific gravity (SG) • Bond Rod mill work index (RWI) • Bond Ball mill work index (BWI) • Abrasion Index (AI).

Results were:

• Cañariaco variability crushing work indices were in the range of 4 kWh/t to 12 kWh/t, which are typical of porphyry copper deposits but are significantly lower than those reported in the original PA. The 80th percentile of the CWi was found to be 13 kWh/t.

• The variability testwork indicates an average bond rod mill work index (RWi) of just over 10 kWh/t. This, in conjunction with the BWi of just under 10 kWh/t, indicates that the ore is a soft to moderately soft type

• Ball mill work indices in the range of 9 to 12 kWh/t confirm the results of the process development composites by showing a soft to moderately soft ore type

• The average abrasion index was less than 0.1, which may indicate that the selected variability composites correspond to a non-abrasive mineralized material. At the 80th percentile, the abrasion index is around 0.12, which is the basis of the steel values used for the operating cost.

16.1.4 Flotation Tests

2006

Initial flotation tests were performed in 2006 by SGS on three lithology composites, breccia-rich, chalcopyrite-rich, and copper secondary mineral-rich. Results discussed are summarized from Bonson et al. (2008):




• Several rougher flotation tests were performed at different particle sizes (K80 of 75, 106, 150, 212 and 250 μm) and at a pH of 10. A P80 of 150 μm was recommended for the later stages of testwork.

• Rougher flotation tests were carried out to determine the best reagents to obtain a high copper recovery. The best flotation results were obtained with the collector AP-3894 and methyl isobutyl carbinol (MIBC) frother.

• Flotation tests conducted at pH values of 7.2, 11 and 12 indicated that a pH of 10 is optimum for the rougher flotation stage. This pH gave acceptable results for all of the composites tested.

• Flotation performance (recovery and grade) was not improved by the addition of a sulphidizing reagent.

• Concentrate grade is sensitive to cleaning flotation pH and regrind size. Finer regrind improved concentrate grade for all the composites tested. Results indicated performance at a P80 of 37 μm (80% passing 400 mesh) achieved satisfactory concentrate grade for each type of mineralized material.

• A recovery of 89% Cu with a concentrate grade of 25.5% Cu was achieved for the chalcopyrite-rich composite. For the breccia composite, the final concentrate grade was 26.6% Cu with 81% recovery. For the composite containing higher levels of secondary minerals, a recovery of 84.5% Cu with a final concentrate grade of 26.5% Cu was achieved.

2008

The same material used to provide samples for the variability grinding work was used to create four composites which were used as the basis of flotation testwork conducted at SGS Metallurgical Laboratory in 2008. Results discussed are summarized from Bonson et al. (2008).

The effect of grind size (P80 at 75 μm, 100 μm, 150 μm, 200 μm and 250 μm) on the flotation performance using rougher kinetic tests was evaluated; although finer grinding improved the copper recovery, a primary grind of P80 125 μm was deemed as optimum.

Tests were completed out at grind size of 125 μm and pH of 10 with collector types of AP-3894, AP-3330, PAX, AP-3477, AP-404, AP-3926 and AP-3302 and with frother (MIBC) addition. Higher copper selectivity was observed for all composites using collectors AP-3330, and AP-3477.




Frother screening tests were carried out at grind size of 125 μm and pH of 10 with frother types of MIBC, D-250, H-76 and TEB and with addition of collectors AP-3477 (25 g/t) and AP-3330 (25 g/t). The results indicate that copper recovery and grade benefit most from the use of AP-3330/MIBC.

Several rougher kinetic tests were conducted at natural pH, pH of 9, 10, 11 and 12. Tests were performed at a grind size of 125 μm with addition of collector AP-3330 (25 g/t) and frother MIBC (20 g/t). At pH 9 and 10, good performance was obtained (recovery and grade) with particularly higher concentrate copper grades occurring at pH 10 for all composites. A noticeable decrease in copper recovery occurred at natural pH for the arsenic-rich and copper secondary mineral-rich samples.

Sulphidization tests were completed at a grind size of 125 μm, pH of 10 with addition of collector AP-3330 (25 g/t) and frother MIBC (10 g/t). The effect of the reagent NaHS addition (50 g/t and 100 g/t) on the mineral response to flotation was evaluated. The mixed, chalcopyrite-rich and copper secondary mineral-rich composites showed better copper recoveries without the addition of NaHS. The arsenic-rich sample did show slightly improved recovery with NaHS addition.

Regrinds were evaluated at times of 4 min, 8 min, and 12 min rather than specific P80 sizes to facilitate laboratory work with the small sample mass available. The chalcopyrite-rich sample required 4 min of regrinding to achieve optimal grade-recovery relationship in batch tests. Between 8 min and 12 min was required for the other three composites. Locked cycle tests and detailed mineralogy were recommended in order to further define the regrind requirements for each composite. This corresponds to a P80 of 50 μm for chalcopyrite-rich and arsenic-rich mineralization, and a P80 of 35 μm for secondary copper mineral-rich and mixed sulphide type materials.

Several cleaner flotation tests were conducted at pH 11, 11.5 and 12. Tests were carried out at a grind size of 125 μm with addition of AP-3330 (25 g/t) and frother MIBC (20 g/t). A pH of 11.5 achieved satisfactory results for all composites tested.

2010

In 2010, SGS performed additional testwork on three selected composites, a low-arsenic composite, a copper secondary mineral composite, and a composite containing high levels of chalcopyrite and arsenic. These composites were derived from the same inventory of material which supplied material for the 2008 composites.

Primary flotation was conducted with six stages (1, 2, 4, 8, 10, 12 and 18 min.) of roughing. Collector and frother type and dosage, and pulp pH were varied to evaluate




their effect on mineral flotation performance. The duration of each flotation test was 18 minutes.

The effect of grind size (P80 at 120, 150, 175, 200, and 250 μm) on the flotation performance of the three composites (at the conditions developed in 2008) was evaluated. A slight decrease in recovery was noted when going to a coarser grind with all three composites. However, the relative impact of going from 150 μm to 200 μm was fairly minor for all composite types. The results were sufficiently encouraging to develop a comparison model which utilized preliminary capital and operating costs at the various sizes versus the revenue developed. For all the mineralization types, the increase in throughput and decrease in overall cost outweighed any recovery loss and the required cost increases in regrinding. The results of the comparison model for all three cases indicated that it was beneficial to move from a 150 μm to a 200 μm primary grind as it allowed a greater amount of metal to be recovered per unit cost of expenditure.

Collector tests were performed to assess increases in recovery of both copper and gold into the concentrate. Rougher tests indicated that the best overall recovery performance for copper and gold was achieved by a 75% A-3894 and 25% PAX combination.

Cleaner tests showed that cleaning requires a high pH level of approximately 12 in order to produce satisfactory copper concentrate grades. The mixed reagent did not perform as well as the sole use of A3894. Additional mineralogical and size analysis is underway to further investigate the cause of the cleaning seen with the mixed reagent suite. Although there will be a compensating effect in the impurity roasting step, which will see an uplift in final copper concentrate grade suitable for sale, AMEC recommends that more material be obtained for a series of locked-cycle tests to examine this aspect in more detail.

Flotation testwork was performed to develop the tailings flotation flowsheet and provide sample suitable for acid-base account testing.

16.1.5 Concentrate Generation

Additional testwork was performed on all three 2010 composites to produce concentrate feed for proof-of-concept roasting testwork by Outotec. Testwork utilized a primary grind of 150 μm and a regrind of approximately 35 μm. The collector used was 25 g/t of A3894.

During the production of the concentrates, thickening tests performed on the samples showed that:




• Required underflow slurry densities could be achieved for all the flocculant addition rates at the three loading levels, 0.30 t/m2h, 0.50 t/m2h, and 0.70 t/m2h.

• Required overflow slurry clarities be achieved for most of the Magnafloc 1011 flocculant addition rates at the three loading levels, 0.30 t/m2h, 0.50 t/m2h, and 0.70 t/m2h.

All concentrates show reasonable settling performance.

Concentrate roasting tests were performed using a rotating kiln at a fixed temperature. The objective was to reduce the As and Sb in the concentrate to levels that would have low or no penalty elements and thus enhance the marketability of the product. Each of the three concentrates was tested at four different temperatures, 600°C, 650°C, 700°C and 725°C with a residence time of 30 minutes. A stable reducing environment was achieved by adding a flow of SO2 gas. The produced calcines were analyzed both for sintering tendencies and elemental composition.

The roasting tests showed that the expected arsenic content in the head grade of the calcine was similar to the low-arsenic composite, which gave a residual As result under 0.1%. Composites with higher arsenic levels also produced calcines with residual As levels of between 0.1% to 0.3%, unless there was calcium present either in the gangue or from the reagent systems.

The low-arsenic composite calcine showed the best performance at 700°C, producing a concentrate that would not incur a smelter penalty. The secondary copper mineral composite calcine also showed good performance at 700°C, and typically would not incur a smelter penalty. The results for the high-chalcopyrite and arsenic composite calcine at 700°C, showed a level of arsenic that would incur a penalty; however, Outotec indicated that this may due to the presence of excess lime used for pH control in the cleaning circuit.

The levels of antimony were found to reduce by 50% to 60% to a level where this element would be below penalty levels.

The silver and gold content in the calcine was not analyzed; however, these elements are expected behave similarly to copper. The average increase in concentrate grade was 6% for copper, and both silver and gold should follow the same trend.

The roasting tests did not show any problems related to sintering. All concentrates had a P80 of 50 µm or less, with approximately 20% of the material smaller than 10 µm. This suggests that 5% of the calcine will report to the scrubbing circuit for capture at that point.




16.1.6 Testwork Results

2008

A standard copper flotation flowsheet was recommended in 2008, incorporating crushing, grinding, rougher flotation, rougher concentrate regrinding, and two stage cleaner flotation to produce a copper concentrate. Under the recommended scenario, the first cleaner tail would be scavenged in a mechanical cell and the cleaner scavenger concentrate sent back to the regrind mill. The rougher tails and the cleaner scavenger tails would be discharged to a tailings containment facility. The cleaner concentrate would be dewatered and shipped offshore to a smelter for refining.

A recovery of 90% Cu was suggested for PA-level work, based on a P80 of 125 μm and a rougher flotation mass pull of 20%. A concentrate grade of 27% Cu was projected. Recoveries of 55% for gold and 65% for silver were recommended. No molybdenum recovery was proposed, due to the very low tenor of the Mo mineralization. It was noted at the time that arsenic and antimony in the concentrates could incur smelter penalties.

2010

In 2010, it was recognized that there was an opportunity to optimize the metallurgy of the deposit and to use another approach to deal with the levels of arsenic present in the concentrate. This second phase of work for Cañariaco Norte began in 2010 with the initial goal being continuation of the definition of mineralization characteristics sufficient to support a pre-feasibility study.

The early work in this program consisted of testing for the optimum rougher feed size, regrind feed size, rougher and cleaner reagent additions. Products from this testwork were further analyzed for mineralogy, chemistry and settling quality. Results from this testwork indicated that it was possible to utilize a coarser grind and still maintain recoveries through the employment of a dual collector system. Later work on the test material concentrated on establishing that the improvements were robust and the solution proposed for dealing with the penalty elements would work. This latter testwork, the application of partial roasting, has been to a proof-of-concept level and development work is recommended to define design parameters for feasibility-level design.

At the effective date of this Report, the principal elements of the final flowsheet have been established with indicative recoveries of 89.7% Cu, 55% Au, and 50% Ag, >26% Cu concentrate delivered to the roaster, >30% Cu concentrate delivered to the




smelter, and arsenic as a penalty element controlled to a level where no impact on concentrate marketability is anticipated over the life-of-mine.

16.2 Recoveries

Copper recoveries are expected to vary with feed grade ranging from 93.1% at 0.6% Cu and 58.8% at 0.1% Cu. The metallurgical response of gold and silver has not been assessed to the same degree as copper; however, based on the testwork completed to date, gold and silver recoveries are projected to be 55% and 50% respectively.

Given that the results for the three composites were weighted equally, further definition of the distribution of these sample types and their representation in the LOM plan is recommended. This can be achieved through a variability flotation testwork program, which is currently planned for the next phase of work.

16.3 Proposed Process Route

The process design is for a concentrator with a nominal processing capacity of 95,000 t/d. Where data were not available at the time of flowsheet development, AMEC developed criteria for the sizing and selection of equipment based on comparable industry applications, benchmarking, and the use of modern modelling and simulation techniques.

16.3.1 Process Description

The mineral processing and the roasting/acid plants are based on conventional technology and industry proven equipment. Run-of-mine (ROM) mineralized material from the open pit will be crushed and conveyed to the concentrator where the mineralization will be ground to liberate the mineral values from the host rock and then separated by flotation. The bulk copper–silver–gold sulphide concentrate produced will be filtered and introduced into the roasting plant. There the concentrate will undergo a partial oxidative roast, which will remove the arsenic and part of the sulphur into the gaseous phase. The gas will be scrubbed to capture particulate matter and the arsenic into solution. This solution will report to a stabilization circuit that will remove any valuable copper and convert the arsenic into scorodite, which will be filtered and transferred to the tailings management facility (TMF).

After removal of the arsenic, the gas containing sulphur dioxide will be processed through a modular plant to produce saleable sulphuric acid.

The copper precipitate and solid calcine produced from the roaster will be agglomerated and stockpiled. Trucks will transport the combined concentrate to the




port facilities, where it will be placed on ocean-going vessels for transport to overseas smelters.

Key elements of the process design are summarized in Table 16-1, and the process route proposed is included as Figure 16-1. A block diagram of the route is shown in Figure 16-2.

16.3.2 Crushing/Conveying

The primary crushing station will be a fixed 60 x 89 inch gyratory crusher. Mine haul trucks with a capacity of up to 290 tonnes each will dump ROM ore directly into the dump pocket of the crusher.

Crushed mineralized material will fall into a pocket below the crusher, which in turn discharges to an apron belt feeder. The feeder will draw mineralization from the surge pocket at a controlled rate using a variable-speed drive and will discharge to sacrificial coarse ore conveyor No. 1. The conveyor will incorporate a magnet to remove tramp metal and a metal detector to detect remaining tramp metal in the bulk material and thus prevent damage to downstream equipment.

Coarse ore conveyor No. 1 will feed coarse ore conveyor No. 2 for delivery from the crushing station to the coarse ore stockpile (COS). The overland conveyor system from the crusher discharge belt feeder to the COS is sized for a design capacity of 6,900 t/h.

16.3.3 Coarse Ore Storage and Reclaim

The COS will be conical and have a live capacity of 85,000 tonnes and additional dead capacity of 280,000 tonnes. The purpose of the stockpile is to create a buffer between the mining and crushing operations and the process plant so that the plant can continue to receive ore from the stockpile when there are intermittent stoppages of feed from the crusher or the mine. A track dozer will be used to push ore into the feeders when ore needs to be reclaimed from the dead section of the stockpile during upstream maintenance activities.

Ore will be reclaimed from the stockpile with four in-line belt feeders installed in a concrete tunnel below the stockpile. The belt feeders will discharge mineralization at a controlled rate onto the semi-autogenous grind (SAG) mill feed conveyor.




Table 16-2: Key Process Design Criteria

Project Unit Value Throughput Project life years 22 Annual t/a 34,675,000 Daily t/d 95,000 Operating Time Days per year days 365 Plant availability % 94 Run of Mine Ore Moisture content %wt 3.0 Run-of-mine top size mm 1,000 Nominal feed grade

Cu % 0.40 Au g/t 0.05 Ag g/t 1.80

Design feed grade Cu % 0.50

Overall plant recovery (nominal) Cu LOM average % 89.7 Au LOM average % 55.0 Ag LOM average % 50.0

Copper grade in flotation bulk copper concentrate % 26.0 Copper grade in roaster calcine % 30.6 Primary Crushing Haul truck capacity t 290 Dump pocket capacity No. of haul trucks 1.5 Crusher type Gyratory

Crusher OSS mm 168 Discharge P80 mm 125

Coarse Ore Storage and Reclaim Live storage t 85,000 Grinding Circuit configuration - SAB Grinding Parameters Crusher work index (average) kWh/t 9.4

Bond rod mill work index (average) kWh/t 10.2 Bond ball mill work index (average) kWh/t 11.0

JKTech ore breakage test results A - 52.1 b - 1.0 ta - 0.64

Primary grinding Mill type - SAG P80 mm 3.4

Secondary grinding




Project Unit Value Mill type - Ball Ball Mill circulating load % 250 Cyclones P80 μm 200

Flotation Rougher Concentrate

Recovery, Cu % of mill feed 91 Grade, Cu % 4.00

Flotation cells Residence time min 18

Regrind Work index kWh/t 15.00 Circuit product size, P80 μm 30 Cleaner Stage 1 Concentrate

Recovery, Cu % of mill feed 90 Grade, Cu % 20


Cleaner Stage 2 Concentrate

Recovery, Cu % of mill feed 89.7 Grade, Cu % 26.00


Cleaner Scavengers Concentrate

Recovery, Mass % of mill feed 0.59 Grade, Cu % 2.00

Copper Concentrate Thickening and Filtration Thickener underflow density wt% solids 65 Tailings Thickening Thickener underflow density wt% solids 55 Reagents Collector 1 - AP 3894 Collector 2 - PAX Frother - MIBC Flocculant - - Lime - -




Figure 16-1: Process Overview – General Diagram

Figure 16-2: Process Overview – Block Diagram




The SAG mill feed conveyor will deliver the reclaimed ore from the stockpile to the SAG mill and also receive oversize material from the pebble circuit and SAG mill balls, which will be discharged back into the SAG mill. The SAG mill feed conveyor will have a design capacity of 6,300 t/h.

16.3.4 Grinding

The grinding circuit configuration will consist of an open-circuit SAG mill followed by two ball mills in closed circuit with cyclones. Pebbles—material greater than 15 mm leaving the SAG mill—will first be screened out by a trommel screen and then passed over the SAG mill discharge screen to remove the fines and wash the material before recycle back to the SAG mill feed conveyor. The undersize from the two screens will be combined and transferred to the cyclone feed pumpbox. The average final product size from the grinding circuit will be 80% passing 200 µm. Total line throughput will average 95,000 t/d.

The SAG mill will be a 38 ft diameter x 20 ft effective grinding length (EGL) mill, powered by a 24 MW gearless motor drive. The mill will be operated at a ball load of between 15% and 18% by volume, a total mill loading of 26%, and a rotation at 78% of critical speed. The grates will initially be installed with a 75 mm gap, wearing to approximately 90 mm. The pulp lifters will be radial; grate flux will be achieved by adequate discharge cavity depth and proper throat design.

The SAG mill discharge, with a top size of 90 mm, will pass through a trommel screen from where most of the finely ground material will report directly to the SAG discharge mill pumpbox. The remainder will pass onto a 3.6 m wide x 7.3 m long vibrating screen equipped with a 15 mm screen deck. Undersize from the vibrating screen will collect in the SAG discharge pumpbox and be pumped to the ball mill discharge pumpbox. Given the coarse nature of the feed, two installed pumps are considered necessary in this position. The washed pebbles in the vibrating screen oversize will be conveyed back to the SAG mill feed belt. Solid lime will be added to the SAG mill feed to suppress the flotation of pyrite while recovering the copper, gold, and silver minerals in the rougher flotation circuit.

The grinding circuit will have two 26 ft x 40 ft overflow ball mills, each driven by two 7,500 kW single-pinion, fixed-speed drives on either side of the mill, operating in parallel. The SAG mill transfer pump will feed a distribution box that splits the feed between two cyclone feed pumpboxes, one coupled with each ball mill. Each pumpbox will have a single variable-speed pump feeding a cluster of ten 30" cyclones. The average 80% passing particle size will be 200 µm. Cyclone overflows will contain approximately 35% solids. Each cyclone cluster overflow stream will be sampled and the sample analyzed by a particle size analyzer before the stream feeds into the




flotation circuit. Cyclone cluster underflow will report to the ball mill. The ball mills will operate at a ball load of between 30% and 33% by volume, a total mill loading of 35%, and a rotation at 76% of critical speed. A grizzly will protect each cyclone feed pump from large rocks or balls that may be discharged from the mills during upset conditions. Ball mill discharge will join the SAG mill discharge in the mill discharge pumpbox.

Collectors will be added to the mill discharge pumpbox to prepare the material for subsequent flotation.

16.3.5 Flotation

The flotation circuit is designed to recover the minerals containing copper, gold, and silver into a bulk concentrate. The concentrate will subsequently be treated by partial roasting to prepare it for commercial sale. The design of the flotation circuit maximizes the recovery of these valuable minerals at the lowest overall capital and operating cost.

Depending on the mine source location, the mill feed material may at times contain amounts of clay minerals and pyrite. Copper minerals and some pyrite, along with significant entrained and locked non-sulphide gangue, will be recovered in rougher flotation and then be reground to further liberate the minerals from each other. The rougher tailings will still have significant levels of pyrite, which will be removed in a flotation circuit positioned near the TMF to produce material that can be used as cyclone sand for construction of the tailings dam.

The rougher concentrate will be upgraded in two stages of cleaning and sent to the bulk concentrate thickener.

The average 80% passing particle size in the rougher circuit will be 200 µm. Cyclone overflow will be at approximately 35% solids. The cyclone overflow stream will be sampled and the sample analyzed by a particle size analyzer before the stream reports to a distributor and is divided between two rows of six 300 m3 tank-type flotation cells (12 cells total). At 95,000 t/d, the total rougher residence time will be 18 minutes to obtain an average copper recovery of 92% before cleaning. A mixture of A3894 and potassium amyl xanthate (PAX) will be used as a collector.

The rougher concentrate will collect in a series of launders and be directed to a single pumpbox. Two rougher concentrate pumps, one operating and one standby, will feed the regrind circuit.

Frother (MIBC) is added in the flotation circuit, and additional lime is added to the roughers to achieve a pH of approximately 9.8.




The copper cleaner circuit will have one regrind and two cleaner upgrading steps.

The rougher concentrate will be sent directly to densifying cyclones before entering the regrind mill circuit, consisting of two M10,000 IsaMill units. These regrind mills will reduce the concentrate particle size to 30 µm for discharge to the 1st cleaner circuit of four 100 m3 mechanical tank cells. The 1st cleaner concentrate reports to the 2nd cleaner feed. The tails from the 1st cleaners enter the 1st cleaner scavengers, consisting of four 100 m3 mechanical tank cells. The 1st cleaner scavenger concentrate is recirculated back to the feed of the regrind mill system, while the 1st cleaner scavenger tails join the bulk flotation tails.

The 2nd cleaners will consist of four column cells and use recirculating pumps to feed a sparger assembly. The 2nd cleaner concentrate is the final concentrate and is sent to the concentrate thickener. The 2nd cleaner tails are recycled back to the feed of the 1st cleaners.

Lime and frother are the principal reagents in the cleaning circuit. Lime is used as the primary pH modifier throughout and is added at the regrind mills as required to depress pyrite.

16.3.6 Concentrate Dewatering

The final concentrate will be fed into the copper concentrate thickener where flocculant will be added to assist material settling. While the thickener overflow is recycled to the process water system, the thickened underflow is stored in a concentrate storage tank. Material from the concentrate storage tank is fed in batches to the pressure filters for the reduction of overall concentrate moisture content to 8%.

After filtration, the concentrate will be transferred by conveyor to a storage silo with sufficient capacity for eight hours of concentrate delivery. Concentrate from the silo is fed to the roaster circuit for partial roasting to remove impurities.

16.3.7 Tailings Thickening and Disposal

The role of the tailings thickener is primarily to recycle solution at the process plant that would otherwise report to the TMF with the flotation tailings, thereby significantly reducing the volume of decant solution to be pumped back to the plant from the TMF. The tailings thickener underflow will be pumped to the final tailings pumpbox at a density of 55%. Any discharges from the process plant are also added to the tailings pumpbox at this point.




From the pumpbox at elevation 2,930 m, the tailings will be pumped by means of a two-stage pumping system (with installed spare) through a HDPE pipeline to a high point at elevation 2,990 m approximately 650 m north of the plant site, along the main access road. From there the tailings will flow by gravity to the pyrite flotation circuit at elevation 2,985 m.

The tailings line will enter the pyrite flotation plant at the TMF and discharge to a mixing, dampening tank where potassium amyl xanthate (PAX) and frother will be added before the material passes to two 300 m3 self-aerated flotation cells. The pyrite-rich froth is removed and diverted directly to the TMF for subaqueous encapsulation. The pyrite flotation underflow material is pumped to a cyclone plant on the northern abutment of the tailings storage embankment. The cyclone plant will be installed at 2,750 m elevation for use during the initial years of operation and will be moved to 2,980 m elevation for final deposition as the embankment height increases.

The cyclone plant will separate the tailings flow into a sands component and a slimes component. The sand, which comprises approximately 24% of the total tailings stream, will be used for construction of the tailings embankment, while the slimes, the remaining 76%, will be used to form a beach on the upstream side of the embankment.

The tailings sands will exit the cyclones at a pulp density of approximately 74%, requiring the use of positive displacement pumps. Due to the high pressures needed for pumping this material, the slurry will be pumped through a HDPE-lined carbon steel pipe to smaller diameter spigot pipes. The spigots will be spaced at 50 m intervals to evenly distribute the sand slurry along the downstream side of the embankment, where it will be spread and compacted by dozers.

The slimes component will be pumped to the upstream side of the embankment through a HDPE pipeline to spigots spaced 50 m apart along the embankment crest. The slimes will form a beach to help raise the embankment by the centreline construction method.

16.3.8 Roasting and Stabilization

The filtered flotation concentrate from the process plant will contain arsenic at levels that would impair its marketability. Arsenic will therefore be removed from the concentrate by the industrially proven “Partial Roasting” process, which converts the arsenic to its (air oxidation) trioxide (As2O3) form.

The arsenic trioxide, recovered by scrubbing the roaster off-gas with water, will need to be stabilized before disposal in a separate lined impoundment within the TMF,




separated from the flotation tailings. The arsenic is oxidized and reacts with dissolved iron to produce stable, highly insoluble scorodite (ferric arsenate).

During the partial roast, some of the sulphur in the concentrate is oxidized to sulphur dioxide (SO2) and must be removed from the scrubbed (arsenic-free) off-gas. The technology selected for removal is conversion to concentrated sulphuric acid in a conventional package plant. The sulphuric acid will be an upgraded, marketable by-product. The partially roasted concentrate will be agglomerated as required for shipment and smelting.

The concentrate is introduced into a fluid bed roaster operating at a temperature and controlled air flow rate conducive to partial oxidation to release the arsenic and a portion of the sulphur into the gas stream, while maintaining a suitable level of sulphur in the concentrate for smelting requirements.

Figure 16-3 shows the general arrangement and flowsheet for the planned roasting facilities.

The calcined concentrate from the roaster and the dust cyclone underflow will be processed through a calcine cooler and then agglomerated together with the precipitated copper product before being transported by screw conveyor to a storage bin. This is the final concentrate product that will be trucked to the port facilities.

In the stabilization circuit, the scrubber slurry passes to conditioning leach tanks where sulphur dioxide (SO2), in a concentrated bleed stream from the acid manufacturing plant, is added. The role of the SO2 is to create an acidic reducing environment where arsenic and iron are leached out while copper is precipitated with the SO2 to form a copper sulphide. After this controlled reaction, the discharge from the tanks is sent to a fine copper thickener where the precipitated copper sulphide and insolubles (silica, etc.) are thickened to a level that permits filtration. The filtered material will be stored in a surge bin that discharges to agglomeration.




Figure 16-3: Partial Roasting and Arsenic Removal Circuit

The overflow from the fine copper thickener, containing dissolved arsenic and iron, is sent to the impurity stabilization circuit. Air is introduced together with limee to create scorodite. This material is precipitated out of solution in these tanks and is transported to a stabilized product thickener. There the material is thickened and then pumped to a surge tank feeding a filtration circuit that filters out the stable arsenic-containing sludge. The filtered cake is stored in a contained stockpile before being trucked to a dedicated, lined impoundment within the TMF.

Calcine from the roaster and dust cyclones (dry) and copper sulphide precipitate (wet) from the stabilization circuit are withdrawn from their respective storage bins and fed to an agglomerator where the moisture content is adjusted to meet specifications for transport. Intense mixing in this unit results in uniform wetting, blending, and formation of granular agglomerates. The agglomerated material is non-dusting and has favourable material handling properties.

The acid plant will be a modular package plant that will utilize current industry technology and meet environmental guidelines.

In a typical acid plant flowsheet, the roaster off-gas first passes through a scrubber to be cleaned of all fine dust and particulate matter. The SO2-containing gas is then sent to a droplet eliminator and on to a cooling tower where the gas is cooled by means of




circulating weak acid. The gas passes through wet electrostatic precipitators for removal of the finest particles and acid aerosols and is then diluted with atmospheric air and conveyed to a drying tower irrigated with sulphuric acid (96%) flowing counter-current to the gas. Through this contact, the acid absorbs the moisture in the SO2 gas and heat is released. The sulphuric acid is collected and cooled.

The gas from the tower is transported to a converter section where the cold SO2 gas is reheated before entering the first in a series of catalyst beds. SO3 is formed as the gas passes through successive catalyst beds. The gas then enters an adsorption tower where the SO3 is adsorbed by circulating acid, which is collected and transferred to an acid handling system. The gas stream passes through a final adsorption tower, where the remaining SO3 is recovered, and then through a high-efficiency mist eliminator and into the final discharge stack. The emissions from the acid plant will meet applicable environmental regulations for air and water quality.

16.4 Reagents and Grinding Media

The principal flotation reagents for the process are:

• A3894 – liquid flotation collector, promoter • PAX – potassium amyl xanthate, a strong collector • MIBC – the frother used in flotation • Lime – used to control circuit pH • Flocculant – reagent used to help settle solids in the thickeners.

Solid and liquid reagents will be stored in a contained, fenced area, the solid reagents either under cover or in containers and the liquids in either tote containers or bulk tanks. As required, the reagents will be delivered to the concentrator mixing area and prepared for distribution into the system. Each reagent mixing tank will be ventilated to control emissions. The storage and mixing facilities in this area will also have a separate sump system.

Reagents will generally be delivered in tote bags or tote containers. Storage capacity will be provided for a six-week supply of unmixed reagents. Lime required for pH control will be shipped in trucks as pebble quicklime from Chiclayo on the Peruvian coast.

Steel grinding media will be trucked in bulk from Lima and be loaded into storage bins at site. The media for the IsaMill regrind mills will be shipped from Europe in 1 tonne tote bags and will be loaded into the mills by vendor-supplied equipment.




Sulphur dioxide provided as a bleed stream from the acid production plant will be used as a reductant in the conditioning leach circuit.

Limestone will be necessary as a neutralizing reactant for any excess SO2 (sulphuric acid) produced from the acid plant that cannot be upgraded for sale.

Flocculant will be used as a settling aid in the fine copper and stabilized product thickeners.

16.4.1 Process Water

Water used for processing purposes will be derived from several sources:

• Water reclaimed from the TMF • Water from pit dewatering • Internal plant water recycled from the tailings thickener and the copper

concentrate thickener • Water from the waste rock seepage recovery system • Fresh water from Rio Cañariaco for specific plant requirements.

Most process water will be used in grinding and flotation, with lesser amounts used for washing, flushing, and general cleanup. Fresh water will be used where clean water is required. Booster pumps will be added where higher pressures are required.

16.5 Process Control, Sampling, and Assay

Process control will be from a central control room where the operator can perform remote start-up and shut-down. In addition, operators will carry portable computing pads (with wireless links) to provide them with control information updates. A plant-wide radio system will be used for communications.

Instrumentation and control equipment will be used and applied for measurement and control of process variables such as pressure, level, flow, temperature, density, weight, and speed. The degree of instrumentation will be sufficient for safe and efficient control of the process by a minimum number of operators.

All instrumentation will be standard catalogue products from suppliers. Where practical, identical components will be used to promote component interchangeability, minimize spare parts, and simplify service and repairs.

Grinding and regrinding cyclone overflows—primary cyclone overflow and cleaner regrind mill discharge—will be continuously sampled for particle size analysis.




An on-stream x-ray fluorescent analyzer (OSA) will provide continuous analyses of solids content, copper, iron, arsenic, antimony, and sulphur content to enable operators and supervisory control systems to optimize flotation and respond to upset conditions.

A laboratory facility will be provided at the mine site for metallurgical, environmental and analytical services. The facility will be housed in its own building, isolated from the mill and will be able to conduct instrumental analysis for up 200 samples per day.

16.6 Concentrate Receiving and Loadout at Port Site

Copper concentrate from the process plant will be transported by truck to the Marine Terminal Muchik (MTM) in northern Peru, a planned port development by Lumina Copper (Lumina). CCC will construct and operate its own concentrate receiving, storage, and reclaim facilities at the terminal and will share marine handling facilities with Lumina. Annual throughput from Cañariaco Norte is projected to average approximately 400,000 tonnes.

Trucks arriving at the terminal will unload the concentrate product into a truck unloading station consisting of a hopper feeding a conveyor system into the storage building. All trucks will be weighed before and after unloading and will be washed on arrival and before departure. Wash water will be collected and treated at the port facility.

16.7 Proposed Tailings Management Facility

The TMF presented in this subsection was designed to manage tailings for mine operations at a processing rate of 95,000 t/d for 19 years as per the preliminary mine plan. Subsequent to completion of this facility design an updated mine plan was developed that extended the mine life to 22 years.

16.7.1 Location

Two locations with suitable topography for the construction of an embankment and sufficient volume for containment of the projected life-of-mine tailings were identified: the Quebrada Yerma valley and the Río Cañariaco valley. A trade-off study was undertaken to assess the availability of construction materials such as waste rock, local borrow, and cyclone sand; to compare downstream and centreline embankment construction methods for each option; and to evaluate alternative tailings depositions technologies such as high-density thickening. Capital and LOM operating costs were then compared and weighed against the potential socioeconomic impacts and physical




footprint. The trade-off concluded that the Quebrada Yerma valley was the preferable location for building either a cyclone sand or waste rock embankment with centreline construction, or a waste rock embankment with downstream construction for the deposition of thickened, sloped tailings.

16.7.2 Design Considerations

The design criteria utilized are commensurate with appropriate governance literature and/or as appropriate, industry standards for this level of study. The design process has observed all applicable regulations and statutes, and provisions were made to satisfy the interpreted regulatory requirement where feasible.

The following design criteria were adopted for design of the tailings facilities:

• The design incorporates sufficient freeboard to manage the projected tailings volume and inflow design flood (IDF) at different stages of the mine life. For Years 1 to 5 of the mine life, the IDF corresponds to the one-in-1,000 year, five-day storm event. For Years 6 to 19, the IDF corresponds to the probable maximum flood (PMF) 30-day storm event. The criteria for closure are less demanding than for operations because an embankment spillway has been designed and will be constructed to safely manage the PMF 24-hour event. In accordance with the preceding IDF criteria, a minimum of 3 m freeboard will be provided and maintained above the tailings and water level. This comprises 2 m for dynamic settlement, static settlement, and wave run-up due to wind action, and 1 m of dry freeboard as a contingency.

• The tailings starter embankment will be constructed to provide storage for a minimum of one year of tailings production as well as the design storm water and freeboard requirements.

• The impoundment will be operated to ensure that the tailings remain saturated so as to inhibit oxidation.

• A sulphide flotation circuit will be installed adjacent to the TMF to remove sulphide minerals (primarily pyrite) from the tailings. The non-sulphide tailings will feed the cyclone sand plant and will be non-acid generating (NAG). The coarse cyclone underflow will be used to construct the tailings embankment. The fine sand overflow stream will be deposited in the TMF.

The main components of the TMF are as follows:

• Starter embankment. • TMF embankment • Seepage collection (SC) impoundment




• Construction diversion structures • Scorodite management facility.

16.7.3 Starter Embankment

The TMF starter embankment is an earthfill/rockfill embankment, with appropriate filter and transition zones to contain the tailings solids, constructed using locally-borrowed fill materials. The design includes start-up water being supplied from the freshwater reservoir upstream of the waste rock management facility. The embankment will be constructed with 2H:1V upstream and downstream slopes, with a 10 m wide crest. It has been sized to store one year of tailings production from the process plant, as well as appropriate design freeboard allowances, including inflow design flood storage, dynamic and static settlements, wave run-up, and contingency freeboard. The maximum height of the starter embankment will be 105 m (measured vertically at the embankment centreline), with a total volume of 12.5 Mm3.

The starter embankment has been sized to store one year of tailings production from the process plant as well as appropriate freeboard allowances for storage of the IDF, dynamic and static settlements, wave run-up, and contingency. At present there is no plan to store water in the TSF for use during start-up. The slope design required for satisfactory stability will depend primarily on the ground conditions encountered under the proposed embankment structure. This will be analyzed during the next stage of study using limit-equilibrium models once data are available from the planned geotechnical site investigations.

16.7.4 TMF Embankment

The TMF embankment will be raised in stages as a compacted cycloned sand shell above the starter embankment using cycloned sands and centreline construction methods. The cyclone sand plant (CSP) will initially be installed on the right abutment of the TMF embankment at approximately elevation 2,820 masl to produce the cycloned sand required for the downstream shell. As the embankment is raised, the CSP will be moved to a platform at elevation 2,960 masl. This relocation concept will be reassessed at the next stage of study, as there may be an opportunity to move the sands to the embankment for construction by means of gravity flow.

Feed to the cyclone plant will be the tailings from a sulphide flotation circuit that will remove the sulphide minerals. The sulphide flotation concentrate will have acid-generating potential and will therefore be stored subaqueously within the tailings impoundment. Subaqueous deposition limits oxygen contact with the sulphide minerals, thereby substantially reducing the potential to produce acid.




The TMF embankment has been designed with a final downstream slope of 2H:1V, similar to other tailings embankments currently in operation or under construction. The embankment will ultimately be 270 m high and require a total volume of 78 Mm3 of cycloned sand. There is precedent for cyclone sand embankments of this height both within Peru (currently in operation) and other countries. At Los Pelambres, both the Quillayes and El Mauro tailings embankments are currently being constructed as cycloned sand operations with proposed final heights of over 200 m.

The current design requires approximately 80% of the CSP coarse underflow tailings to be available for placement on the embankment at all times during construction to maintain adequate storage and freeboard. This aspect will be addressed during subsequent stages of design when additional information becomes available on the variability of tailings particle size distribution. In the event that the embankment raising schedule cannot be fully met using cyclone underflow tailings, then a minor amount of earthfill / rockfill material may be needed to augment the embankment fill.

16.7.5 Seepage Collection Impoundment

A seepage collection (SC) impoundment has been designed downstream of the main TMF embankment. The SC embankment design is based on similar designs previously permitted and constructed in Peru and includes a compacted earthfill / rockfill embankment with filter / transition zones and a low-permeability, bituminous geomembrane liner on the upstream face.

Trade-off studies should be undertaken to determine the most appropriate design considering the availability of local borrow materials and costs for supplying and installing synthetic and/or imported materials. Slope stability design for the SC embankment will also depend primarily on foundation conditions, for which data will be available after the planned site investigation program has been completed.

16.7.6 Construction Diversion Structures

The SC impoundment has been specifically designed to retain cyclone sand drainage, embankment seepage, and runoff from the embankment and contributing catchment. The SC impoundment is sized to contain the accumulated precipitation from the 1:25-year wet year. In addition, the SC embankment design includes an emergency spillway sized to manage the peak flow during a PMF event. The embankment will reach a maximum height of 50 m (measured vertically at the embankment centreline), creating 1.25 Mm3 of storage volume.




It is anticipated that the SC impoundment will collect water during the operational life of the TMF. During post-closure, the SC embankment could serve as a reservoir for ongoing treatment before release to the downstream environment, depending on the water quality at that time.

16.7.7 Scorodite Management Facility

Mineral processing over the operating life of the mine will result in the production of approximately 30,000 m3 of scorodite—the stable ferric arsenate byproduct from the partial roast process incorporated in the process flowsheet to manage arsenic levels in the concentrate. Scorodite represents approximately 0.011% of the total waste material to be stored in the TMF. It is currently planned to provide a separate lined containment facility for this byproduct. The facility will be capped at closure.

16.7.8 Closure

The design objective for the TMF is to keep the PAG tailings saturated at all times within the impoundment. To achieve this, the PAG tailings will be deposited subaqueously, encapsulated with NAG sands, and be flooded with a minimum 2 m of water cover. Based on the current water balance, the median depth of water cover in the embankment will be 8 m. It is recommended that the water quality and excess water discharge schedule be re-evaluated in the next phase of work to reduce the water cover at closure while determining if there is potential to discharge water and what treatment technology, if any, may be required.


In the opinion of the QPs, the metallurgical test work conducted to date supports the declaration of Mineral Resources based on the following:

• The metallurgical testwork completed on the Project has been appropriate to establish a process route that is applicable to the mineralization types.

• Tests were performed on samples that were representative of the mineralization for the purposes of establishing an optimal conceptual process flowsheet.

• The process route proposed uses conventional technology

• Recovery factors from the tests are appropriate to the mineralization types and selected process route based on the available testwork data. If put into operation, the plant will see recovery factors will vary on a day to day basis depending on grade and mineralization type.




• Recoveries used to support Mineral Resource estimation were variable for copper, based on the copper grade, ranging from 58.8% at 0.1% Cu to 93.1% at 0.6% Cu, 55% for gold and 50% for silver.

• Copper recoveries within the process plant are expected to vary with feed grade ranging from 93.1% at 0.6% Cu and 58.8% at 0.1% Cu. The metallurgical response of gold and silver has not been assessed to the same degree as copper; however, based on the testwork completed to date, gold and silver recoveries are projected to be 55% and 50% respectively.




17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

Mineral Resource estimates were prepared by AMEC employees David G. Thomas, P.Geo., Senior Resource Geologist, and Rodrigo Alves Marinho, (C.P.G., AIPG), Principal Geologist.

17.1 Database

The 2010 AMEC mineral resource estimate uses 15 additional core drill holes that were not considered in the November 2008 estimate because analytical results were not available to SRK at the previous estimate cut-off date. The estimation database comprises 225 core holes (67,485.06 m of core, of which 66,584.29 m have assay intervals) from the Ingemmet, Placer, Surcusal Peruana, and CCC drill programs. The drill database was provided by CCC as MS Excel® spreadsheets. The database cut-off date is 8 November, 2010.

AMEC imported the collar, survey, lithology, alteration, and assay data into GEMS® (version 6.2.4), a commercial mining software program. GEMS® validation routines were used to check for overlapping intervals, missing intervals, and consistent drill hole lengths between tables, and no errors were reported.

17.2 Topography

Topographic contour limits were based on a photogrammetric interpretation undertaken on aerial photograph coverage provided by Horizons South America S.A.C. Aerial Mapping Services, which was accurate to 1 m (x, y, z). Errors noted in the digital terrain model (DTM) by Bonson et al., (2008) were surveyed using a Sokkia total station instrument. These points were used in conjunction with the original topographic data to minimize the differences found. Although AMEC did not find a constant elevation difference between the surveyed collars and the DTM, AMEC is of the opinion that further investigation is required for the next level of study since more precision will be required for earth-movement estimates.

17.3 Models

CCC provided interpretations of lithological units on north–south and east–west vertical sections that were spaced 100 m apart. As the lateral extent of the lithological interpretations in some areas of the deposit are not fully defined by drill data, AMEC expanded the interpretation beyond the limits of the available drill hole information to cover the block model extents. Although AMEC considers such extrapolation to be




reasonable, AMEC recommends that additional drilling be performed to increase the level of confidence of the lithological interpretation in these areas.

Drilling completed post-2008 was added to the existing interpretations. AMEC created bench polygons and extruded the bench plan polygons to the mid-point distance to the adjacent polygons to create lithological solids.

East–west oriented, 100 m spaced vertical sections for alteration types and intensities were prepared by CCC. These were simplified by AMEC, and the resulting vertical polygons extended beyond the limits of the available drill hole information to cover the block model extents. AMEC recommends a full review of the alteration interpretation be completed on vertical sections reconciled to bench plans. This work should be completed for the next level of study.

Estimation domains were defined following evaluation of statistical distributions of lithological and alteration units. AMEC created a total of seven domains for copper and four domains for gold and silver (Table 17-1; refer to Table 7-1 for a description of the lithological unit codes). These domains were tagged to blocks and back-tagged to composites and were used as the basis for matching samples and blocks during the estimation process. For the definition of arsenic domains, AMEC used a probabilistic approach considering an arsenic cut-off grade of 250 ppm as the threshold for low- and high-grade populations.

AMEC defined a supergene-enriched domain that was considered only for the estimate of copper grades. The supergene layer in the Cañariaco Norte deposit is thin and laterally discontinuous.

Copper mineralization at Cañariaco Norte is related to porphyry intrusions, breccias and potassic, phyllic, argillic and advanced argillic alteration; however, the limits of mineralization are not well known in the deposit. AMEC recommends building a mineral zonation model for the next phase of study. A combination of lithology, alteration and mineral zones should provide a more robust support for the mineralized envelope and the definition of grade estimation domains.




Table 17-1: Estimation Domains

Element Domain Code Lithological Units Alteration Units Copper 10 VC, CQFP Weak argillic

20 VC, CQFP All but weak argillic 30 BFP, IBXP Weak argillic 40 BFP, IBXP All but weak argillic 50 CQP, IBXH, TBX Weak argillic 60 CQP, IBXH, TBX All but weak argillic 70 FP Weak argillic, propylitic, intermediate argillic, phyllic

Gold and Silver

10 VC, CQFP, CQP, FP All but advanced argillic 20 VC, CQFP, CQP, FP Advanced argillic 30 BFP, IBXH, TBX, IBXP All but advanced argillic 40 BFP, IBXH, TBX, IBXP Advanced argillic

17.4 Composites

Although the nominal sample length for assays is 2 m, sample lengths in the Cañariaco Norte assay database range from 0.45 m to 100.3 m; such long intervals correspond to non-mineralized zones or to intervals of non-sampling. In order to normalize the weight of influence of each sample, AMEC regularized the assay intervals by compositing the drill hole data into 6 m lengths using no geological or domain boundaries. AMEC then back-tagged the 6 m composites using the lithological and alteration solid shapes and assigned estimation domain codes.

17.5 Exploratory Data Analysis

Exploratory data analysis comprised basic statistical evaluation of the 6 m composites for copper, gold, silver and arsenic.

Box plots indicated that copper displayed similar grade distributions in those domains that had no weak argillic alteration. The biotite feldspar porphyry and polylithic igneous breccia units (BFP and IBXp) are preferentially mineralized, and have higher average grades for all metals. CVs for copper are low, around and below 1, in all domains but domain 10, which consists of volcanic rocks and the crowded quartz porphyry, where the coefficient of variation (CV) is 1.5 for uncapped copper composite values.

Average gold grades are very low in all domains, but box plots indicate that the combination of biotite feldspar porphyry (BFP) and the different types of breccia (IBxP, TBx, and IBxH) host most of the higher-grade gold mineralization where such mineralization is associated with any alteration type other than advanced argillic. All




domains have low CV values for gold, and this confirms the low variability of the gold grades.

AMEC used the same domains for silver as defined for gold due to their good correlation. Silver displays similar average grades and grade ranges to gold. A higher spread of grade ranges was observed in domain 10.

High-grade arsenic values are concentrated in the northwest and south-centre parts of the deposit and AMEC could not clearly define domains based on lithology, alteration or a combination of both. From the analysis of cumulative distribution plots, AMEC defined a threshold of 250 ppm As for low- and high-grade arsenic populations. By splitting the data, AMEC reduced the global CV from 1.7 to 0.7 and 1.1 for low- and high-grade samples, respectively. AMEC recommends that, once a mineralogical model has been constructed, the arsenic distribution be evaluated using the resulting interpreted units.

Contact analyses were completed for copper and gold composite values and AMEC defined soft, firm and hard boundaries from this analysis. To represent the firm contacts, AMEC shared samples only during the first estimation pass.

17.6 Variography

AMEC used Sage2001 software to construct down-hole and directional correlograms for the estimation domains for copper, gold, silver and arsenic. For arsenic, AMEC also created correlograms for the 250 ppm As indicator threshold and subsequent grade correlograms for low- and high-grade domains. Domains that displayed soft boundaries were grouped.

For copper and gold, AMEC used spherical models to fit the experimental correlograms. For silver a combination of exponential and spherical models was used, depending on domain.

17.7 Grade Capping

AMEC evaluated probability plots by estimation domains to define grade outliers for copper, gold, silver and. Outlier values typically occur in the upper 1% of the distribution. Copper and gold values were capped at the thresholds defined. For silver, outlier values were controlled by using a restricted search ellipse, with a radius of 25 m x 25 m x 15 m, during grade estimation. AMEC did not restrict extremely high arsenic values. The grade thresholds for the different element outliers are shown in Table 17-2.




Table 17-2: Outlier Thresholds for Copper, Gold and Silver Element Domain Code Threshold Copper (%) 10 1.10

20 2.00 30 2.50 40 2.50 50 1.10 60 1.20 70 1.10

Gold (ppb) 10 - 20 130 30 530 40 -

Silver (ppm) 10 12 20 12 30 17 40 17

17.8 Estimation Methodology

The block model consists of regular blocks (15 m x 15 m x 15 m) and no rotation is used. The block size was chosen such that geological contacts are reasonably well reflected and to support an open pit mining scenario.

AMEC estimated copper, gold, silver, and arsenic grades by estimation domains using ordinary kriging (OK) interpolation for the majority of domains. Inverse distance weighting to the second power (ID2) was used to interpolate Au and Ag in domains 20 and 40 where variography was not considered sufficiently robust.

The process included:

• Grade estimation was completed in three passes • Sample sharing was based upon the matrix determined from contact profiles • Search orientations for all domains were based upon variogram orientations. • A minimum of 3–11 and a maximum of 9–15 drill hole composites were required

for estimation; this varied by element and estimation pass.

The estimate of arsenic grades was divided into low- and high-grade populations and the final grade calculated by weighting grades by the respective probability.

17.9 Density

AMEC assigned density values to blocks based upon the lithological codes. In AMEC’s opinion, these density values are reasonable for use in mineral resource estimation at this preliminary level of study, but recommends executing a continuous




program of bulk density determinations from core samples, using preferably the same laboratory and determination procedures.

17.10 Model Validation

AMEC validated the Cañariaco Norte block model to ensure appropriate honouring of the input data. A nearest neighbour (NN) model was created to validate the OK model. The validation comprised:

• Detailed visual inspection of block grade versus composited data in section and plan view. The visual inspection of block grade versus composited data showed a good reproduction of the data by the model.

• A comparison between the OK and NN estimates was completed to check for global bias in the copper, gold, silver and arsenic grade estimates. Differences were within acceptable levels and no global biases were noted in the estimates.

• Swath plot validation compared average grades from OK and NN models along different directions. Except in areas where there is currently limited drilling, the swath plots indicated good agreement for all variables.

• The degree of smoothing due to kriging was assessed by considering change of support correction using Hermetian polynomials. AMEC considered blocks from all copper domains and the results show a smoothing of 4% in copper grades but 7% more tonnes above the 0.2% Cu cut-off, resulting in a difference of only 1% in contained metal. The kriging smoothing is within acceptable ranges.

17.11 Mineral Resource Classification

AMEC used the following criteria to pre-classify blocks into categories as:

• Measured mineral resources: composites from a minimum of three drill holes within 75 m radius from a block centroid, or samples from two drill holes with the closest sample within 25 m of the block centroid.

• Indicated mineral resources: composites from a minimum of two drill holes within 110 m distance of the block centroid.

Blocks that were not classified as Measured or Indicated categories, but had a composite within 135 m from the block centroid were classified as Inferred. Remaining blocks were not classified. AMEC used a semi-automated process to smooth the initial classification and avoid islands or isolated blocks of different categories.




17.12 Assessment of Reasonable Prospects of Economic Extraction

AMEC assessed the classified blocks for reasonable prospects of economic extraction by applying preliminary economics for potential open pit mining methods. The assessment does not represent an economic analysis of the deposit, but was used to determine reasonable assumptions for the purpose of determining the mineral resource. Mining and process costs, as well as process recoveries were defined from on-going AMEC studies for the Project.

A large bulk mining open pit operation is envisioned for Cañariaco Norte, featuring large rope shovels working on 15 m benches. A single pit with internal phases is projected to provide mill feed at a rate of 95,000 t/d.

AMEC defined a pit shell, optimized using Whittle® (version 4.3) software, based on the parameters listed in Table 17-3. The metal prices used by AMEC represent long-term estimates for mineral resources. The metallurgical recovery for copper is variable depending on the grade range and is shown in Table 17-4.

AMEC defined a marginal cut-off of 0.14% Cu (see details in Table 17-5). AMEC selected a base case 0.2% Cu cut-off for mineral resource reporting based upon the operating cut-offs of other comparable copper mines in South America.




Table 17-3: Optimization Parameters for Resource Pit Shell

Mining Costs Unit Value Waste Mining Reference Cost US$/t mined 1.11 Mining Sustaining Capex Allowance US$/t mined 0.16

Total Reference Mining Costs US$/t mined 1.27 Process +Tailings + G&A Cost US$/t milled 3.32 Mill Sustaining Capex Allowance US$/t milled 0.28

Closure Costs Allocation US$/t milled 0.10 Total Ore Based Costs US$/t milled 3.70 Cu Price US$/lb 2.50 Au Price US$/oz 1,035.00 Ag Price US$/oz 17.25 Selling Cost US$/lb 0.40 Cu Recovery % variable by grade, refer to Table 17-3 Au Recovery % 55 Ag Recovery % 50 Overall Pit Slope(s) Degree 38

Note: There is an incremental mining cost which increases with height and depth ($0.015 per 15 m bench above 2,800 m, $0.03 per 15 m bench below 2,800 m). No dilution or mining loss adjustments were applied. The overall pit slope assumption is based on limited geotechnical data and will likely change when results from the planned geotechnical drilling become available.

Table 17-4: Copper Process Recoveries

% Cu Head Grade Recovery 0.60 93.1% 0.50 91.8% 0.40 89.7% 0.30 86.3% 0.20 79.4% 0.15 72.5% 0.10 58.8%

Table 17-5: Marginal Cut-Off Calculation

Parameters Value Processing Cost (US$/t) 3.70 Recovery (%) 58.8 Price (US$/lb) 2.5 Selling Cost (US$/lb) 0.40 Cut-Off Cu (%) 0.14




17.13 Mineral Resource Statement

Mineral Resources for the Project were classified under the 2005 CIM Definition Standards for Mineral Resources and Mineral Reserves by application of a cut-off grade that incorporated process-based operating costs and recovery parameters, and constraint of the Mineral Resources to a pit shell based on commodity prices, recovery parameters and operating costs.

Mineral resources are tabulated in Table 17-6. The Qualified Person for the Mineral Resource estimate is David Thomas, P.Geo. Mineral resources are reported at a cut-off grade based on the assumption of long-term metal prices as follows: copper price of US$2.50 per pound, a gold price of US$1,035 per ounce and a silver price of US$17.25 per ounce, and have an effective date of 8 November 2010. Mineral resources that are not mineral reserves do not have demonstrated economic viability.

The numbers reported in Table 17-6 differ from the CCC press release of 1 November 2010 because of the inclusion of additional Ag and As assay results from ICP tests in the estimate for some holes drilled during the 2008 campaign that did not have the final assay results for the full suite of elements at the time of the press release. Differences exist in the contained metal tonnages between this table and the press release, although the tonnes and grades are the same; these differences are not material.

The sensitivity of the mineral resource to a reduction or increase in copper cut-off grades is included as Table 17-7, with the base case highlighted.




Table 17-6: Mineral Resource Statement for Cañariaco Norte at a 0.2% Cu Cut-off Grade (David Thomas P. Geo., Effective Date 8 November 2010)

Grade Contained Metal

Category Tonnage

Mt Cu %

Au g/t

Ag g/t

Copper (Blb)

Gold (Moz)

Silver (Moz)

Measured 406.6 0.44 0.07 1.9 3.977 0.958 24.656 Indicated 596.5 0.38 0.06 1.6 4.964 1.081 30.403 Measured + Indicated 1,003.0 0.40 0.06 1.7 8.941 2.039 55.059 Inferred 293.3 0.33 0.05 1.5 2.165 0.448 13.547

Table 17-7: Mineral Resource Statement for Cañariaco Norte Showing Sensitivity to Various Cut-offs (David Thomas P. Geo., Effective Date 8 November 2010).

Cut-Off Grade Contained Metal

Cu (%) Tonnage

Mt Cu %

Au g/t

Ag g/t

Copper (Blb)

Gold (Moz)

Silver (Moz)

Measured

0.14 425.6 0.43 0.07 1.9 4.049 0.986 25.467 0.2 406.6 0.44 0.07 1.9 3.977 0.958 24.656 0.3 338.1 0.48 0.08 2.0 3.588 0.853 21.399 Indicated

0.14 680.7 0.35 0.05 1.5 5.283 1.166 33.124 0.2 596.5 0.38 0.06 1.6 4.964 1.081 30.403 0.3 414.3 0.43 0.06 1.7 3.945 0.843 23.135 Measured + Indicated

0.14 1,106.2 0.38 0.06 1.6 9.332 2.152 58.592 0.2 1,003.0 0.40 0.06 1.7 8.941 2.039 55.059 0.3 752.4 0.45 0.07 1.8 7.533 1.696 44.534 Inferred

0.14 419.4 0.28 0.04 1.3 2.634 0.561 17.002 0.2 293.3 0.33 0.05 1.5 2.165 0.448 13.547 0.3 157.7 0.41 0.06 1.7 1.433 0.281 8.539

Notes to accompany Mineral Resource and Sensitivity Tables:

1. Mineral resources that are not mineral reserves do not have demonstrated economic viability 2. Mineral Resources base case is reported at a 0.2% Cu cut-off grade 3. Mineral Resources are reported as undiluted. 4. A Lerchs–Grossmann pit shell was used to constrain the Mineral Resources to assess reasonable prospects

of eventual economic extraction 5. Mineral Resources are reported using assumed long-term prices as follows: copper price of US$2.50/lb,

gold price of US$1,035/oz and silver price of US$17.25/oz 6. Rounding as required by reporting guidelines may result in apparent summation differences between

tonnes, grade and contained metal content 7. Tonnage and grade measurements are in metric units. Contained gold ounces are reported as troy ounces,

contained copper pounds as imperial pounds





The QPs are of the opinion that the Mineral Resources for the Project, which have been estimated using core drill data, have been performed to industry best practices, and conform to the requirements of CIM Definition Standards (2005).

Areas of uncertainty that may materially impact the Mineral Resource estimates include:

• Long-term commodity price assumptions • Long-term exchange rate assumptions • Operating and capital assumptions used • Metal recovery assumptions used • Any changes to the slope angle of the pit wall as a result of more detailed

geotechnical information would affect the pit shell used to constrain the mineral resources.




18.0 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORT ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES

As the Project is not at a development stage, this section is not relevant to the Report.




19.0 OTHER RELEVANT DATA AND INFORMATION

This section reports the results of an updated Preliminary Assessment on the Project, which is based on a Project report completed in mid-January 2011. AMEC cautions that the Preliminary Assessment is preliminary in nature and that there is no certainty that the Preliminary Assessment will be realized. The Preliminary Assessment is based upon Measured and Indicated mineral resources; mineral resources that are not mineral reserves do not have demonstrated economic viability.

Work performed and reported in this subsection was based on a pre-feasibility study commissioned by CCC, and is based on interim information from that study. Additional geotechnical drilling and rock quality assessment is required to complete the open pit slope design to a level consistent with generally accepted pre-feasibility requirements. The above ground structures in the tailings management facility were designed to a pre-feasibility level; however, geotechnical investigations in the TMF area have not been completed and as a result there is a risk that the TMF design may require revision. As a result, the information from the interim study findings are presented as a Preliminary Assessment. Results of this Preliminary Assessment supercede earlier economic evaluations reported for the Project.

19.1 Proposed Mining Operation

A single pit with internal phases is proposed to provide process feed at a rate of 95,000 t/d. The proposed mine plan is based on diluted Measured and Indicated mineral resources contained in a Lerchs-Grossmann (LG) optimized pit shell generated at metal prices of $2.25/lb Cu, $930/oz Au, and $15/oz Ag.

19.1.1 Dilution

Revenue is generated from the sale of copper concentrates, which contain payable co-products of gold and silver. Copper metal recoveries are variable with head grade. To capture the multi-element and variable recovery complexity, net smelter return (NSR) values were calculated for cut-off application and block valuation during pit optimization.

Mineralization loss and dilution adjustments were performed by applying a “mixing zone” across economic boundaries. No additional mineralization loss was applied. The NSR calculations and dilution/mining loss adjustments were performed using custom scripts in MineSight®.




This procedure used three scripts to:

• Calculate the NSR from the pre-diluted grades

• Adjust the dilution/mining losses across economic boundaries defined by the marginal NSR cut-off, with Inferred material being treated as waste. A second set of block model grade items (CuDil, AuDil, AgDil, AsDil) was populated.

• Calculate the diluted NSR (NSRD) from the diluted grades.

The NSRD was then used for pit optimization and cut-off application for mine planning.

19.1.2 Pit Geotechnical Design Recommendations

The open pit mine design relating to the angle of the pit wall slopes has been developed based on geotechnical logging of drill core plus rock quality evaluation and compressive strength testing of a limited number of core samples. The amount of geotechnical data available is not sufficient, however, to support development of a pre-feasibility level design for the pit wall slopes. As a result, the mine plan developed for this report must be considered to be at a preliminary level.

A site investigation program has been proposed to support geotechnical pit design at a pre-feasibility level. The data from the site investigation are considered to be key in understanding the joint patterns and geotechnical complexities associated with the highwall (approximately 925 m high) and the sensitivity of the highwall slope to pit depressurization and rock mass strength.

A preliminary geotechnical analysis was performed to provide credible slope angles to be used to develop preliminary LG pit shapes. The geotechnical analysis considered three design sectors (Table 19-1).

For the purposes of pit optimization, sectors 1 and 2 (Figure 19-1) were combined and flattened to account for ramp and geotechnical berm locations, resulting in overall slopes of 35° and 44° for Sectors 1 + 2 and 3, respectively.

19.1.3 Pit Hydrogeology

Packer test data indicate that permeable features such as faults or fracture zones are present at depth. Pit dewatering will therefore be undertaken by a combination of vertical dewatering wells, specifically constructed in these permeable features, and horizontal drains in areas where the hydraulic conductivities are lower than approximately 1 x 10-7 m/s.




Table 19-1: Design Sector Slope Recommendations

Design Sector

Inter-ramp Angle

(degrees)

Geotechnical Berm Overall Slope

Width (m)

Vertical Interval (m)

Max Slope Angle (degrees)

Max Slope Height (m)

1 38 50 250 35 925 2 38 50 250 35 925 3 48 50 250 45 500

Figure 19-1: Pit Design Sectors

To determine potential dewatering requirements, the bedrock was divided into three regimes. The upper regime, generally less than 100 m deep, could be dewatered with vertical wells if the hydraulic conductivities were high enough; this should be confirmed moving into the next phase of the study. The intermediate regime, at 100 m to 300 m depth, is generally of low permeability and will be depressurized with horizontal drains.

In this case there is evidence that higher-permeability structures are present down to 300 m, indicating that vertical wells may be required in these areas at depth within the pit. The deeper regime, below 300 m, will be depressurized by horizontal drains.

The model and water balance will be refined on receipt of additional hydrogeologic information from planned drilling, and the inflow predictions will be re-assessed during the next phase of study.




19.1.4 Pit Optimization

The reference mining cost developed for the pit shell generation is US$1.27/t total for waste and mineralized material mined. Costs were based on a first-principles cost estimate from a previous preliminary mine plan. Incremental haulage costs were based on the increased distance from the pit rim (estimated at level 2,800 m) and are as follows:

• For every bench 15 m above the pit rim, an increase in haulage costs of US$0.015/t

• For every bench 15 m below the pit rim, an increase in haulage costs of US$0.030/t.

Applying this reference mining cost and the elevation-based cost adjustment factors to the preliminary mine plan results in an average mining cost of $1.42/t mined.

A combined mineralization-based cost of US$3.57/t milled was used for pit optimization. Because the mineralization-waste delineation was performed using an NSR block value, the combined mill feed-based cost of US$3.57/t milled represents the marginal breakeven cut-off grade. Only Measured and Indicated materials were processed. Inferred material was treated as waste.

Sensitivity optimization runs were performed to determine the impact of changes to pit slopes and operating costs. The results show that the LG pit shells are more sensitive to a pit slope reduction of 2° than an increase of 2°. The LG pit shells are more sensitive to a 10% increase in operating cost than a 10% decrease.

19.1.5 Mine Production Schedule

A mine plan was generated using LG pit shells adjusted by minimum mining width. The next iteration of mine planning will require pit phase designs, considering pioneering access to upper benches, internal ramp access, minimum mining widths, and depressurizing and dewatering requirements.

The input parameters utilized for the mine plan development included a target mine life between 18 and 22 years, a life-of-mine (LOM) head grade of no less than 0.40% Cu, and a strip ratio of less than 1:1. To meet these criteria, a series of constant LOM elevated cut-offs were tested on the collection of nested pit shells.

Internal to the selected ultimate LG pit shell, three intermediate phase mining shapes were established. Shells 3, 9, and 15 were selected for phase development based on their ability to maximize grade and reduce pre-stripping requirements early in the LOM,




while keeping the process plant at full production capacity in high-level Whittle™ schedules. These shells were adjusted for a minimum mining width of 60 m (2 x the 30 m re-blocked block dimension). Because the starter pit had a relatively low east-side wall, which was predominantly in relatively competent rock, it was re-generated using a 40° overall slope in the upper wall of sectors 1 and 2.

During the next iteration of mine planning, AMEC recommends that cut-off optimization be performed in conjunction with the use of a more-detailed shell selection methodology.

The selected mine plan resulted in a total of 728.2 Mt of mill feed and 713.5 Mt of waste (0.98:1 strip ratio) over a 22-year mine operating life. The LOM plan focuses on achieving the required mill feed production rate, mining higher-grade material early in the schedule, and balancing strip ratios. The mine was divided into four phases (Figure 19-2) with the intent of maximizing grade in the early years, reducing pre-stripping requirements in the early years, and maintaining process plant production at 95,000 t/d (except in Year 1, when the plant is expected to achieve 80% of its capacity.

Figure 19-2: Proposed Pit Phases (figure looks northeast)

The projected LOM production schedule is included as Table 19-2. Figure 19-3 and Figure 19-4 show the projected material movement by year and the annual projections for mill feed, respectively.

Preproduction stripping was scheduled within Year -1, with Year 1 representing the commencement of processing at a throughput projection at 80% of maximum, allowing for plant ramp-up after commissioning. Approximately 8 km of pit access construction will be required during Year -2 and another 4 km in Year -1.




Table 19-2: LOM Production Schedule

Year Ore Mined

(Mt) Plant Feed

Mt)

Waste Mined (Mt)

Total Mat (Mt)

StripRatio

NSR($/t)

Copper (%)

Gold (g/t)

Silver(g/t)

Arsenic(g/t)

-1 1.3 - 46.2 47.5 35.65 - - - - -

1 26.4 27.7 45.3 71.7 1.71 17.6 0.46 0.08 2.46 468

2 34.7 34.7 40.3 75.0 1.16 19.3 0.50 0.09 2.14 305

3 34.7 34.7 40.3 75.0 1.16 17.2 0.45 0.08 1.83 261

4 34.7 34.7 38.7 73.4 1.12 14.9 0.39 0.06 1.95 354

5 34.7 34.7 38.3 73.0 1.11 17.3 0.45 0.08 2.03 278

6 34.7 34.7 38.3 73.0 1.11 16.4 0.43 0.08 1.96 231

7 34.7 34.7 38.3 73.0 1.11 16.4 0.43 0.08 1.82 184

8 34.7 34.7 38.3 73.0 1.11 15.9 0.41 0.08 1.67 156

9 34.7 34.7 38.3 73.0 1.11 14.2 0.37 0.07 1.52 163

10 34.7 34.7 38.3 73.0 1.11 11.6 0.32 0.05 1.32 208

11 34.7 34.7 38.3 73.0 1.11 11.2 0.31 0.04 1.38 307

12 34.7 34.7 38.3 73.0 1.11 13.4 0.36 0.05 1.60 302

13 34.7 34.7 38.3 73.0 1.11 14.9 0.39 0.06 1.70 263

14 34.7 34.7 38.3 73.0 1.11 15.8 0.42 0.07 1.72 194

15 34.7 34.7 37.5 72.2 1.08 16.1 0.42 0.07 1.76 200

16 34.7 34.7 29.3 64.0 0.85 15.5 0.41 0.07 1.63 153

17 34.7 34.7 25.9 60.6 0.75 12.0 0.33 0.05 1.27 168

18 34.7 34.7 13.5 48.1 0.39 11.0 0.30 0.04 1.30 264

19 34.7 34.7 6.8 41.4 0.20 13.2 0.36 0.05 1.50 241

20 34.7 34.7 4.3 38.9 0.12 16.5 0.43 0.07 1.76 279

21 34.7 34.7 1.7 36.4 0.05 17.5 0.46 0.07 1.76 174

22 7.0 7.0 0.4 7.4 0.06 18.7 0.49 0.07 1.99 236 Total 728.2 728.2 713.5 1,441.7 0.98 15.1 0.40 0.07 1.71 243




Figure 19-3: Material Movement by Period

Figure 19-4: Mill Feed Characteristics by Year

Notes: Although the term “ore” is used in these figures, no mineral reserves have been declared for the Project. The term is used to denote the mineralized material expected as mill feed and material moved.




The mining schedule attempts to reflect the challenges of developing multiple phases in steep terrain. The average mining rate is approximately 185 kt/d with a peak of 205.4 kt/d in Years 2 and 3. Material re-handle required during access and mining of the upper benches was not included in the estimate for this level of study.

The mine plan and equipment productivities assume that a small, short-term stockpile will be provided near the primary ore crusher to act as a buffer between the mine and mill; however, no blending of stockpiled material has been included in this schedule.

Detailed pit phase and access design will be required during the next phase of mine planning. Mine scheduling will need to consider the time required to develop access to the upper parts of the phases and the requirements to bridge drainages.

19.2 Waste Rock Management Facility

A single elongated waste rock management facility (WRMF) will be developed adjacent to the pit (Figure 19-5). “Bottom-up” methods will be used for construction. Preliminary waste rock characterization suggests that potentially acid generating (PAG) rock will be present, and so encapsulation storage methods may be required. More rock geochemistry information is required to develop a waste rock characterization block model and plan for handling these materials. This work is recommended for the next phase of study.

Figure 19-5: Proposed Layout, WRF




19.3 Proposed Equipment

Large-scale mining equipment was selected for the conventional bulk open pit mining approach and to meet the required ore and waste mining rates. Mining will be performed on 15 m benches, with production drilling performed by electric 311 mm diameter blast hole drills. Rope shovels of 43 m3 bucket size were selected for loading activities, as they were a good match to the material movement requirements and the 15 m mining bench. A 40 m3 front-end loader (FEL) will assist with production and re-handle from the short-term stockpile and will act as a backup to the rope shovels. Haul trucks of 290 t capacity were selected in an effort to minimize the fleet size and match the loading unit and material movement requirements.

A fleet of suitably-sized support and ancillary equipment has been specified. Mine support activities include access road construction, pit cleanup, in-pit ramp maintenance, haul road maintenance, dozing, road grading, haul road dust control, and water diversion and management.

Primary equipment requirements were calculated from engineering estimates of productivities, activities, and quantities of material moved. Haul profiles from a previous preliminary plan were used to estimate truck requirements. Support and ancillary equipment unit requirements were determined by estimates of activities.

Equipment requirements are as summarized in Table 19-3.




Table 19-3: Mine Equipment Requirements

Equipment Type No. of Units Primary Equipment P&H 4100C Shovel 2 L2350 Front-end Loader 1 930E Haul Truck 21 Sandvik 1190E Drill 3 Sandvik DP1500 Drill 2 Subtotal Primary Equipment 29 Support Equipment Cat D11R Track Dozer 2 Cat D10T Track Dozer 3 Cat 24M Motor Grader 1 Cat 16M Motor Grader 2 Cat RTD834 Rubber-tired Dozer 2 Cat 785C Water Truck 2 Cat 365DL Track Excavator 1 Cat 330DL Track Excavator 1 HM400 Articulated Truck 3 Subtotal Support Equipment 17 Ancillary Equipment General Service Truck 2 Hiab 105 Crane Truck (10 t) 1 Mine Rescue Ambulance 1 Komatsu WA320 Wheel Loader w/ Tire Handler 2 Adanac Fuel Lube Truck 3 Cat P36000 Forklift 1 Personnel Bus 2 Pickup Truck 16 Lighting Plant 20 kW 8 Lowbed Tractor Trailer 1 Cat 988 MCS Cable Reeler 1 Liebherr LTM 55 Mobile Crane 1 Flygt Pump 104 kW 1 Flygt Pump 65 kW 2 McElroy Pipe Fuser 2 Mobile Jaw Crusher 1 Subtotal Ancillary Equipment 45 Total 91




19.4 Proposed Water Management

The surface water management plan for the Cañariaco Norte project will preserve to the maximum extent possible the “non contact” status of surface waters. Waters that come into contact with Project facilities will be contained and treated if necessary so that any water released to the environment always meets the applicable water quality regulatory guidelines.

A system of impoundments, embankments, diversions, and spillways will be developed immediately before construction commences to handle runoff from construction-related activities (Figure 19-6). This system will continue to be developed through construction and operations to ensure that water discharged to the environment meets the rebulatory guidelines.

The Project will affect two main catchment areas: the upper reaches of both the Cañariaco valley and the Quebrada Yerma valley. Water will be impounded upstream of the WRMF to supply the freshwater makeup requirements of the process plant. A diversion channel will be constructed to direct excess water from Río Cañariaco around the WRMF. Contact water from the facilities in the area of the pit, process plant, and WRMF will be collected immediately downstream of the WRMF and be used as mill makeup water; the excess will be discharged to Río Cañariaco via water treatment facilities. Because most of this water is consumed by the mill, the site water balance indicates that such discharges are only expected toward the end of the mine life.

Precipitation on the TMF and runoff draining into the TMF will be impounded to ensure a minimum water cover is maintained over the tailings. Excess water will be directed to the process plant as makeup water. At closure, a TMF spillway will be constructed on the east saddle embankment to route storm water flows to the Quebrada Yerma. The TMF spillway will be designed for the PMF. The site-wide water balance estimates that a water release from the TMF would not be required until the end of the mine life.




Figure 19-6: Layout Plan, Proposed Water Management Structures




19.5 Capital Costs

Capital costs were based on a combination of budget pricing from local suppliers contacted specifically for this Project and relevant in-house data from AMEC and Ausenco. Labour rates were calculated using typical Peruvian wages and benefits based on recent cost surveys. These were blended with expatriate support labour to achieve appropriate crew mixes.

The estimate was developed in accordance with pre-feasibility study criteria, philosophy, and preliminary project design. All costs are expressed in 3rd quarter (Q3) 2010 U.S. dollars. No allowance has been included for escalation, interest or financing fees, taxes or duties, or working capital during construction. The level of accuracy for the estimate is ±20% of estimated final costs, as per AACE Class 4 definition.

Capital cost estimates are as follows:

• Mine capital costs include the mine mobile equipment, pre-stripping, mine access roads, earthworks associated with water diversion around the mine, dewatering, and drainage, and the explosives facilities. Mine mobile equipment including drills, haul trucks and shovels would be supplied under a lease agreement with the equipment supplier. The mobile equipment capital cost represents a LOM total of $292 million. The explosives emulsion plant will be provided by a contractor based on bulk supply and is included in the contractor’s rate build-up. The mine capital costs total $172 million.

• The direct capital cost of the process plant is $381 million, and includes all process equipment and structures from the primary crusher to the tailings management and reclaim water systems, as well as concentrate handling and loadout on site.

• Site and services capital totals $133 million, and covers infrastructure and facilities required to support the mine/mill operations, including plant site preparation, civil costs, and permanent accommodations.

• Tailings embankment construction and facilities, including seepage infrastructure and reclaim costs total $101.5 million.

• Direct capital costs for the construction of the Project access road from Corral Quernado Road to the site total $39 million.

• Water diversion capital totals $51 million, and consists of the construction of a diversion channel around the proposed waste management facility, including the freshwater diversion embankment and associated seepage infrastructure and facilities.




• Costs for the port site facilities to be constructed by CCC include port site development and utilities, and concentrate receiving, storage, and handling. The shiploader and dock will be constructed by Lumina; CCC has included a usage fee, which is carried in the operating costs. CCC direct construction costs at the port total $27 million.

• Indirect costs for the project total $296 million and include the services required to support the construction activities.

• Contingency accounts for unforeseen costs within the project scope and was calculated using an average (of both direct and indirect capital costs) factor of 18%, which is believed to be appropriate for the level of engineering work performed in the preparation of this pre-feasibility progress report. Contingency has been applied to all the capital cost estimates in this report, excluding working and sustaining capital. Input variables used in calculating the contingency are a result of information gathered from all parties involved in the basis of the estimate. The contingency is $236.5 million.

The total estimated capital cost to design and build the Cañariaco Norte Copper Project is $1.437 billion (Table 19-4). This cost is based on a leased mining fleet and Owner-operated process plant, and self-performed pre-production development. The estimate covers the direct field costs of executing the Project, plus the Owner’s indirect costs associated with design, construction, and commissioning.

19.6 Operating Costs

Average salary and wage rates used were based on AMEC in-house data from recent projects in South America. The work schedule is based on an operating time-frame of 24 h/d, 7 d/wk, 365 d/a. Based on a WTI average price of US$80.00 per barrel, a diesel cost of $0.72/L, including delivery, was used for operating cost calculations.

The power price for the project was based on costs submitted by a Peruvian utility provider at $50.51/MWh. The cost includes the price of power, connection fee, transmission costs, and the cost of generation.

Transportation costs are included within the costs for supplies and materials calculated for mining and processing and are based on conventional highway truck supply. Costs per tonne were based on quotations from local trucking firms. Transport cost for supplies and materials priced from the closest port to site is $25/t. Transport from Lima is $35/t, and an allowance of $260/t, based on quotations from freight-forwarders, is included to cover both sea and land freight for bulk goods procured internationally.




Table 19-4: Summary of Capital Costs Area 2010 ($000) Prestripping 70,008 Mining Infrastructure & Equipment 101,942 Processing Plant & Acid Plant 381,277 On Site Infrastructures 133,444 Tailings Management 101,527 Project Access Road 38,795 Water Diversion and Reservoir 50,753 Port Site Facilities 26,887 Total Direct 904,633 Owner’s Cost 52,899 Indirects 243,076 Total Indirects 295,975 Contingency 236,522 Total Capital Cost 1,437,160 Escalation (excluded) — Working Capital (included in Financial Analysis) — Total 1,437,160

The mine operating cost estimate incorporates costs for operating and maintenance labour, staff, and operating and maintenance supplies for each year. Operating and maintenance supplies are based on a combination of in-house and vendor-supplied data and are exclusive of taxes. Consumables (fuel, explosives, supplies etc.) were calculated from expected use, unit consumptions, and allowances for minor items, based on in-house information combined with “first principle” costs developed for project specifics. All mining costs are based on production Years 1 to 22. Pre-production costs have been capitalized and are included in the capital cost estimate. Mining costs are included in Table 19-5.

Processing costs include the costs of operating and maintaining the processing facilities, from the primary crusher through to concentrate loadout, as well as treatment of concentrate in the on-site roasting facility, production of by-product sulphuric acid, provision of process and reclaim water pumping, and tailings management.

Tailings management includes pumping mill tailings to the tailings management facility and treatment of some of the tailings to remove pyrite and other sulphide materials for continuous placement of sands on the tailings embankment. The processing costs account for the expenses associated with purchasing consumables, equipment maintenance, personnel, and power consumption.




Table 19-5: LOM Mining Cost by Category ($000)

Operating Costs Total $/t mined $/t milled

Admin/Overhead 114,152 0.08 0.16

Drilling 101,710 0.07 0.14

Blasting 303,170 0.21 0.42

Loading 205,359 0.14 0.28

Hauling 944,514 0.66 1.30

Support 266,757 0.19 0.37

Ancillary 45,587 0.03 0.06

Dewatering 14,000 0.01 0.02 Total 1,995,248 1.38 2.74

Consumables costs include items such as crusher liners, mill liners, grinding media, all chemical reagents, and an allocated cost for office / laboratory supplies. The reagent costs are inclusive of freight for shipping the items to site. Workforce organization, salaries, and hourly costs were based on AMEC in-house data for recent South American projects. Equipment maintenance supplies and materials are estimated as a percentage of the capital cost of equipment. Power consumption was derived from the estimated load of individual pieces of equipment on the equipment list. Processing operating costs are included in Table 19-6.

The general and administrative (G&A) operating costs are the labour and overhead expenses for cost centres that are not directly linked to the mining and process disciplines. G&A for each cost centre was estimated either from first principles or input from AMEC based on other operations. Maintenance costs have been assigned to G&A to cover maintenance costs not specific to either the process plant or mine. G&A costs are included in Table 19-7.

19.7 Markets

19.7.1 Treatment Charges

The anticipated quality of the copper concentrate to be produced is based on metallurgical testing carried out by SGS Lakefield in Santiago and Outotec (Sweden) AB. Other than arsenic, which has been managed by the partial roast process now included in the process flowsheet design, no other significant deleterious constituents are anticipated, although it should be noted that antimony is near the normal penalty levels before roasting of the concentrate, and significantly below penalty limits post-roast. Overall, the concentrates are considered to be of good quality.




Table 19-6: Average Annual Processing Costs

Item Average ($/a) $/t Milled

Labour 24,526,600 0.71

Reagents and Consumables 37,529,300 1.08

Power 39,083,800 1.13

Maintenance Supplies 6,729,800 0.19 Total 107,869,500 3.11

Table 19-7: Summary of G&A Cost Estimate

Detail Average Total

($/a) $/t milled

Site Services 6,414,900 0.19 Access Road Maintenance 1,253,500 0.04 Environment 2,250,000 0.06 Freight Charges 1,406,600 0.04 Miscellaneous 3,500,000 0.10 Total 14,825,000 0.43

The estimation of treatment and refining charges for the study is based on long term projections (post 2013) contained in a report prepared by Neil S. Seldon & Associates Ltd. (NSA) and current AMEC in-house data.

The key determinants for future treatment and refining charges (TC/RC) are the supply/demand balance for copper concentrates, smelter economics, and spot market activity as assessed by NSA.

NSA indicated that the annual benchmark numbers have essentially troughed and that a gradual increase can be expected over the next two to three years, resulting in treatment charges approaching equilibrium levels around 2013. Looking forward to post 2012, charges will ultimately be required to approach a sustainable economical level for the smelting industry. Given such economics and the expectation of higher copper prices; reflecting the falling dollar value, price participation or some form of price sharing, such as a step up in the level of TC/RC at higher prices, could eventually reappear.

AMEC used TC and RC levels of $75/t and $7.50/t, respectively. Following discussions with NSA regarding the probable range on long-term projection, the benchmarked figures used by AMEC were determined to be reasonable and fell within the range considered by NSA. AMEC used a penalty limit of 0.2% for arsenic, which differs from NSA’s recommendation of 0.1%. AMEC does not consider the differences




to be material because the smelter term assumptions are at a high level and preliminary in nature, given that no contact has been made with smelters regarding specific Cañariaco concentrates.

19.7.2 Concentrates

From a chemical quality standpoint, the Cañariaco Norte concentrate production is anticipated to be readily acceptable to smelters. At 30% copper, and with sulphur content at approximately 20%, the higher copper and lower sulphur content of the partially roasted concentrate is seen as important to smelters today as the copper content of concentrates for mines such as Escondida are falling, and as a result, smelter input blends have fallen to ±28%.

19.7.3 Acid

The partial roasting facility would generate approximately 140 kt/a of by-product sulphuric acid. Caliper Metals Corp (CMC), on behalf of NSA, was commissioned to produce a high-level evaluation of the marketability of the acid both domestically in Peru and in the export market.

Acid markets in Chile and Peru are centred on the consumption of acid for SX/EW leach processes and metallurgical acid supply. Chile in particular has evolved into an intensive user of acid to supply its large and growing base of SX/EW production. It is anticipated that total SX/EW production volumes will remain significant and will fully utilize domestic Chilean acid-producing capacity and still require additional imports.

The acid market dynamics in Peru are considerably different and on a reduced scale. Operations and developments in copper mining are still in the early stages relative to potential. Peru’s geographic proximity to Chile’s SX/EW-consuming northern regions offers competitive advantages over deliveries of Chilean domestic acid from central regions. The Chile/Peru supply demand balances remain in net deficit for the period examined ending in 2017. From a Peruvian acid producer perspective, this is particularly positive as domestic options increase while Chile remains a viable export alternative.

AMEC used a long-term acid price of $90/t based on preliminary discussions with CMC and NSA in October 2010. A more detailed market study for domestic and export acid markets is recommended for the feasibility study.




19.7.4 Contracts and Agreements

It is expected that normal commercial terms would operate for any smelting agreement that may be reached. The terms contained within the sales contracts are likely to be typical of, and consistent with, standard industry practice, and be similar to such contracts elsewhere in the world.

19.8 Taxation

The information on taxation is derived from the Federation of International Trade Associations (FITA) Global Trade Portal, the website of the Federation of International Trade Associations (FITA, 2007), and was current as at 2007. The QPs have not verified this information, and have relied upon the website for the data presented.

Detailed taxation regime review would be undertaken with feasibility-level studies; the data presented in this subsection is for information purposes only.

There is a taxation rate of 27 percent on net profits for Peruvian-resident companies. The distribution of profits to non-residents is generally taxed at 4.1 percent, resulting in an effective taxation rate of 30 percent. Branch companies are subject to the same tax as the Peruvian companies at the rate of 30 percent. The corporate income tax is calculated on the basis of a tax unit (UIT), which is recalculated annually to account for the effects of inflation.

A special regime applies to small businessmen and to some low income companies whereby the tax rate 2.5 percent of monthly income. Capital gains are taxed in Peru at the rate of 30 percent. Dividends paid by subsidiaries to their parent companies are subject to a four percent withholding.

The fiscal year begins on 1 January and ends on 31 December of the same year. A levy on the income to any individual is based on a system of annual units fixed by the Peruvian Government. A unity in January 2007 was equal to 3,100 PEN, such that from zero to 27 units the tax rate was 15 percent, rising to 21 percent taxation on an income of 27 to 54 units. Beyond 54 units, the tax rate was 27 percent. Various deductions and allowances reduce the individual's taxable income. A total of 20 percent of the salary and wages paid by a company to its employees is tax-exempt.

The standard value-added tax (VAT) rate is 19 percent, two percent of which goes to the local level as a municipal promotion tax. In special circumstances, a reduced rate of two percent VAT may be payable. Exemptions include some basic foodstuffs, urban passenger transport, the international transport of cargo, life assurance polices, some financial products, books and the construction and maintenance of ships.




19.9 Financial Analysis

The results of the economic analysis represent forward-looking information (cashflows, net present value, internal rate of return, production rate, and total metal produced) that are subject to a number of known and unknown risks, uncertainties and other factors that may cause actual results to differ materially from those presented here. As with most projects at this level of assessment, risks exist that may affect the development of the Project. Factors that could pose a risk to the Cañariaco Norte Project include changes in world commodity markets, equity markets, costs and supply of labour and materials relevant to the mining industry, extent of mineral resources actually contained in the deposit, geotechnical conditions, actual recoveries achieved in processing the mineralized material, marketing of concentrate, technological change, water management, local community support, environmental permitting, change in government, or changes to regulations affecting the mining industry.

The Project has been evaluated using a discounted cash flow (DCF) analysis. Cash inflows consist of annual revenue projections for the mine and two years of preproduction. Cash outflows such as capital, operating costs, and taxes are subtracted from the inflows to arrive at the annual cash flow projections. To reflect the time value of money, annual net cash flow (NCF) projections are discounted back to the project valuation date using several discount rates. The discount rate appropriate to a specific project depends on many factors, including the type of commodity and the level of project risks, such as market risk, technical risk, and political risk. The discounted present values of the cash flows are summed to arrive at the project’s net present value (NPV). In addition to NPV, internal rate of return (IRR) and payback period are also calculated. The IRR is defined as the discount rate that results in an NPV equal to zero. Cash flows are taken to occur at the end of each period. Capital cost estimates have been prepared for initial development and construction of the project and for ongoing operations (sustaining capital). The resulting net annual cash flows are discounted back to the date of valuation end-of-year 2010 dollars and totalled to determine NPVs at the selected discount rates. The IRR is calculated as the discount rate that yields a zero NPV. The payback period is calculated as the time needed to recover the initial capital spent.

The mine plan is based on Measured and Indicated mineral resources only, with Inferred mineral resources considered as waste. The results of the economic analyses based on these mineral resources are considered to be a Preliminary Assessment (PA) under NI 43-101.




19.9.1 Basis of Base Case Financial Analysis

The base case financial analysis incorporated the following assumptions:

• Conventional truck-and-shovel open pit mining method

• Average LOM process rate of 34.7 Mt/a

• Average LOM copper recovery of 89.7%

• Transport and insurance costs of $781 million over the life of the mine

• To reflect some of the advantage that CCC is expected to gain from the current high price scenario, AMEC used the reverting cost curve values in Table 19-8 to estimate future metal prices for each case.

• Life-of-mine operating costs of $1.24/lb recovered Cu, including smelter and transport costs, for the base case.

Table 19-8: Base Case Metal Prices

Copper Gold Silver Year $/lb $/oz $/oz

2015 2.92 1,159 20.96

2016 2.77 1,132 20.00

2017 2.65 1,111 19.25

2018 2.56 1,094 18.64

2019 2.50 1,080 18.15

2020 2.44 1,069 17.76

2021 2.40 1,060 17.43

2022 2.37 1,052 17.17

2023 2.34 1,046 16.95

2024 2.32 1,041 16.76

2025 2.25 1,015 15.85

2026 2.25 1,015 15.85

• The capital cost estimate was prepared assuming Owner equipment and Owner operation of the equipment. For the purposes of the financial analysis, mine mobile equipment was considered to be leased, rather than purchased.

• Start-up capital costs for the base case of $1.437 billion for the mine construction phase (excluding leased equipment) and $70 million in sustaining capital to replenish mining and other equipment during the life of the mine.




• Royalty assumptions of 1% on the first $60,000, 2% on the next $60,000, and 3% on the remainder. Over the life of the mine, the NPI royalty is estimated to total $330 million.

• Working capital of a maximum of $66 million in the first year of operation

• Decommissioning and reclamation costs of $58 million

• Corporate tax rate of 30% plus an 8% labour profit sharing tax

• No inflationary adjustments

• No salvage value.

19.9.2 Results of Base Case Financial Analysis

The pre-tax cumulative cash flow for the base case is $5,998 million with an IRR of 26.8%. The cash flow analysis shows that the project will generate a positive cash flow in all years except Year -2 and Year -1 on a pre-tax basis. The annual positive cash flow results in a payback period of 2.9 years. At an 8% discount rate, the pre-tax net present value (NPV) of the project is $1,983 million.

The after-tax cumulative cash flow for the base case is $3,855 million with an IRR of 18.8%. The cash flow analysis shows that the project will generate a positive cash flow in all years except Year -2 and Year -1 on an after-tax basis. The annual positive cash flow results in a payback period of 4.1 years. At an 8% discount rate, the after-tax net present value (NPV) of the project is $1,063 million.

The cashflow, on an annualized basis, for the base case is included as Table 19-9.

19.9.3 Base Case Sensitivity Analysis

Sensitivity analysis was performed on the base case taking into account variations in metal prices, operating costs and capital costs. For the purposes of the analysis, variations in grade were considered to mirror the variations in metal price.

The results from the analysis showed that the Project sensitivity was (in order from highest to lowest) metal price, operating expenditure, capital expenditure. These are shown in Table 19-10 and Figure 19-6.

Considering that the Project is priced in US dollars, the effects of exchange rate variation do not apply in the current model, although in reality some equipment, supplies, and services will be priced in Euros or the local Peruvian currency.




Table 19-9: Base Case Cashflow Analysis

Project year Life of Mine 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25

Production year -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22

Metal prices Gold US$/oz 1,159 1,132 1,111 1,094 1,080 1,069 1,060 1,052 1,046 1,041 1,015 1,015 1,015 1,015 1,015 1,015 1,015 1,015 1,015 1,015 1,015 1,015 Silver US$/oz 20.96 20.00 19.25 18.64 18.15 17.76 17.43 17.17 16.95 16.76 15.85 15.85 15.85 15.85 15.85 15.85 15.85 15.85 15.85 15.85 15.85 15.85 Copper US$/lb 2.92 2.77 2.65 2.56 2.50 2.44 2.40 2.37 2.34 2.32 2.25 2.25 2.25 2.25 2.25 2.25 2.25 2.25 2.25 2.25 2.25 2.25 Pay Metal Gold Kozs 791 37.70 52.03 47.74 36.18 43.39 44.25 46.33 46.03 41.12 30.39 23.46 29.53 33.77 38.50 40.79 39.06 27.23 21.58 26.41 37.16 40.68 7.68 Silver Kozs 18,014 976.39 1,069.81 918.38 977.62 1,015.86 983.14 912.68 838.83 759.61 663.46 691.29 799.24 849.12 862.40 882.95 815.92 634.23 648.94 753.14 880.13 880.15 200.25 Copper Klbs 5,555,529 253,064 336,744 295,653 259,975 304,860 281,809 282,018 273,986 247,269 200,857 196,419 237,031 260,894 275,006 279,781 269,536 208,298 193,218 235,688 287,149 309,493 66,780 Pay Value Gold US$000 832,910 43,688 58,910 53,045 39,585 46,867 47,302 49,105 48,432 43,018 31,629 23,815 29,972 34,272 39,082 41,404 39,645 27,638 21,901 26,807 37,717 41,286 7,790 Silver US$000 308,989 20,461 21,399 17,676 18,224 18,441 17,456 15,909 14,399 12,872 11,122 10,957 12,668 13,459 13,669 13,995 12,932 10,052 10,286 11,937 13,950 13,950 3,174 Copper US$000 13,289,674 739,664 932,530 784,407 666,692 760,981 688,536 677,430 649,317 579,758 466,952 441,942 533,320 587,011 618,764 629,507 606,456 468,670 434,740 530,298 646,086 696,360 150,255 Total US$000 14,431,573 803,814 1,012,839 855,127 724,501 826,288 753,294 742,443 712,148 635,648 509,703 476,714 575,960 634,741 671,515 684,906 659,033 506,360 466,927 569,042 697,753 751,597 161,220 Concentrate transport Land freight US$000 (192,019) (8,747) (11,639) (10,219) (8,986) (10,537) (9,740) (9,748) (9,470) (8,547) (6,942) (6,789) (8,193) (9,017) (9,505) (9,670) (9,316) (7,200) (6,678) (8,146) (9,925) (10,697) (2,308) Port storage & handling US$000 (110,274) (5,023) (6,684) (5,869) (5,160) (6,051) (5,594) (5,598) (5,438) (4,908) (3,987) (3,899) (4,705) (5,179) (5,459) (5,553) (5,350) (4,135) (3,835) (4,678) (5,700) (6,143) (1,326) Ocean freight US$000 (457,188) (20,826) (27,712) (24,331) (21,394) (25,088) (23,191) (23,208) (22,547) (20,349) (16,529) (16,164) (19,506) (21,470) (22,631) (23,024) (22,181) (17,142) (15,901) (19,396) (23,631) (25,470) (5,496) Insurance charges US$000 (21,647) (1,206) (1,519) (1,283) (1,087) (1,239) (1,130) (1,114) (1,068) (953) (765) (715) (864) (952) (1,007) (1,027) (989) (760) (700) (854) (1,047) (1,127) (242) Total US$000 (781,128) (35,801) (47,555) (41,701) (36,627) (42,916) (39,655) (39,667) (38,524) (34,757) (28,223) (27,567) (33,268) (36,618) (38,603) (39,275) (37,836) (29,235) (27,115) (33,074) (40,302) (43,437) (9,371) Smelting Treatment US$000 (640,244) (29,164) (38,808) (34,072) (29,961) (35,133) (32,477) (32,501) (31,575) (28,496) (23,148) (22,636) (27,317) (30,067) (31,693) (32,243) (31,063) (24,005) (22,267) (27,162) (33,092) (35,667) (7,696) Cu refining US$000 (416,665) (18,980) (25,256) (22,174) (19,498) (22,865) (21,136) (21,151) (20,549) (18,545) (15,064) (14,731) (17,777) (19,567) (20,625) (20,984) (20,215) (15,622) (14,491) (17,677) (21,536) (23,212) (5,009) Au refining US$000 (3,955) (188) (260) (239) (181) (217) (221) (232) (230) (206) (152) (117) (148) (169) (193) (204) (195) (136) (108) (132) (186) (203) (38) Ag refining US$000 (7,205) (391) (428) (367) (391) (406) (393) (365) (336) (304) (265) (277) (320) (340) (345) (353) (326) (254) (260) (301) (352) (352) (80) Total US$000 (1,068,069) (48,723) (64,752) (56,852) (50,031) (58,621) (54,227) (54,249) (52,690) (47,551) (38,629) (37,761) (45,561) (50,142) (52,856) (53,784) (51,799) (40,017) (37,126) (45,272) (55,166) (59,435) (12,823) Net Smelter Return US$000 12,582,376 719,289 900,533 756,574 637,843 724,751 659,412 648,527 620,934 553,340 442,851 411,386 497,131 547,981 580,057 591,847 569,397 437,108 402,686 490,696 602,284 648,724 139,025 Acid Revenue US$000 269,306 12,267 16,324 14,332 12,602 14,778 13,661 13,671 13,282 11,986 9,737 9,521 11,490 12,647 13,331 13,562 13,066 10,097 9,366 11,425 13,920 15,003 3,237 NSR Plus Acid Revenue US$000 12,851,683 731,557 916,856 770,906 650,445 739,529 673,073 662,198 634,215 565,327 452,587 420,907 508,621 560,627 593,388 605,409 582,463 447,205 412,052 502,121 616,204 663,727 142,263 Production costs Mining US$000 (1,995,248) (94,354) (116,781) (98,501) (98,888) (114,731) (100,459) (101,503) (113,388) (101,481) (94,191) (95,615) (91,220) (88,101) (88,610) (106,956) (96,632) (83,925) (71,582) (69,094) (68,598) (72,446) (28,192) Process US$000 (2,265,380) (86,296) (107,870) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (107,869) (21,695) G&A US$000 (311,342) (11,860) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (14,825) (2,982) Acid US$000 (62,838) (2,862) (3,809) (3,344) (2,941) (3,448) (3,188) (3,190) (3,099) (2,797) (2,272) (2,222) (2,681) (2,951) (3,111) (3,165) (3,049) (2,356) (2,185) (2,666) (3,248) (3,501) (755) Lease Payment - Mining US$000 (316,092) (31,842) (38,104) (37,069) (37,051) (37,051) (9,054) (682) (854) (1,171) (2,003) (2,325) (1,928) (2,330) (21,271) (21,904) (21,843) (21,674) (21,522) (4,405) (1,143) (433) (433) Total US$000 (4,950,902) (227,214) (281,389) (261,608) (261,574) (277,925) (235,394) (228,070) (240,035) (228,143) (221,161) (222,856) (218,524) (216,077) (235,686) (254,720) (244,218) (230,649) (217,984) (198,860) (195,683) (199,074) (54,057) Royalty payment Net Smelter Return US$000 12,582,376 719,289 900,533 756,574 637,843 724,751 659,412 648,527 620,934 553,340 442,851 411,386 497,131 547,981 580,057 591,847 569,397 437,108 402,686 490,696 602,284 648,724 139,025 Stage 1 royalty US$000 (125,824) (7,193) (9,005) (7,566) (6,378) (7,248) (6,594) (6,485) (6,209) (5,533) (4,429) (4,114) (4,971) (5,480) (5,801) (5,918) (5,694) (4,371) (4,027) (4,907) (6,023) (6,487) (1,390) Stage 2 royalty US$000 (112,624) (6,593) (8,405) (6,966) (5,778) (6,648) (5,994) (5,885) (5,609) (4,933) (3,829) (3,514) (4,371) (4,880) (5,201) (5,318) (5,094) (3,771) (3,427) (4,307) (5,423) (5,887) (790) Stage 3 royalty US$000 (99,424) (5,993) (7,805) (6,366) (5,178) (6,048) (5,394) (5,285) (5,009) (4,333) (3,229) (2,914) (3,771) (4,280) (4,601) (4,718) (4,494) (3,171) (2,827) (3,707) (4,823) (5,287) (190) Net profits royalty US$000 (337,871) (19,779) (25,216) (20,897) (17,335) (19,943) (17,982) (17,656) (16,828) (14,800) (11,486) (10,542) (13,114) (14,639) (15,602) (15,955) (15,282) (11,313) (10,281) (12,921) (16,269) (17,662) (2,371) Closure & salvage Closure costs US$000 (57,650) (57,650) Total US$000 (57,650) (57,650) Earnings Earnings before taxes, depreciation & amortization US$000 7,505,260 484,564 610,251 488,402 371,536 441,662 419,696 416,472 377,352 322,383 219,941 187,509 276,983 329,912 342,100 334,734 322,963 205,243 183,787 290,340 404,252 446,991 28,185




Project year Life of Mine 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22 23 24 25

Production year -3 -2 -1 1 2 3 4 5 6 7 8 9 10 11 12 13 14 15 16 17 18 19 20 21 22

Deductible interest US$000 Earnings before taxes, depreciation & amortization US$000 7,505,260 484,564 610,251 488,402 371,536 441,662 419,696 416,472 377,352 322,383 219,941 187,509 276,983 329,912 342,100 334,734 322,963 205,243 183,787 290,340 404,252 446,991 28,185 Taxation Earnings before taxes US$000 7,505,260 484,564 610,251 488,402 371,536 441,662 419,696 416,472 377,352 322,383 219,941 187,509 276,983 329,912 342,100 334,734 322,963 205,243 183,787 290,340 404,252 446,991 28,185 Deductions Depreciations US$000 (1,487,717) (200,436) (139,740) (140,126) (140,216) (102,619) (46,785) (45,656) (45,871) (46,107) (45,743) (45,430) (45,329) (45,309) (46,217) (46,385) (46,911) (46,959) (46,805) (46,496) (46,469) (45,742) (26,365) Total Deductions (200,436) (139,740) (140,126) (140,216) (102,619) (46,785) (45,656) (45,871) (46,107) (45,743) (45,430) (45,329) (45,309) (46,217) (46,385) (46,911) (46,959) (46,805) (46,496) (46,469) (45,742) (26,365) Net taxable income US$000 6,017,543 284,128 470,511 348,275 231,320 339,042 372,911 370,816 331,481 276,276 174,198 142,079 231,655 284,602 295,883 288,349 276,052 158,284 136,983 243,845 357,783 401,250 1,820 Labour Profit Sharing Tax US$000 (481,403) (22,730) (37,641) (27,862) (18,506) (27,123) (29,833) (29,665) (26,518) (22,102) (13,936) (11,366) (18,532) (22,768) (23,671) (23,068) (22,084) (12,663) (10,959) (19,508) (28,623) (32,100) (146) Taxable Income subject to Corporate Tax US$000 5,536,139 261,398 432,870 320,413 212,814 311,919 343,078 341,150 304,963 254,174 160,263 130,713 213,122 261,834 272,212 265,281 253,968 145,621 126,024 224,337 329,160 369,150 1,675 Corporate income tax US$000 (1,660,842) (78,419) (129,861) (96,124) (63,844) (93,576) (102,923) (102,345) (91,489) (76,252) (48,079) (39,214) (63,937) (78,550) (81,664) (79,584) (76,190) (43,686) (37,807) (67,301) (98,748) (110,745) (502) Total tax paid US$000 (2,142,245) (101,150) (167,502) (123,986) (82,350) (120,699) (132,756) (132,010) (118,007) (98,354) (62,015) (50,580) (82,469) (101,318) (105,334) (102,652) (98,274) (56,349) (48,766) (86,809) (127,371) (142,845) (648) Net earnings US$000 5,363,014 383,415 442,749 364,416 289,186 320,963 286,940 284,461 259,345 224,029 157,926 136,929 194,514 228,593 236,766 232,082 224,689 148,894 135,022 203,532 276,881 304,147 27,537 Capital expenditure Construction capital US$000 (1,437,159) (606,908) (818,909) (11,342) Sustaining capital US$000 (70,438) (7,460) (1,977) (4,746) (4,453) (5,804) (1,975) (1,722) (861) (2,499) (4,191) (1,865) (19,217) (3,358) (5,770) (4,089) (451) Working capital US$000 0 (65,754) (16,482) 6,409 1,277 (5,660) 11,448 1,828 (2,705) 3,915 3,379 (260) (342) (226) (5,399) (4,927) 2,985 5,542 3,696 3,291 (1,013) (1,631) 60,628 Debt drawdown US$000 0 Debt repayment US$000 0 Total US$000 (1,507,597) (606,908) (818,909) (77,096) (23,942) 4,431 (3,469) (10,113) 5,643 (147) (4,427) 3,054 880 (4,451) (2,207) (19,443) (8,756) (10,697) (1,104) 5,092 3,696 3,291 (1,013) (1,631) 60,628 Net project cash flow Pre-Tax US$000 5,997,662 (606,908) (818,909) 407,468 586,310 492,833 368,067 431,549 425,340 416,324 372,925 325,437 220,821 183,059 274,776 310,469 333,344 324,037 321,859 210,334 187,484 293,632 403,239 445,360 88,813 After tax US$000 3,855,417 (606,908) (818,909) 306,319 418,808 368,847 285,717 310,850 292,583 284,314 254,917 227,083 158,807 132,478 192,307 209,150 228,010 221,385 223,585 153,985 138,718 206,823 275,869 302,515 88,165




Table 19-10: Results of Sensitivity Analysis for Base Case with NPV @ 8%

Change in Factor Factor -30% -20% -10% 0% 10% 20% 30%

Exchange rate 932 932 932 932 932 932 932 Capital expenditure 1,323 1,193 1,062 932 802 671 541 Operating expenditure 1,312 1,185 1,059 932 806 679 552 Metal price (195) 182 557 932 1,307 1,681 2,056

Figure 19-7: Graph of Sensitivity Analysis Results for Base Case

(500)

0

500

1,000

1,500

2,000

2,500

-40% -30% -20% -10% 0% 10% 20% 30% 40%

NPV

@ 8

% (U

S$ m

illio

n)

Change in Factor

Candente - Sensitivity of NPV @ 8%

Capex

Opex

Price

Xrate

19.9.4 Alternative Cases

Three alternative cases were evaluated. These used the parameters of the base case with the exception of the capital costs. The other three scenarios included:

• Full Owner-operated mine and process facilities

• Contractor mining

• Full lease of process and electrical equipment in addition to mining equipment.

The capital cost for the full Owner-operated and financed option is $1.599 billion for start-up and $206 million to sustain the operation. If the contract mining option were considered, then the cost could be reduced to $1.405 billion and $63 million for start-up and sustaining capital, respectively. The option involving full lease of mine and




major process and electrical equipment reduces to $1.191 billion and $207 million. Results of the evaluation of the three alternative cases are summarized in Table 19-11.

Table 19-11: Results of Financial Analysis for Alternative Cases

Item Unit Value at $2.25/lb Cu Value at $3/lb Cu Full Owner Operated – pre-tax

IRR % 25.8 35.8

CNCF* C$000,000 6,017 9,842

NPV 8% C$000,000 1,968 3,504

NPV 10% C$000,000 1,501 2,776

Payback Years 3.0 2.3 Contract Mining – pre-tax

IRR % 27.0 37.8

CNCF* C$000,000 5,754 9,579

NPV 8% C$000,000 1,919 3,456

NPV 10% C$000,000 1,476 2,751

Payback Years 2.8 2.1 Equipment Leasing – pre-tax

IRR % 29.9 42.0

CNCF* C$000,000 5,847 9,672

NPV 8% C$000,000 1,989 3,525

NPV 10% C$000,000 1,547 2,822

Payback Years 2.6 1.9 *Cumulative Net Cash Flow

19.10 Risks and Opportunities

A risk and opportunity analysis was performed.

Key risks identified included:

• Geotechnical design of +900 m pit highwall: The preliminary pit highwall is over 900 m in vertical height, making the proposed Cañariaco Norte pit one of the world’s deepest. The current study design basis for the overall pit and mine bench slope angles is pending completion of a geotechnical drilling program to assess rock quality in the pit wall areas. The geotechnical program is planned to proceed upon receipt of drilling permits. The current design is based on assessment of scoping level information, visual inspection of available core, and limited physical testing of available material. If future geotechnical investigations result in shallower overall pit slope angles, then the stripping ratio will increase and have a negative impact on the project mining costs.




• Obtaining permits for Project: Prior to development, the project will require many applications for permits to allow for timely execution of Project work. Receipt of the necessary permits for Project development will depend on completion of an Environmental and Social Impact Assessment that is acceptable to the relevant regulatory agencies in Peru and negotiation of surface access rights for the project area from the local community. A comprehensive ESIA for the Cañariaco project is currently underway under the guidance of AMEC Earth and Environmental. AMEC notes that the time period required for negotiation of drilling permits with the local community has taken longer than anticipated by CCC. Other applications could potentially encounter similar delays. By continuing to work closely with the local communities, CCC is confident that this risk can be managed.

• Main access road routing: The design of the main access road is based on low-quality contour data. This may lead to a less-than-optimum design with a risk that the road may not be constructible in some sections. Aerial or land reconnaissance followed by more detailed geotechnical investigations along the proposed road routing is recommended to decrease the risk in the design of this road.

• Landslides and diversion channel: Naturally-occurring landslides were observed to have taken place in the vicinity of the proposed diversion channel. The functionality of the channel may be at risk. Further assessment of the valley slopes will be required to assess the severity of this risk and whether it can be mitigated through regular maintenance of the channel.

• Width of tailings line access road: The narrowness of the proposed road may cause issues during the construction of the pipeline by restricting access to the site, since the equipment necessary to install the pipeline will block the road for extended periods of time. Similarly, during operation, repair, and maintenance of the pipeline there may be restricted access to and from the plant site. Road movements will need to be scheduled and prioritized for vehicular access.

• Coarse Ore Stockpile and Dusting: The open-air design of the coarse ore stockpile is typical of many operating mines however it may result in non-compliant dust loading during dry periods. Dust suppression or a stockpile cover may be required.

• Stockpile feed conveyor incline angle: The design incline angle of the stockpile feed conveyor is 14°. Testwork is required to confirm that this design angle is appropriate for this material.

• Delivery of long-lead capital equipment: Market conditions may increase delivery schedules for long-lead capital equipment items. Equipment may need to be secured and purchased earlier than anticipated.




• Port availability: Current assumptions are that the marine facilities to be developed by Lumina will be for Phase 2 development, i.e., fully loaded vessel sizes of up to 50,000 DWT (instead of 33,000 DWT) and a berth in 15 m water depth. This should be confirmed in negotiations with Lumina, since acceptable berth availabilities are only achieved with Phase 2 implementation.

Opportunities recognized, and which will be evaluated during more detailed studies, include:

• 18 month mine plan: The mine plan for the first 18 months of operations should be optimized to maximize the IRR and reduce project payback period.

• Mineral resource estimation: Additional drilling should be undertaken in the vicinity of Cañariaco Norte with the objective of providing sufficient data to support potential upgrading of mineralization that is currently classified as Inferred to higher confidence categories. The potential exists to discover and delineate additional mineralization at the Cañariaco Sur and Quebrada Verde targets

• Leasing of equipment: The possibility of leasing equipment for the mine, mill, and electrical infrastructure should be evaluated to determine if it may improve the project economics.

• Low-grade material recoveries: Testwork is required to evaluate process recoveries for low-grade material. Further testwork and optimization of process parameters could result in higher recoveries.

• Concentrate copper grade: The target grade for the copper concentrate should be assessed to optimize the overall costs related to handling, transportation and treatment.

• Final tailings thickener: A trade-off study is required to determine if overall tailings and reclaim water handling costs may be reduced by elimination of the tailings thickener from the process flowsheet.

• Concentrate roasting requirements: The arsenic distribution in the resource should be evaluated in more detail and optimized in the mine plan to reduce roasting associated costs.

• Concentrate shipping costs: The potential to reduce land shipping costs from the plant to the port by transporting dry concentrate should be evaluated.

• Mineralized material hardness: The mill tonnage throughput should be optimized through a mineralization hardness model in the block model.

• Tailings cyclone system: The tailings cyclone system was selected based on a preliminary level trade-off assessment comparing rock fill and cyclone sands dam




construction. A more in depth trade-off assessment is recommended to confirm the dam construction methodology that would incur the lowest life of mine costs. .

• Stockpile feed conveyor: A trade-off study should be completed to evaluate a single stockpile feed conveyor versus splitting it into two sections. Having a short conveyor feeding the stockpile will reduce the amount of structural steel required for the cantilevered head because the shorter conveyor will transmit lower forces to the structure.

• Coarse ore conveyor motor: A trade-off study should be completed for a wound rotor motor versus a wrap-around motor on the coarse ore stockpile conveyor.




20.0 INTERPRETATION AND CONCLUSIONS

The QPs, as authors of this Report, have reviewed the data for the Project and are of the opinion that:

• Legal opinion provided to AMEC indicates that the mining tenure held by CCC in the Project area is valid, and sufficient to support declaration of Mineral Resources.

• Surface rights are held by the Comunidad Campesina San Juan Bautista de Cañaris, and access for a further three-year term for exploration purposes is currently being negotiated. Additional surface rights negotiations will be required to support infrastructure construction and mining operations in the proposed mining area. Negotiations will be required for surface rights for additional lands including road rights-of-way, and powerline facilities.

• Although no metal production royalty is held by any third-party, royalties are payable to the Peruvian Government.

• The permits held by CCC for the Project are sufficient to ensure that exploration activities are conducted within the regulatory framework required by the Peruvian Government. Additional permits will be required for Project development.

• Environmental permits for Project development have to be secured. This process will determine the precise number of environmental management plans that the regulatory authorities will require to ensure compliance with environmental design and permit criteria. A comprehensive ESIA for the Cañariaco Project is currently underway.

• The time period required for negotiation of drilling permits with the local community has taken longer than anticipated by CCC. Other applications could potentially encounter similar delays. By continuing to work closely with the local communities, CCC is confident that this risk can be managed.

• Sufficient work has been undertaken on the permitting and environmental aspects of the project to gain an understanding of the regulatory requirements that will need to be met to construct, operate, and close the mine.

• The existing and planned infrastructure, availability of staff, the existing power, water, and communications facilities, the methods whereby goods are transported to the mine, and any planned modifications or supporting studies are well-established, or the requirements to establish such, are well understood by CCC, and can support the declaration of Mineral Resources.

• The geologic understanding of the deposit settings, lithologies, and structural and alteration controls on mineralization is sufficient to support estimation of Mineral Resources.




• The mineralization style and setting is well understood and can support declaration of Mineral Resources.

• The exploration programs completed to date are appropriate to the porphyry copper style of the deposit.

• Work completed in the period 1967 to date has consisted of geological mapping, prospecting, stream sediment, grab, rock chip and soil geochemical sampling, trenching and pitting, ground IP/resistivity and magnetic geophysical surveys, petrographic studies, bulk sampling for metallurgical testing, re-logging and re-sampling of historic drill core, and core drilling. Completed exploration programs were appropriate to the mineralization style. To date, a deposit and two exploration targets have been identified.

• Sampling methods are acceptable, meet industry-standard practice, and are acceptable for Mineral Resource estimation purposes.

• The quality of the CCC copper, gold, and silver analytical data is reliable and that sample preparation, analysis, and security are generally performed in accordance with exploration best practices and industry standards.

• Historic drill data have been validated by CCC’s re-logging and re-assay programs, and data from these programs are reliable, and can be used to support Mineral Resource estimates.

• The results of ongoing metallurgical testwork indicate that the mineralization responds positively to typical grinding and conventional flotation methodologies.

• Metallurgical testwork to date has shown that the mineralization is amenable to being processed using conventional technologies, and acceptable recoveries were returned. Metallurgical testwork completed on the Project has been appropriate to establish process routes that are applicable to the mineralization types and was performed on samples that were representative of the mineralization.

• A concentrator flowsheet was developed that appears to be viable with respect to the known mineralization of the deposit. The Outotec partial roast process has been confirmed to reduce the levels of impurities in the concentrate to below smelter penalty limits and will produce a saleable concentrate throughout the mine life.

• Mineral Resources, which were estimated using core drill data, have been performed to industry best practices, and conform to the requirements of CIM Definition Standards (2005).

• Reviews of the environmental, permitting, legal, title, taxation, socio-economic, marketing and political factors and constraints for the Project support the declaration of Mineral Resources using the set of assumptions outlined. Mineral




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

• The proposed open pit mining method is appropriate to the style of mineralization

− Mining will proceed using a conventional truck-and-shovel fleet − Production forecasts are achievable with the proposed equipment and plant − The predicted mine life of 22 years is achievable based on the projected

annual production rate and the Measured and Indicated Mineral Resources estimated

− There is some upside for the Project if the Inferred Mineral Resources that are identified within the LOM production plan can be upgraded to higher confidence Mineral Resource categories. Additional potential exists if significant mineralization can be identified within the adjacent Cañariaco Sur and Quebrada Verde targets.

• The open pit mine design relating to the angle of the pit wall slopes has been developed based on geotechnical logging of drill core plus rock quality evaluation and compressive strength testing of a limited number of core samples. The amount of geotechnical data available is not sufficient, however, to support development of a pre-feasibility level design for the pit wall slopes. As a result, the mine plan developed for this Report must be considered to be at a preliminary level.

• The preliminary pit highwall is over 900 m in vertical height, making the Cañariaco Norte pit one of the world’s deepest. If future geotechnical investigations result in shallower overall pit slope angles, then the stripping ratio will increase and have a negative impact on the project mining costs.

• AMEC does not provide detailed taxation analysis, but has attempted to apply the relevant allowances, deductions, and taxes to the maximum advantage of CCC based on other projects in Peru. Additional considerations on the likely taxation regime are required during more detailed studies.




21.0 RECOMMENDATIONS

The following work is recommended; no phase is contingent on another, and all can be conducted concurrently. CCC should:

• Initiate insertion of fine pulp duplicate samples for precision analysis, as this is currently not part of the CCC sampling program. Such checks should be instituted in all future exploration programs. AMEC estimates that this would add approximately 5% to the total costs of the sample analytical programs.

• Construct a relational database using a commercial database software package, for more detailed studies. Costs of such a program would depend on whether the database was constructed in-house or a third-party consultant was contracted to complete the exercise. Estimated cost: US$10,000.

• Drill 11 geotechnical core holes (6,690 m) to provide additional information in the proposed pit area on rock mass properties, joint patterns, hydrogeology and geological modeling/resource estimation. Estimated cost: US$1.1 M.

• Drill additional holes to increase the level of confidence of the lithological interpretation in areas outside the existing drill hole limits but constrained by the resource pit. The proposed geotechnical holes will cover part of the recommendation; however an additional 10 holes (6,000 m) is suggested in addition. Estimated cost: US$900,000 to US$950,000.

• Complete a full review of the alteration interpretation on vertical sections reconciled to bench plans. Construct a mineral zonation model for the next phase of study. Evaluate arsenic distribution using the interpreted units. Estimated cost: US$20,0000 to US$350,000.

• Institute a continuous program of specific gravity determinations from core samples preferably using the same laboratory and determination procedures as those used in the 2010 testing program. AMEC estimates that this would add approximately 2–5% to the total analytical program costs.

The total cost of the recommendations ranges from US$2.21 million to US$2.41 million.

At the completion of this work, results should be reviewed and incorporated as appropriate into the ongoing preliminary feasibility study or future detailed studies.




22.0 REFERENCES

AMEC, 2011: Prefeasibility Progress Report: unpublished internal report prepared by AMEC E&C Services Inc., for Candente Copper Corporation, 24 February 2011.

Berger, R.B., Ayuso, R.A., Wynn, J.C., and Seal, R.R., 2008: Preliminary Model Of Porphyry Copper Deposits: U.S. Geological Survey Open-File Report 2008–1321, 55 p.

Bonson, C., Nowak, M., Doerksen, G., Johnston, A., and van Egmond, R., 2008: Technical Report Cañariaco Norte Project, Department of Lambayeque, Peru: technical report prepared by SRK Consulting Ltd for Candente Resource Corp., effective date 11 July 2008

Bonson, C., Campbell, R., Bender, M., Doerksen, G., Johnston, A., Meyer, T., Nowak, M., Pilotto, D., Van Egmond, R., Critikos, P., Ostolaza, R., and Huanani, A., 2008: Revised Preliminary Economic Assessment Technical Report, Cañariaco Norte Project, Peru: technical report prepared by SRK Consulting Ltd for Candente Resource Corp., effective date 30 November 2008

Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2000: CIM Standards for Mineral Resources and Mineral Reserves, Definitions and Guidelines: Canadian Institute of Mining, Metallurgy and Petroleum, August, 2000 http://www.jogmec.go.jp/mric_web/tani/cimstandard.pdf

Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2003: Estimation of Mineral Resources and Mineral Reserves, Best Practice Guidelines: Canadian Institute of Mining, Metallurgy and Petroleum, November 23, 2003, http://www.cim.org/committees/estimation2003.pdf.

Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2005: CIM Standards for Mineral Resources and Mineral Reserves, Definitions and Guidelines: Canadian Institute of Mining, Metallurgy and Petroleum, December 2005, http://www.cim.org/committees/CIMDefStds_Dec11_05.pdf.

Canadian Securities Administrators (CSA), 2005: National Instrument 43-101, Standards of Disclosure for Mineral Projects, Canadian Securities Administrators.

Collins, J., McCrea, J., and Rokosh, J., 2006: Cañariaco Copper Project, Peru, Preliminary Assessment and Economic Evaluation Report: technical report prepared by Merit Consultants for Candente Resource Corp., effective date 28 June 2006




Currie, J.A., 2004: Technical Report On The Cañariaco Copper Porphyry Prospect Department Of Lambayeque Northern Perú Latitude 6° 05’ S Longitude 70° 17’ E, technical report prepared for Candente Resource Corp., August 2004

Foulkes, J., 2010: The Cañariaco Norte Copper Project, December 2010 Update: unpublished PowerPoint presentation, Candente Resource Corp., 3 December 2010.

Huanqui, F., Freeze, J.C., and Coder, J.M., 2002: Geological Report On The Cañariaco Copper-Gold Porphyry Prospect Department Of Lambayeque Northwest Perú NTS 13-E 70° 17’ E 6° 05’ S: technical report prepared for Candente Resource Corp., effective date 9 August 2002

Journel, A., G., and Huijbregts, Ch., J., 1978: Chapter 5: Selection and Estimation of Reserves: in Mining Geostatistics p. 472-490, Blackburn Press, 600 p.

McCrea, J.A., 2005: Technical Report on the Cañariaco Copper Porphyry Prospect Department of Lambayeque Northwest Peru: technical report prepared for Candente Resource Corp., effective date 22 April 2005

McCrea, J.A., 2006: Technical Report on the Cañariaco Copper Porphyry Prospect Department of Lambayeque Northwest Peru: technical report prepared for Candente Resource Corp., effective date 27 April 2006

Sillitoe, R.H., 2000: Role of Gold-Rich Porphyry Models in Exploration, in S.G. Hagerman and P.H. Brown, eds., Gold in 2000, Reviews in Economic Geology, v. 13, pp. 311–346.

Sillitoe, R.H., 1999: Comments on Geology and Potential of the Cañariaco Porphyry Copper Prospect, Northern Perú: Unpublished internal geological report for prepared for Billiton Exploration and Mining Perú B.V., 10 p.

Sillitoe, R.H., 2010: Porphyry Copper Systems: Economic Geology, v. 105, pp. 3–41.

Sinclair, W.D., 2006: Consolidation and Synthesis of Mineral Deposits Knowledge - Porphyry Deposits: report posted to Natural Resources Canada website 30 January 2006, 14 p., <http://gsc.nrcan.gc.ca/mindep/synth_dep/porph/index_e.php>, accessed 28 August 2010

Stone, D.M.R., Godden, S., Van Egmond, R., and Tosney, J.R., 2007: Updated Technical Report and Preliminary Economic Assessment on the Cañariaco Norté Porphyry Copper Project Starter Pit Option: technical report prepared




by Minefill Services Inc. for Candente Resource Corp., effective date 18 April 2007

Waller, S., and Freeze, J., 2010: The Cañariaco Norte Copper Project, December 2010 Update: unpublished PowerPoint presentation, Candente Resource Corp., 3 December 2010.

Wilson, A.J., 1999: Cañariaco Project, Lambayeque, Northern Perú: Unpublished internal geologic report prepared for Billiton Exploration and Mining Perú B.V




23.0 DATE AND SIGNATURE PAGE

The effective date of this Technical Report, entitled “Candente Copper Corporation, Cañariaco Project, Lambayeque Department, Peru, NI 43-101 Technical Report on Pre-feasibility Study Progress Report” is 18 January 2011.

“signed and sealed”

David G. Thomas, P.Geo., MAusIMM

dated: 4 March 2011


Jay Melnyk, P.Eng.

dated: 4 March 2011


Tony Lipiec, P.Eng.

dated: 4 March 2011


Alexandra Kozak, P.Eng.

dated: 4 March 2011

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