EX-99.1 2 ex99-1.htm EXHIBIT 99.1
IMPORTANT NOTICE
Recognizing that Bema Gold Corporation (Bema) has legal and regulatory obligations in a number of global jurisdictions, AMEC E&C Services (AMEC) consents to the filing of this report with any stock exchange and other regulatory authority and any publication by Bema, including electronic publication on Bema’s website accessible by the public, of this report.
This report was prepared as a National Instrument 43-101 Technical Report, in accordance with Form 43-101F1, for Bema by AMEC. This report has been readdressed to Kinross Gold Corporation (Kinross) as a result of Kinross’ acquisition of Bema on February 27, 2007. 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 to be used by Bema and Kinross, subject to the terms and conditions of its contract with AMEC. That contract permits Bema and Kinross to file this report as a Technical Report with Canadian Securities Regulatory Authorities pursuant to provincial securities legislation. Except for the purposes legislated under provincial securities laws, any other use of this report by any third party is at that party’s sole risk.

CERTIFICATE OF AUTHOR
Larry B. Smith, R. Geo, C. P. Geo
6575 Stone Valley Drive
Reno, Nevada 89523
Tel: (775) 331-2375
Fax: (775) 331-4153
larry.smith@amec.com
I, Larry B. Smith, R. Geo, C.P. Geo., am a Registered Geologist and Chartered Professional Geologist, and Manager of Mining & Metals Consulting of AMEC E&C Services, Inc. at 780 Vista Boulevard, Suite 100 in Sparks, Nevada 89434. I have been employed by AMEC since February, 1998.
I am registered as a Professional Geologist in the state of Wyoming (PG-324), am a Fellow and Chartered Professional Geologist in the Australasian Institute of Mining and Metallurgy (Registration number 209301) and am a Certified Professional Geologist with the American Institute of Professional Geologists (CPG-10313). I graduated from Boise State University with a Bachelor of Science in geology in 1972 and subsequently obtained a Master of Science degree in Economic Geology from the Colorado School of Mines in 1982.
I have practiced my profession continuously since 1972 and have been involved in: mineral exploration for uranium, copper, gold, silver, nickel, lead, zinc, and industrial minerals in the United States, Canada, Mexico and Central America; exploration data evaluation, geological modeling and resource modeling of gold, copper, iron, manganese and industrial mineral deposits in the United States, Canada, Australia, Colombia, Chile, Bolivia, Brazil, Greenland, Bosnia and Niger.
As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.
I am currently a Consulting Geologist and have been so since February 1998.
I am responsible for the preparation of Sections 1, 3, 6 to 16, and 20 to 22 of the technical report titled Technical Report, Cerro Casale Project, Chile and dated 06 August 2006 (the “Technical Report”) relating to the Cerro Casale Project in Region Three of northern Chile. I was the Qualified Person responsible for the preparation of the previous technical report on the Cerro Casale Project, which was dated 22 March 2005 and which was filed with the Canadian securities regulators on 24 March 2005. As part of preparation of this report, I visited the Cerro Casale Project on January 12 and 13, 2005 and reviewed exploration programs, exploration data, geological models, core and reverse-circulation sampling practices, sample preparation, assaying, resource estimates and reserve estimates for the purpose of preparing a technical report on all mining operations and mineral resources and reserves within the joint venture. No additional exploration or site work has occurred on the project since my site visit in 2005.
The subject of this update to the Technical Report are revisions in mine designs, process designs, operating costs and capital costs that were the result of work by Bema and Mine and Quarry Engineering Services. In preparation of this update to the Technical Report I was assisted in review of mine planning, cost estimating and financial analyses by William A. Tilley, P.E, an AMEC employee and Qualified Person in the area of mining engineering. Bill was assisted in review of mine designs by Mark Hertel, an AMEC employee, in review of capital costs by Manuel Romero, an AMEC employee, and in review of cash flows and financial analyses by Simon Handelsman, Associate Financial Analyst. I was also assisted in review of new metallurgical testwork, revised process designs and concepts, and process operating costs by Jerry Jergensen, Associate Metallurgist.
I have had no other prior involvement with the property that is the subject of the Technical Report.

AMEC E&C Services, Inc. 780 Vista Boulevard, Suite 100 Sparks, Nevada 89434 Tel +1 775 331 2375 Fax +1 775 331 4153 |  | www.amec.com |  |

William A. Tilley is registered as an Engineer/Mining in the state of Arizona (Registration Number 32391). He graduated with a degree in Bachelor of Science in Mining Engineering from Montana College of Mineral Science and Technology in 1988. He has practiced his profession continuously since 1988 and has been involved in project evaluation for copper, gold, silver, uranium, and industrial minerals in the United States, Canada, Mexico, South America, Europe, and South Africa.
I certify that, to the best of my personal knowledge, information and belief, that the technical report contains all scientific and technical information required to be disclosed to make the report not misleading.
I and William A. Tilley are independent of Bema Gold Corporation, Arizona Star Resources and Kinross Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.
I have read National Instrument 43-101 and certify that the Technical Report has been prepared in compliance with that Instrument. I further certify that, as of the date of this certificate, the Technical Report contains all of the information required under Form 43-101F1 in respect of the property that is the subject of the report.
Dated at Reno, Nevada, this 21st day of March 2007.

“Signed and Sealed” |  |  |
Larry B. Smith, P.Geo., C.P. Geo. |  |  |

AMEC E&C Services, Inc. 780 Vista Boulevard, Suite 100 Sparks, Nevada 89434 Tel +1 775 331 2375 Fax +1 775 331 4153 |  | www.amec.com |  |

CERTIFICATE OF AUTHOR
William A. Tilley PE
2001 West Camelback Street, Suite 300
Phoenix, Arizona 85015
Tel: (602) 343-2400
Fax: (602) 343-2499
bill.tilley@amec.com
I, William A. Tilley, PE, am a Registered Engineer, and Manager, US Consulting for AMEC E&C Services, Inc. at 2001 West Camelback Street, Suite 300, Phoenix, Arizona 85015. I have been employed by AMEC since February 2004.
I am registered as an Engineer/Mining in the state of Arizona (Registration Number 32391). I graduated with a degree in Bachelor of Science in Mining Engineering from Montana College of Mineral Science and Technology in 1988.
I have practiced my profession continuously since 1988 and have been involved in project evaluation for copper, gold, silver, uranium, and industrial minerals in the United States, Canada, Mexico, South America, Europe, and South Africa.
As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.
I am responsible for the preparation of Sections 2, 4, 5, and 17 to 19 of the technical report titled Technical Report, Cerro Casale Project, Chile and dated 6 August 2006 (the “Technical Report”) relating to the Cerro Casale Project in Region Three of northern Chile. I have not visited the site.
The subject of this update to the Technical Report are revisions in mine designs, process designs, operating costs and capital costs that were the result of work by Bema and Mine and Quarry Engineering Services. I was assisted in review of mine designs by Mark Hertel, an AMEC employee, in review of capital costs by Manuel Romero, an AMEC employee, and in review of cash flows and financial analyses by Simon Handelsman, Associate Financial Analyst.
I have had no other prior involvement with the property that is the subject of the Technical Report.
I certify that, to the best of my personal knowledge, information and belief, that the technical report contains all scientific and technical information required to be disclosed to make the report not misleading.
I am independent of Bema Gold Corporation, Arizona Star Resources and Kinross Gold Corporation in accordance with the application of Section 1.5 of National Instrument 43-101.
I have read National Instrument 43-101 and certify that the Technical Report has been prepared in compliance with that Instrument. I further certify that, as of the date of this certificate, the Technical Report contains all of the information required under Form 43-101F1 in respect of the property that is the subject of the report.
Dated at Phoenix, Arizona, this 21 st day of March 2007.

“Signed and Sealed” |  |  |
William A. Tilley PE |  |  |

AMEC E&C Services, Inc. 780 Vista Boulevard, Suite 100 Sparks, Nevada 89434 Tel +1 775 331 2375 Fax +1 775 331 4153 |  | www.amec.com |  |

Larry B. Smith
AMEC E&C Services, Inc.
6575 Stone Valley Drive
Reno, Nevada 89523
Telephone: 775-331-2375
Fax: 775-331-4153
Email:larry.smith@amec.com
CONSENT of AUTHOR

TO: | British Columbia Securities Commission Alberta Securities Commission Saskatchewan Securities Commission Manitoba Securities Commission Ontario Securities Commission Commission des valeurs mobilieres du Quebec Nunavut Legal Registry Officer of the Administrator, New Brunswick Nova Scotia Securities Commission Registrar of Securities, Prince Edward Island Securities Commission of Newfoundland Registrar of Securities, Government of the Yukon Territories Securities Registry, Government of the Northwest Territories

AND TO: | Kinross Gold Corporation

I, Larry B. Smith, do hereby consent to the filing of the technical report prepared for Bema Gold Corporation and re-addressed herein to Kinross Gold Corporation titled Technical Report, Cerro Casale Project, Chile and dated 06 August 2006 (the “Technical Report”) with the securities regulatory authorities referred to above.
I further consent (a) to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication of the Technical Report by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, and (b) to the publication of the Technical Report by Kinross Gold Corporation on its company website or otherwise.
Dated this 21st day of March 2007.

“Signed and Sealed” |  |  |
Larry B. Smith |  |  |

AMEC E&C Services, Inc. 780 Vista Boulevard, Suite 100 Sparks, Nevada 89434 Tel +1 775 331 2375 Fax +1 775 331 4153 |  | www.amec.com |  |

William A. Tilley PE
AMEC E&C Services, Inc.
2001 West Camelback Street, Suite 300
Phoenix, Arizona 85015
Telephone: 602-343-2400
Fax: 602-343-2400
Email:bill.tilley@amec.com
CONSENT of AUTHOR

TO: | British Columbia Securities Commission Alberta Securities Commission Saskatchewan Securities Commission Manitoba Securities Commission Ontario Securities Commission Commission des valeurs mobilieres du Quebec Nunavut Legal Registry Officer of the Administrator, New Brunswick Nova Scotia Securities Commission Registrar of Securities, Prince Edward Island Securities Commission of Newfoundland Registrar of Securities, Government of the Yukon Territories Securities Registry, Government of the Northwest Territories

AND TO: | Bema Gold Corporation

I, William A. Tilley, do hereby consent to the filing of the technical report prepared for Bema Gold Corporation and re-addressed herein to Kinross Gold Corporation titled Technical Report, Cerro Casale Project, Chile and dated 06 August 2006 (the “Technical Report”) with the securities regulatory authorities referred to above.
I further consent (a) to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication of the Technical Report by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, and (b) to the publication of the Technical Report by Kinross Gold Corporation on its company website or otherwise.
Dated this 21st day of March 2007.

“Signed and Sealed” |  |  |
William A. Tilley PE |  |  |

AMEC E&C Services, Inc. 780 Vista Boulevard, Suite 100 Sparks, Nevada 89434 Tel +1 775 331 2375 Fax +1 775 331 4153 |  | www.amec.com |  |

| K INROSS G OLD C ORPORATION C ERRO C ASALE P ROJECT, C HILE T ECHNICAL R EPORT

C ONTENTS

1.0 | SUMMARY | 1-1

2.0 | INTRODUCTION | 2-1

| 2.1 | Sources of Information | 2-1

| 2.2 | Qualified Persons | 2-1

| 2.3 | Terms of Reference | 2-2

| 2.4 | Units of Measure | 2-2

| 2.4.1 | Common Units | 2-2

| 2.4.2 | Common Chemical Symbols | 2-5

3.0 | RELIANCE ON OTHER EXPERTS | 3-1

4.0 | PROPERTY DESCRIPTION AND LOCATION | 4-1

| 4.1 | Project Ownership and Agreements | 4-1

| 4.2 | Mineral, Surface, and Water Rights | 4-2

| 4.2.1 | Mineral Rights | 4-2

| 4.2.2 | Surface Rights | 4-5

| 4.2.3 | Water Rights | 4-5

| 4.2.4 | Conveyance Rights of Way | 4-6

| 4.3 | Royalties | 4-6

| 4.4 | Other Costs | 4-6

| 4.5 | Environmental Exposures | 4-6

| 4.5.1 | Environmental Approval of Power Supply Infrastructure | 4-7

| 4.5.2 | Environmental Approval of Port Facilities | 4-7

| 4.5.3 | Acid Rock Drainage (ARD) Potential | 4-7

| 4.5.4 | Impacts on Surrounding Water Systems from Water Supply Operations Conducted at the Piedra Pomez Well Field | 4-8

| 4.6 | Environmental Approvals and Permits | 4-8

5.0 | ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY | 5-1

| 5.1 | Accessibility | 5-1

| 5.2 | Climate | 5-1

| 5.3 | Local Resources | 5-3

| 5.4 | Infrastructure | 5-3

| 5.5 | Physiography | 5-4

6.0 | HISTORY | 6-1

7.0 | GEOLOGICAL SETTING | 7-1

| 7.1 | Regional Geology | 7-1

| 7.2 | District Geology | 7-1

| 7.3 | Cerro Casale Deposit Geology | 7-3

| 7.3.1 | Introduction | 7-3

| 7.3.2 | Lithology | 7-3

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| 7.3.3 | Structure | 7-7

| 7.3.4 | Weathering and Oxidation | 7-8

8.0 | DEPOSIT TYPES | 8-1

9.0 | MINERALIZATION | 9-1

| 9.1 | Introduction | 9-1

| 9.2 | Cerro Casale Deposit | 9-1

| 9.2.1 | Alteration | 9-1

| 9.2.2 | Mineralization | 9-3

| 9.3 | Eva Deposit | 9-10

| 9.3.1 | Geology | 9-10

| 9.3.2 | Alteration and Mineralization | 9-13

| 9.4 | Cerro Roman | 9-13

| 9.4.1 | Geology | 9-13

| 9.4.2 | Alteration and Mineralization | 9-13

| 9.5 | Estrella Prospect | 9-16

| 9.5.1 | Geology | 9-16

| 9.5.2 | Alteration and Mineralization | 9-16

| 9.6 | Anfiteatro Prospect | 9-16

| 9.6.1 | Geology | 9-16

| 9.6.2 | Alteration and Mineralization | 9-17

| 9.7 | Romancito Sur | 9-17

| 9.7.1 | Geology | 9-17

| 9.7.2 | Alteration and Mineralization | 9-17

| 9.8 | Other Areas | 9-18

10.0 | EXPLORATION | 10-1

| 10.1 | Cerro Casale | 10-1

| 10.2 | Eva | 10-1

| 10.3 | Cerro Roman | 10-1

| 10.4 | Estrella | 10-2

| 10.5 | Anfiteatro | 10-2

| 10.6 | Romancito | 10-2

| 10.7 | Other Areas | 10-2

11.0 | DRILLING | 11-1

| 11.1 | Drilling Methods | 11-5

| 11.1.1 | Reverse Circulation Drilling | 11-5

| 11.1.2 | Diamond Drilling Equipment | 11-5

| 11.2 | Geological Logging Practices | 11-6

| 11.2.1 | Reverse Circulation Chip Logging | 11-6

| 11.2.2 | Core Logging | 11-7

| 11.2.3 | Geotechnical Logging | 11-7

| 11.3 | AMEC Review of Logging | 11-8

| 11.4 | Core and RC Recovery | 11-8

| 11.5 | Topography | 11-8

| 11.6 | Drill Hole Collar Surveys | 11-9

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| 11.7 | Downhole Surveys | 11-10

12.0 | SAMPLING METHOD AND APPROACH | 12-1

| 12.1 | Reverse-Circulation Drill Sampling | 12-1

| 12.2 | Drill Core Sampling | 12-2

| 12.3 | List of Significant Assays | 12-2

13.0 | SAMPLE PREPARATION, ANALYSES, AND SECURITY | 13-1

| 13.1 | Sample Preparation | 13-1

| 13.1.1 | Reverse-Circulation Samples | 13-1

| 13.1.2 | Core Samples | 13-1

| 13.2 | Assaying | 13-2

| 13.3 | Assay Quality Assurance and Quality Control (QA/QC) | 13-3

| 13.3.1 | On-Site Procedures | 13-3

| 13.3.2 | Assay QA/QC - Pre-1995 | 13-4

| 13.3.3 | Assay QA/QC - 1995 and 1996 | 13-5

| 13.3.4 | Assay QA/QC - 1996 and 1997 | 13-8

| 13.3.5 | Assay QA/QC - 1998 | 13-11

| 13.3.6 | Assay QA/QC - 1999 | 13-23

| 13.4 | Density | 13-33

14.0 | DATA VERIFICATION | 14-1

| 14.1 | Database Development and Integrity Checks | 14-1

| 14.1.1 | Data for 1991 to Early 1996 Drilling Campaigns | 14-1

| 14.1.2 | Data for Late 1996 through 1997 Drilling Campaign | 14-2

| 14.1.3 | Data for 1998 and 1999 Drilling by Placer Dome | 14-2

| 14.2 | AMEC Data Verification | 14-2

| 14.2.1 | Database | 14-2

| 14.2.2 | Geological Interpretations | 14-3

| 14.2.3 | Sampling and Assaying | 14-3

15.0 | ADJACENT PROPERTIES | 15-1

16.0 | MINERAL PROCESSING AND METALLURGICAL TESTING | 16-1

| 16.1 | Scope of Facilities | 16-2

| 16.2 | Design Criteria | 16-3

| 16.3 | Metallurgical Recoveries | 16-4

| 16.4 | Supporting Data and Test Work | 16-5

| 16.4.1 | Ore Classifications and Rock Types | 16-5

| 16.4.2 | Mineralogy | 16-5

| 16.4.3 | Comminution | 16-6

| 16.4.4 | Selection of Optimum Grind Size | 16-6

| 16.4.5 | Flotation | 16-7

| 16.4.6 | Concentrate Quality | 16-8

| 16.4.7 | Cyanidation of Cleaner Flotation Tails | 16-8

| 16.4.8 | Thickening | 16-9

| 16.4.9 | Filtration and Transportable Moisture Limits | 16-9

| 16.4.10 | Slurry Rheology | 16-9

| 16.4.11 | Cyanide Destruction and Water Treatment | 16-9

| 16.5 | Supporting Data and Test Work | 16-10

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| 16.5.1 | Comminution | 16-10

| 16.5.2 | Flotation | 16-12

| 16.5.3 | Regrinding | 16-15

| 16.5.4 | Cyanidation of Cleaner Tailings | 16-15

| 16.5.5 | Solid-Liquid Separation | 16-16

| 16.5.6 | Concentrate Pumping and Slurry Pipeline | 16-16

| 16.5.7 | Heap Leaching | 16-17

17.0 | MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES | 17-1

| 17.1 | Mineral Resource Estimates | 17-1

| 17.1.1 | Geological Models and Data Analysis | 17-1

| 17.1.2 | Histograms, Cumulative Frequency Plots, and Boxplots | 17-2

| 17.1.3 | Grade Scatterplots | 17-7

| 17.1.4 | Contact Profile Analysis | 17-8

| 17.1.5 | Estimation Domains | 17-8

| 17.2 | Evaluation of Extreme Grades | 17-9

| 17.3 | Variography | 17-10

| 17.4 | Estimation | 17-11

| 17.4.1 | Validation | 17-12

| 17.5 | Mineral Resource Classification | 17-13

| 17.6 | Mineral Resources | 17-14

| 17.7 | Mineral Reserves | 17-15

18.0 | OTHER RELEVANT DATA AND INFORMATION | 18-1

19.0 | REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION AND DEVELOPMENT PROPERTIES | 19-1

| 19.1 | Mining Operations | 19-1

| 19.1.1 | Economic Modeling | 19-1

| 19.1.2 | Pit Shell Optimization | 19-5

| 19.1.3 | Pit Access and Phase Design | 19-5

| 19.1.4 | Stockpile and Dump Design | 19-6

| 19.1.5 | Production Schedule | 19-8

| 19.1.6 | Equipment | 19-9

| 19.2 | Recoverability | 19-10

| 19.3 | Markets | 19-10

| 19.4 | Contracts | 19-10

| 19.5 | Environmental | 19-10

| 19.6 | Taxes | 19-10

| 19.7 | Operating Cost Estimates | 19-11

| 19.7.1 | Operating Cost Summary | 19-11

| 19.7.2 | Mine Operating Costs | 19-11

| 19.7.3 | Processing Plant and Heap Operating Costs | 19-12

| 19.7.4 | General and Administrative Operating Costs | 19-13

| 19.8 | Capital Cost Estimates | 19-13

| 19.8.1 | Capital Cost Review | 19-15

| 19.8.2 | Sustaining Capital Cost Review | 19-16

| 19.9 | Economic Analysis | 19-16

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| K INROSS G OLD C ORPORATION C ERRO C ASALE P ROJECT, C HILE T ECHNICAL R EPORT

| 19.10 | Payback | 19-20

| 19.11 | Mine Life | 19-20

20.0 | INTERPRETATIONS AND CONCLUSIONS | 20-1

| 20.1 | Technical Basis | 20-1

| 20.2 | Mineralization and Alteration | 20-1

| 20.3 | Drilling | 20-1

| 20.4 | Sampling, Sample Preparation, and Assaying | 20-1

| 20.5 | Assay QA/QC | 20-2

| 20.6 | Density | 20-3

| 20.7 | Data Verification | 20-3

| 20.8 | Geological Interpretations | 20-3

| 20.9 | Metallurgical Processing | 20-4

| 20.9.1 | Comminution | 20-4

| 20.9.2 | Flotation | 20-4

| 20.9.3 | Dump Leaching of Oxide Ores | 20-5

| 20.9.4 | Other Plant and Process Issues | 20-5

| 20.10 | Mineral Resource and Mineral Reserve Estimates | 20-5

| 20.10.1 | Resource Classification | 20-5

| 20.10.2 | Mineral Resources | 20-5

| 20.10.3 | Mineral Reserves | 20-7

| 20.11 | Mine Designs and Production Plans | 20-8

| 20.12 | Operating Cost Estimates | 20-8

| 20.13 | Capital Cost Estimates | 20-9

| 20.14 | Economic Analysis | 20-9

| 20.14.1 | Sensitivity Analysis | 20-10

| 20.15 | Permitting and Environmental Studies | 20-10

21.0 | RECOMMENDATIONS | 21-1

22.0 | REFERENCES | 22-1

23.0 | DATE AND SIGNATURE PAGE | 23-1

T ABLES

Table 1-1: | Mineral Reserve Summary (MQes, June 2006) | 1-3

Table 1-2: | Mineral Resource Summary (MQes, June 2006) | 1-3

Table 4-1: | Area of Interest | 4-2

Table 4-2: | Mineral Concessions within Aldebarán Area of Interest | 4-4

Table 7-1: | Major Lithological Units at Cerro Casale | 7-4

Table 11-1: | Cerro Casale Drilling | 11-1

Table 13-1: | Check Assays by Chemex, 1996 and 1997 (MRDI, 1997b) | 13-9

Table 13-2: | Acme and Chemex Analyses of Standard, 1996-1997 (MRDI, 1997b) | 13-9

Table 13-3: | 1998 Standards and Blanks Used at Cerro Casale - Gold | 13-12

Table 13-4: | 1998 Standards and Blanks Used at Cerro Casale - Copper | 13-12

Table 13-5: | 1998 Check Assay Statistics | 13-22

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| K INROSS G OLD C ORPORATION C ERRO C ASALE P ROJECT, C HILE T ECHNICAL R EPORT

Table 13-6: | 1999 Check Assay Statistics | 13-33

Table 13-7: | Summary Statistics for Bulk Density Determinations, by Rock Type, All Sulphides | 13-34

Table 13-8: | Summary Statistics for Bulk Density Determinations, by Oxidation State, All Rock Types | 13-34

Table 13-9: | Specific Gravity for Mineralization Domains | 13-35

Table 16-1: | Metal Recoveries into Concentrate | 16-4

Table 16-2: | Mineral Reserves by Metallurgical Rock Type (MQes, 2006) | 16-5

Table 16-3: | Indicative Concentrate Quality Analyses | 16-8

Table 16-4: | Comparative Grindability Parameters | 16-10

Table 16-5: | Power Draw Estimates for Cerro Casale Comminution Circuits | 16-11

Table 16-6: | Estimated Power Required vs. Ore Hardness (Work Index) | 16-12

Table 16-7: | Avgerage Metallurgical Performance for Combined MDBX, VBX, and GDS Ore | 16-14

Table 16-8: | Flotation Cell Selection Criteria | 16-14

Table 17-1: | Gold and Copper Geologic Models or Domains | 17-2

Table 17-2: | Cutting Thresholds or Cap Grades for Gold and Copper Composite Data, (PDTS, 2000) | 17-9

Table 17-3: | Gold and Copper Variogram Parameters for Estimation Domains (PDTS, 2000) | 17-11

Table 17-4: | Global Model Mean Grade Values by Domain | 17-14

Table 17-5: | Cerro Casale Mineral Resources (MQes, June 2006) | 17-15

Table 17-6: | Mineral Reserves (MQes, June 2006) | 17-16

Table 19-1: | Process Costs | 19-3

Table 19-2: | Base and Incremental Mining Costs | 19-4

Table 19-3: | Unit Operating Costs | 19-11

Table 19-4: | Total Project Capital Costs (MQes, 2006) | 19-13

Table 19-5: | Total Pre-production Capital Costs (MQes, 2006) | 19-14

Table 19-6: | Sustaining Capital Costs (MQes, 2006) | 19-15

Table 19-7: | Summary Cashflow (MQes, 2006) | 19-17

Table 19-8: | Summary Cashflow (MQes, 2006) | 19-18

Table 19-9: | Metal Price Sensitivity Analysis (MQes, 2006) | 19-19

F IGURES

Figure 4-1: | Location of the Cerro Casale Gold-Copper Deposit, Northern Chile | 4-1

Figure 4-2: | Mineral Claims and Area of Interest, Aldebarán (PDTS, 2000) | 4-3

Figure 5-1: | Location of Cerro Casale Project, Northern Chile | 5-2

Figure 5-2: | Mill Site | 5-2

Figure 5-3: | Tailings and Waste Rock Site | 5-3

Figure 7-1: | Geology of the Maricunga Volcanic Belt (PDTS, 2000) | 7-2

Figure 7-2: | Surface Geological Map of Cerro Casale (PDTS, 2000) | 7-5

Figure 7-3: | Cross Section 850E, Looking Northwest, Cerro Casale Deposit (PDTS, 2000) | 7-6

Figure 7-4: | Redox Units, Section 850E (PDTS, 2000) | 7-9

Figure 9-1: | Major Gold-Copper Occurrences in the Aldebarán Property (PDTS, 2000) | 9-2

Figure 9-2: | Measured + Indicated Gold Resources, Section 472200E (MQes, 2006) | 9-4

Figure 9-3: | Measured + Indicated Copper Resources, Section 472200E (MQes, 2006) | 9-5

Figure 9-4: | Measured + Indicated Gold Resources, 3832 Elevation (MQes, 2006) | 9-6

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Figure 9-5: | Measured + Indicated Copper Resources, 3832 Elevation (MQes, 2006) | 9-7

Figure 9-6: | Intensity of Stockwork Veining, Section 850E (PDTS, 2000) | 9-8

Figure 9-7: | Potassium Feldspar Alteration, Section 850E (PDTS, 2000) | 9-9

Figure 9-8: | Geological Map of the Eva Deposit (PDTS, 2000) | 9-11

Figure 9-9: | Cross Section of Eva Deposit (PDTS, 2000) | 9-12

Figure 9-10: | Geological Map of the Cerro Roman Deposit (PDTS, 2000) | 9-14

Figure 9-11: | North-South Cross Section of Cerro Roman Deposit (PDTS, 2000) | 9-15

Figure 11-1: | Drill Collar Locations (PDTS, 2000) | 11-2

Figure 11-2: | Average and Median Drill Spacing by Elevation (PDTS, 2000) | 11-4

Figure 11-3: | Drill-Hole Collar Monuments | 11-9

Figure 13-1: | Relative Differences for Rig Duplicates (MRDI, 1997b) | 13-5

Figure 13-2: | Checks of Acme Gold Assays by Chemex (MRDI, 1997b) | 13-7

Figure 13-3: | Precision from Chemex Check Assays of Acme Gold Assays (MRDI, 1997b) | 13-7

Figure 13-4: | Chemex Check Assays of Acme Copper Assays (MRDI, 1997b) | 13-8

Figure 13-5: | 1998 Cerro Casale Standard (Blank) STD05 - Gold | 13-12

Figure 13-6: | 1998 Cerro Casale Standard (Blank) STD05 - Copper | 13-13

Figure 13-7: | 1998 Cerro Casale Standard STD12 - Gold | 13-13

Figure 13-8: | 1998 Cerro Casale Standard STD12 - Copper | 13-14

Figure 13-9: | 1998 Cerro Casale Standard STD13 - Gold | 13-15

Figure 13-10: | 1998 Cerro Casale Standard STD13 - Copper | 13-15

Figure 13-11: | 1998 Cerro Casale Standard STD14 - Gold | 13-16

Figure 13-12: | 1998 Cerro Casale Standard STD14 - Copper | 13-16

Figure 13-13: | 1998 Cerro Casale Standard STD18 - Gold | 13-17

Figure 13-14: | 1998 Cerro Casale Standard STD18 - Copper | 13-18

Figure 13-15: | 1998 Cerro Casale Standard (Blank) STD19 - Gold | 13-18

Figure 13-16: | 1998 Cerro Casale Standard (Blank) STD19 - Copper | 13-19

Figure 13-17: | 1998 Cerro Casale Gold Duplicate Data | 13-20

Figure 13-18: | 1998 Cerro Casale Gold Duplicate Data | 13-21

Figure 13-19: | 1998 Cerro Casale Copper Duplicate Data | 13-21

Figure 13-20: | 1998 Cerro Casale Copper Duplicate Data | 13-22

Figure 13-21: | 1999 Cerro Casale Standard STD12 - Gold | 13-24

Figure 13-22: | 1999 Cerro Casale Standard STD12 - Copper | 13-24

Figure 13-23: | 1999 Cerro Casale Standard STD13 - Gold | 13-25

Figure 13-24: | 1999 Cerro Casale Standard STD13 - Copper | 13-25

Figure 13-25: | 1999 Cerro Casale Standard STD14 - Gold | 13-27

Figure 13-26: | 1999 Cerro Casale Standard STD14 - Copper | 13-27

Figure 13-27: | 1999 Cerro Casale Standard STD18 - Gold | 13-28

Figure 13-28: | 1999 Cerro Casale Standard STD18 - Copper | 13-28

Figure 13-29: | 1999 Cerro Casale Standard (Blank) STD19 - Gold | 13-29

Figure 13-30: | 1999 Cerro Casale Standard (Blank) STD19 - Copper | 13-29

Figure 13-31: | 1999 Cerro Casale Gold Duplicate Data | 13-30

Figure 13-32: | 1999 Cerro Casale Gold Precision Estimate | 13-31

Figure 13-33: | 1999 Cerro Casale Precision Estimate by Data Date | 13-31

Figure 13-34: | 1999 Cerro Casale Duplicate Copper Data | 13-32

Figure 13-35: | 1999 Cerro Casale Copper Precision Estimate | 13-32

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Figure 13-36: | Boxplot of All Density Measurements by Oxidation Categories | 13-35

Figure 16-1: | Generalized Process Flowsheet | 16-13

Figure 17-1: | Boxplot Summary of Gold Composite Data, Un-cut (PDTS, 2000) | 17-3

Figure 17-2: | Boxplot Summary of Gold Composite Data, Cut Grades (PDTS, 2000) | 17-4

Figure 17-3: | Boxplot Summary of Copper Composite Data, Un-Cut (PDTS, 2000) | 17-5

Figure 17-4: | Boxplot Summary of Copper Composite Data, Cut Grades (PDTS, 2000) | 17-6

Figure 17-5: | Gold vs. Copper Scatterplot (PDTS, 2000) | 17-7

Figure 19-1: | Site Plan (PDTS, 2000) | 19-2

Figure 19-2: | North Looking Section through Pit Phases (MQes, 2006) | 19-6

A PPENDICES

Appendix A: List of Significant Assays

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1.0 | SUMMARY

The Cerro Casale project is currently envisioned as a conventional open pit, which produces nominally 150,000 t/d from this gold-copper porphyry deposit over a 17-year mine life. Processing facilities designs a 75,000 t/d cyanide heap leach facility for oxide ores and a 150,000 t/d semi-autogenous grinding (SAG) mill and flotation concentrator for sulphide and mixed ores. Doré bars will be produced on site from leachate recovered from the heap and flotation cleaner tails leach circuits. Sulphide copper concentrates will be pumped to the port at Punta Padrones near Caldera via a 250 km pipeline, and shipped to smelting and refining facilities.
The Cerro Casale deposit is located in Region Three of northern Chile. The city of Copiapo is 145 km to the northwest (180 km by road). The approximate geographic coordinates of the site are 27° 47’ S and 69 o 17’ W. The international border separating Chile and Argentina is approximately 20 km to the east. The deposit is located in an area of major relief, with local variations in topography ranging from 3700 to 5800 m in elevation.
Compañía Minera Casale (CMC), a contractual mining company formed under the laws of the Republic of Chile, owns the Cerro Casale project. The share capital of CMC is owned by Arizona Star Resources Corporation (Arizona Star) 51.0% and Bema Gold Corporation (Bema) 49.0%. The relationship of the CMC shareholders is governed by a Letter Agreement, between Arizona Star and Bema, drafted on 19 June 2006. The General Manager of the project is Bema.
The Cerro Casale gold-copper deposit is located in the Aldebarán sub-district of the Maricunga Volcanic Belt. The Maricunga belt is comprised of a series of coalescing composite, Miocene, and andesitic to rhyolitic volcanic centers that extend for 200 km along the western crest of the Andes. The volcanic rocks are host to multiple epithermal gold and porphyry-hosted gold-copper deposits, including Cerro Casale, Refugio, Marte, and La Copia. The volcanic rocks overlie older sedimentary and volcanic rocks of Mesozoic and Paleozoic age. Reverse faults parallel to the axis of the Andes have uplifted hypabyssal intrusive rocks beneath the extrusive volcanics, exposing porphyry-hosted gold-copper deposits in the Aldebarán area such as Cerro Casale, Eva, Jotabeche, Estrella, and Anfiteatro. Composite volcanic centers are still preserved in the immediate Cerro Casale area at Volcan Jotabeche and Cerro Cadillal. Extensive hydrothermal alteration consisting of quartz-feldspar veinlet stockworks, biotite-potassium feldspar, quartz-sericite, and chlorite occurs in these intrusive centers. Gold-copper mineralization is principally associated with intense quartz-sulphide stockworks, potassic, and phylliic alteration.

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Gold-copper mineralization occurs in quartz-sulphide and quartz-magnetite-specularite veinlet stockworks developed in the dioritic to granodioritic intrusives and adjacent volcanic wall rocks. Stockworks are most common in two dioritic intrusive phases, particularly where intrusive and hydrothermal breccias are developed. Mineralization extends at least 1,450 m vertically and 850 m along strike. The strike of mineralization follows WNW (310°) fault and fracture zones. The main zone of mineralization pinches and swells from 250 m to 700 m along strike and down dip steeply to the southwest. The highest-grade mineralization is coincident with well developed quartz-sulphide stockworks in strongly potassic-altered intrusive rocks.
Oxidation resulting from weathering and/or high oxygen activity in the last phase of hydrothermal alteration overprints sulphide mineralization in the upper portion of the Cerro Casale deposit. Oxidation locally extends deeply along fault zones or within steeply dipping breccia bodies. Oxidation generally goes no deeper than 15 m where vertical structures are absent. Oxide is present in linear oxidation zones as deep as 300 m along major fault and fracture zones, or as pendants along the intersection of multiple fault zones.
Reverse-circulation (RC) and core drilling was performed in multiple campaigns since 1989. Anglo American drilled two RC holes in 1989. The Bema Shareholder Group drilled a large number of RC and core holes between 1991 and 1997. Placer Dome Latin America drilled additional confirmation, infill, and geotechnical core holes in 1998 and 1999. A total of 224 RC and 124 core holes totaling 122,747 m support the Cerro Casale resource estimate. No additional drilling has been performed since 1999.
Metallurgical test work is used to categorize ore types on the basis of metallurgical characteristics for comminution, optimal grind size, flotation response, cyanidation of tails (for gold), and trace element content. Metallurgical recovery equations for gold and copper were developed for eight ore types.
Feasibility-level studies were prepared in 1997 and 2000, with updates in 2004 and 2005. The general project concept has evolved from a 35,000 t/d heap leach project to a mixed sulphide/oxide plant concept in 2004, to its current form. The 2004 and 2005 studies were more conceptual, incorporating design improvements and updated cost estimates. Over the last year, Mine and Quarry Engineering Services, Inc. (MQes) has been evaluating mining and processing alternatives, which are being incorporated into this Technical Report. The current study uses elements from all previous studies, incorporating some revised equipment and operating cost estimates, scale-up factors, and escalation.
Mineral Resources and Mineral Reserve estimates comply with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000), CIM Definition Standards for Mineral Resources and Mineral Reserves (December 11, 2005) and Canadian National Instrument 43-101 of the Canadian Securities Administrators (December 30, 2005).

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Mineral reserves for Cerro Casale are summarized in Table 1-1.
Table 1-1: Mineral Reserve Summary (MQes, 24 June 2006)

|  |  |  |  |  |  | Grades |  |  | Contained Metal |
| Mineral Reserve Category |  |  | Tonnage (Mt) |  |  | Gold (g/t) |  |  | Copper (%) |  |  | Gold (K oz) |  |  | Copper (M lb) |
| Proven |  |  | 205 |  |  | 0.71 |  |  | 0.24 |  |  | 4,706 |  |  | 1,099 |
| Probable |  |  | 830 |  |  | 0.68 |  |  | 0.26 |  |  | 18,228 |  |  | 4,706 |
| Proven + Probable |  |  | 1,035 |  |  | 0.69 |  |  | 0.25 |  |  | 22,934 |  |  | 5,805 |

Notes: 1. Mineral Reserve metal prices are $450/oz Au and $1.50/lb Cu. 2. Metallurgical recovery equations are noted in Table 16-3. 3. The life-of-mine waste-to-ore strip ratio is 2.9:1. 4. Summation errors are due to rounding.
Cerro Casale mineral resources are reported in Table 1-2. Mineral resources are exclusive of mineral reserves.
Table 1-2: Mineral Resource Summary (MQes, 24 June 2006)

|  |  |  |  |  |  | Grades |  |  | Contained Metal |
| Mineral Resource Category |  |  | Tonnage (Mt) |  |  | Gold (g/t) |  |  | Copper (%) |  |  | Gold (K oz) |  |  | Copper (M lb) |
| Measured |  |  | 34 |  |  | 0.40 |  |  | 0.22 |  |  | 436 |  |  | 164 |
| Indicated |  |  | 347 |  |  | 0.40 |  |  | 0.24 |  |  | 4,460 |  |  | 1,835 |
| Measured + Indicated |  |  | 381 |  |  | 0.40 |  |  | 0.24 |  |  | 4,896 |  |  | 1,835 |
| Inferred |  |  | 301 |  |  | 0.35 |  |  | 0.25 |  |  | 3,385 |  |  | 1,657 |

Notes: 1. Mineral Resources are defined with a Lerchs-Grossman pit design based on metal prices of $550/oz Au and $1.75/lb Cu, and average G&A costs of $0.47/t milled, mining costs of $0.80/t mined, stockpile re-handling costs of $0.29/t re-handled, heap leach costs of $1.85/t leached, and plant operating costs of $3.31/t milled. 2. Mineral Resources are incremental to the Mineral Reserves. 3. Summation errors are due to rounding.
Project capital costs are estimated by MQes to be $1,960 million, assuming the installation of all new equipment. Sustaining capital costs are estimated to be $263 million. Life-of-mine on-site operating costs average $7.68/t. Transportation, smelting, and refining charges equate to an incremental operating cost of $1.92/t of ore processed.
MQes financial analyses indicate the base case mine plan has a positive return, with a payback period of 4.9 years.
The MQes cashflow model excludes working capital allocations. In addition, AMEC identified capital and operating cost increases that in its opinion would be appropriate to include. These were incorporated by AMEC in an economic sensitivity analysis which provides a more conservative view of the project. When all these adjustments are considered simultaneously, there is an impact on payback but the internal rate of return remains positive.

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AMEC concludes that the data collection, sample preparation, QA/QC data, density assignments, assay and geological databases, metallurgical test work and interpretation, plant design, mine design and operating cost estimates generally meet or exceed industry standards, and are of sufficient quality to support preparation of mineral resource and mineral reserve estimates.
The Cerro Casale financial model is based on revisions in previously prepared pre-feasibility and feasibility level studies with significant modifications in mine designs and processing plans. Modifications in mine designs have been executed to at least prefeasibility level. Modifications in processing plans and associated capital and operating costs have been executed to a combination of prefeasibility and scoping study levels to test the impact of these new development options. AMEC recommends preparing a comprehensive prefeasibility-level analysis, which includes a revised block model, updated mining plans, current processing concepts, and first principle cost estimates to increase confidence in the financial model and support a decision to further evaluate or develop the project.

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2.0 | INTRODUCTION

Bema Gold Corporation (Bema) commissioned AMEC E&C Services (AMEC) to prepare an updated Technical Report for the Cerro Casale Project gold-copper project, northern Chile. The purpose of this report is to support the public disclosure of updated mineral resources and mineral reserves. Subsequently, Kinross Gold Corporation (Kinross) requested that AMEC re-address the report to Kinross for the purpose of supporting disclosure of Cerro Casale mineral resources and mineral reserves in Kinross’ Annual Information Form (AIF) for fiscal year 2006. The scope of AMEC’s work included reviewing technical and economic aspects of the project prepared by MQes, such as mine plans, processing concepts, cost estimates, and economic parameters, and incorporating the revisions into an updated Technical Report. Resource estimation work and resource models have not changed from that reported in the NI 43-101 Technical Report and Qualified Persons Review prepared by AMEC in March 2005 (2005 Technical Report). The 2005 Technical Report and related work determined that the resource estimates were developed in accordance with industry standard practices and that the mineral resource and mineral reserve estimates in the 2000 Feasibility Study and March 2004 Feasibility Study Updates are compliant with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000), CIM Definition Standards for Mineral Resources and Mineral Reserves (2000) and Canadian National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators. The format and content of the previous report and this current report are intended to conform to Form 43-101F1.

2.1 | Sources of Information

MQes provided the data used to prepare this report, including the resource block model, cash flow model, and copies of previously published reports.
Aspects of this report summarizing project history and geology were derived from previous studies and 43-101 Technical Reports on the Cerro Casale project.
AMEC did not review any new information specific to Sections 7 through 15, which are presented verbatim from the 2005 Technical Report. No additional work has been done in regards to the geology, exploration, drilling, sampling, assaying, data verifications, resource estimates, environmental conditions and permitting since AMEC’s March 2005 Technical Report. Other references are listed in Section 22.

2.2 | Qualified Persons

Larry B. Smith, R.Geo., Ch.P.Geo, AusIMM and William A. Tilley PE, employees of AMEC, served as Qualified Persons responsible for preparing this report.

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Larry Smith visited the property 12 and 13 January 2005 and reviewed pertinent aspects of geology, exploration data, geological models, land status, infrastructure and mine designs, as part of the 2005 Technical Report team. Larry Smith also reviewed additional documentation for geological models, exploration databases, resource estimates and reserve estimates at Placer Dome offices in Santiago on 14 to 20 January 2005. Mr. Smith is the principal Qualified Person for this report, and is responsible for preparing Sections 1, 3, 6 through 16, and 20 through 22. He was assisted in review of metallurgy, changes in processing methods and processing operating costs by Jerry Jergensen, Associate Metallurgist.
Mr. Smith was assisted in review of resource estimates for the 2005 Technical Report by Dr. Stephen Juras, P.Geo. The resource block model used in the most recent work was the same as reviewed by AMEC in 2005. For AMEC’s 2005 Technical Report Larry was also assisted in the review of environmental baseline studies, environmental management provisions, closure plans and environmental permits. by Lydia LeTourneau, Environmental Specialist and an employee of AMEC. These matters have not changed since 2005.
William A. Tilley reviewed the mine planning, cost estimating, and financial analysis, with detailed assistance provided by Mark Hertel and Manuel Romero, employees of AMEC, and Simon Handelsman, Associate Financial Analyst. Mr. Tilley is responsible for preparation of Sections 2, 4, 5, and 17 through 19.
AMEC is not an associate or affiliate of Bema, or of any associated company. AMEC’s fee for this Technical Report is not dependent in whole or in part on any prior or future engagement or understanding resulting from the conclusions of this report. This fee is in accordance with standard industry fees for work of this nature, and is based solely on the approximate time needed to assess the various data and reach the appropriate conclusions.

2.3 | Terms of Reference

Unless stated otherwise, all quantities are in metric units and currencies are expressed in constant 2006 US dollars. This report is written for the entire project; the interests of any particular shareholder must therefore be deduced from the figures presented.

2.4 | Units of Measure

2.4.1 | Common Units

Above mean sea level |  | amsl
Ampere |  | A
Annum (year) |  | a
Billion years ago |  | Ga

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British thermal unit |  | Btu
Candela |  | cd
Centimetre |  | cm
Cubic centimetre |  | cm 3
Cubic feet per second |  | ft 3/s or cfs
Cubic foot |  | ft 3
Cubic inch |  | in 3
Cubic metre |  | m 3
Cubic yard |  | yd 3
Day |  | d
Days per week |  | d/wk
Days per year (annum) |  | d/a
Dead weight tonnes |  | DWT
Decibel adjusted |  | dBa
Decibel |  | dB
Degree |  | °
Degrees Celsius |  | °C
Degrees Fahrenheit |  | °F
Diameter |  | ø
Dry metric ton |  | dmt
Foot |  | ft
Gallon |  | gal
Gallons per minute (US) |  | gpm
Gigajoule |  | GJ
Gram |  | g
Grams per litre |  | g/L
Grams per tonne |  | g/t
Greater than |  | >
Hectare (10,000 m 2) |  | ha
Hertz |  | Hz
Horsepower |  | hp
Hour |  | h (not hr)
Hours per day |  | h/d
Hours per week |  | h/wk
Hours per year |  | h/a
Inch |  | “ (symbol, not” )
Joule |  | J
Joules per kilowatt-hour |  | J/kWh
Kelvin |  | K
Kilo (thousand) |  | k
Kilocalorie |  | kcal
Kilogram |  | kg
Kilograms per cubic metre |  | kg/m 3
Kilograms per hour |  | kg/h
Kilograms per square metre |  | kg/m 2
Kilojoule |  | kJ

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Kilometre |  | km
Kilometers per hour |  | km/h
Kilonewton |  | kN
Kilopascal |  | kPa
Kilovolt |  | kV
Kilovolt-ampere |  | kVA
Kilovolts |  | kV
Kilowatt |  | kW
Kilowatt hour |  | kWh
Kilowatt hours per short ton (US) |  | kWh/st
Kilowatt hours per tonne (metric ton) |  | kWh/t
Kilowatt hours per year |  | kWh/a
Kilowatts adjusted for motor efficiency |  | kWe
Less than |  | <
Litre |  | L
Liters per minute |  | L/m
Megabytes per second |  | Mb/s
Megapascal |  | MPa
Megavolt-ampere |  | MVA
Megawatt |  | MW
Metre |  | m
Meters above sea level |  | masl
Meters per minute |  | m/min
Meters per second |  | m/s
Metric ton (tonne) |  | t
Micrometer (micron) |  | µm
Microsiemens (electrical) |  | µs
Miles per hour |  | mph
Milliamperes |  | mA
Milligram |  | mg
Milligrams per litre |  | mg/L
Milliliter |  | mL
Millimeter |  | mm
Million |  | M
Million tonnes |  | Mt
Minute (plane angle) |  | ‘
Minute (time) |  | min
Month |  | mo
Newton |  | N
Newtons per metre |  | N/m
Ohm (electrical) |  | Ω
Ounce |  | oz
Parts per billion |  | ppb
Parts per million |  | ppm
Pascal (newtons per square metre) |  | Pa
Pascals per second |  | Pa/s

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Percent |  | %
Percent moisture (relative humidity) |  | % RH
Phase (electrical) |  | Ph
Pound(s) |  | lb
Pounds per square inch |  | psi
Power factor |  | pF
Quart |  | qt
Revolutions per minute |  | rpm
Second (plane angle) |  | “
Second (time) |  | s
Short ton (2,000 lb) |  | st
Short ton (US) |  | t
Short tons per day (US) |  | tpd
Short tons per hour (US) |  | tph
Short tons per year (US) |  | tpy
Specific gravity |  | SG
Square centimetre |  | cm 2
Square foot |  | ft 2
Square inch |  | in 2
Square kilometer |  | km 2
Square metre |  | m 2
Thousand tonnes |  | kt
Tonne (1,000 kg) |  | t
Tonnes per day |  | t/d
Tonnes per hour |  | t/h
Tonnes per year |  | t/a
Total dissolved solids |  | TDS
Total suspended solids |  | TSS
Volt |  | V
Week |  | wk
Weight/weight |  | w/w
Wet metric ton |  | wmt
Yard |  | yd
Year (annum) |  | a

2.4.2 | Common Chemical Symbols

Aluminum |  | Al
Ammonia |  | NH 3
Antimony |  | Sb
Arsenic |  | As
Bismuth |  | Bi
Cadmium |  | Cd
Calcium |  | Ca
Calcium carbonate |  | CaCO 3

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Calcium oxide |  | CaO
Calcium sulphate di-hydrate |  | CaSO 4•2H 2 O
Carbon |  | C
Carbon monoxide |  | CO
Chlorine |  | Cl
Chromium |  | Cr
Cobalt |  | Co
Copper |  | Cu
Cyanide |  | CN
Gold |  | Au
Hydrogen |  | H
Iron |  | Fe
Lead |  | Pb
Magnesium |  | Mg
Manganese |  | Mn
Manganese dioxide |  | MnO 2
Manganous hydroxide |  | Mn (OH)2
Molybdenum |  | Mo
Nickel |  | Ni
Nitrogen |  | N
Nitrogen oxide compounds |  | NOx
Oxygen |  | O 2
Palladium |  | Pd
Platinum |  | Pt
Potassium |  | K
Silver |  | Ag
Sodium |  | Na
Sulphur |  | S
Tin |  | Sn
Titanium |  | Ti
Tungsten |  | W
Uranium |  | U
Zinc |  | Zn

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3.0 | RELIANCE ON OTHER EXPERTS

The results and opinions expressed in this Technical Report are based on AMEC’s field observations, discussions with Bema, Placer Dome, and MQes personnel, and the geological and technical data listed in the references.
The results and opinions expressed in this report are conditional upon the aforementioned technical and legal information being current, accurate, and complete as of the date of this report, and the understanding that no information has been withheld that would affect the conclusions made herein. AMEC reserves the right to revise this report and conclusions if additional information becomes known to AMEC subsequent to the date of this report. AMEC does not assume responsibility for Bema’s actions in distributing this report.
Areas where AMEC has relied on the opinions of other experts include the following:
AMEC did not independently verify the validity of mineral exploration and exploitation licenses and surface agreements. AMEC relied upon a report by Grasty Quintana & Cia (1997) regarding legal title of the mining property, water rights, surface permits, environmental permits and non-environmental permits (an independent legal firm for Barrick when they were doing due diligence on Placer Dome). These reports indicate that all exploration and exploitation concessions, environmental permits and well field permits are secure and not under legal challenge. There have been no material changes in this area since 1997.
The main technical documents consulted for the review of environmental matters include:

| · | the Environmental Impact Study prepared by CMC, dated December 2000 and associated baseline studies prepared by SENES Chile S.A in 1999 and 2000

| · | the January 2000 version of Volume 4 of the Aldebarán Project (equivalent to Cerro Casale) prepared by PDTS Limited of Vancouver

| · | the project’s environmental approval “Resolución Exenta No 014” granted by COREMA on 31 January 2002.

AMEC has not reviewed any specific laboratory test results or detailed information on the potential for Acid Rock Drainage (ARD) as the report from Phase 1 and 2 work on Prediction of Drainage Chemistry prepared by the Minesite Drainage Assessment Group in October 1999 was not available for review. Information on ARD potential contained in this Technical Report is strictly derived from a review of the ARD prediction report’s executive summary and an interoffice memorandum prepared by Keith Ferguson of Placer Dome on 27 October 1999.

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Legal information on regulatory requirements is extracted from reference material listed in Chapter 21, which includes copies of legislative instruments published by the Government of Chile.

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4.0 | PROPERTY DESCRIPTION AND LOCATION

The Cerro Casale deposit is located in Region Three of northern Chile. The city of Copiapo is 145 km northwest of the deposit (Figure 4-1). The approximate geographic coordinates of the project are 27° 47’ S and 69 o 17’ W. The international border separating Chile and Argentina is approximately 20 km to the east. The deposit is located in an area of major relief, with local variations in topography ranging from 3700 to 5800 m in elevation. The top of the Cerro Casale deposit is at an elevation of 4450 m.
Figure 4-1: Location of the Cerro Casale Gold-Copper Deposit, Northern Chile

4.1 | Project Ownership and Agreements

CMC is a contractual mining company incorporated under the laws of the Republic of Chile. CMC owns the Cerro Casale project. In accordance with a Letter Agreement dated June 19, 2006 between Bema Gold Corporation (Bema) and Arizona Star Resource Corporation (Arizona Star), Bema and Arizona Star completed the acquisition of Barrick’s (formerly Placer Dome Inc.’s) 51% of the shares of CMC. Pursuant to this Agreement, Bema and Arizona Star terminated the existing Amended and Restated Shareholder’s Agreement dated June 5, 2003.

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Under the terms of this Agreement, Bema and Arizona Star will jointly pay to Barrick US$10 million upon a decision to construct a mine at Cerro Casale and either (a) a gold payment beginning 12 months after commencement of production consisting of 10,000 ounces of gold per year for five years and 20,000 ounces of gold per year for a subsequent seven years; or (b) a cash payment of US$70 million payable when a construction decision is made, at the election of Bema and Arizona Star.
In accordance with the Purchase and Sales Agreement, CMC is now controlled 51% by Arizona Star Resources Corporation and 49% by Bema Gold Corporation. Bema is the project General Manager.
CMC owns the presently valid mineral and water concessions within an Area of Interest, and has applied for additional mineral and water concessions in the region.

4.2 | Mineral, Surface, and Water Rights

4.2.1 | Mineral Rights

The Cerro Casale gold-copper deposit and lesser explored satellite deposits comprising the Aldebarán Project are located within an Area of Interest previously described in the Amended and Restated Shareholders’ Agreement and included in the June 16, 2006 Purchase and Sale Agreement (Table 4-1). Deposits with less exploration to date include Eva, Cerro Roman, Anfiteatro, Estrella, and Romancito Sur (Figure 4-2). CMC has performed drilling on these satellite deposits sufficient for preliminary estimates of gold-copper mineralization.
The Area of Interest as approved by the CMC Board is defined by the following U.T.M. coordinates and comprises approximately 20,000 ha.
Table 4-1: Area of Interest

| Corner # |  | North (m) |  | East (m) |
| 1 |  |  | 6,939,000.00 |  |  | 458,000.00 |
| 2 |  |  | 6,905,000.00 |  |  | 493,000.00 |
| 3 |  |  | 6,939,000.00 |  |  | 493,000.00 |
| 4 |  |  | 6,905,000.00 |  |  | 458,000.00 |

Cerro Casale is located within the area between U.T.M coordinates 6925000N-6927000N and 471900E-473000E.

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Figure 4-2: Mineral Claims and Area of Interest, Aldebarán (PDTS, 2000)

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CMC-owned mining claims within the Area of Interest include 4,105 patented claims in 30 groups (Table 4-2), totaling 19,955 ha. Claim overlaps reduce the actual area to 19,520 ha. All mineral rights are protected according to Chilean law, by payment of a mining patent.
Table 4-2: Mineral Concessions within Aldebarán Area of Interest

| Register No. |  | Names |  | Number of Claims |  | Area (ha) |
| 03203-1219-1 |  |  | NEVADO 1/840 |  |  | 840 |  |  | 4,200 |
| 03203-1220-5 |  |  | CACHITO 1/1298 |  |  | 1,298 |  |  | 6,490 |
| 03203-1247-7 |  |  | HORUS 1/280 |  |  | 160 |  |  | 800 |
| 03203-1248-5 |  |  | OLIMPO 1/293 |  |  | 30 |  |  | 150 |
| 03203-1249-3 |  |  | MARTE 1/300 |  |  | 300 |  |  | 1,500 |
| 03203-3458-5 |  |  | RAHIL 1/48 |  |  | 48 |  |  | 240 |
| 03203-3849-2 |  |  | PACO 1/60 |  |  | 60 |  |  | 300 |
| 03203-3850-6 |  |  | LUIS 1/40 |  |  | 40 |  |  | 200 |
| 03203-3851-4 |  |  | HUGO 1/60 |  |  | 60 |  |  | 300 |
| 03203-3931-6 |  |  | JUPITER 1/190 |  |  | 190 |  |  | 190 |
| 03203-3517-5 |  |  | CHICO I 1/80 |  |  | 80 |  |  | 400 |
| 03203-3518-3 |  |  | CHICO II 1/80 |  |  | 40 |  |  | 400 |
| 03203-3503-5 |  |  | CHICO III 1/40 |  |  | 40 |  |  | 200 |
| 03203-3519-1 |  |  | CHICO IV 1/80 |  |  | 80 |  |  | 400 |
| 03203-3504-3 |  |  | CHICO V 1/70 |  |  | 70 |  |  | 350 |
| 03203-3505-1 |  |  | CHICO VI 1/70 |  |  | 70 |  |  | 350 |
| 03203-3520-5 |  |  | CHICO VII 1/120 |  |  | 120 |  |  | 600 |
| 03203-3521-3 |  |  | CHICO VIII 1/80 |  |  | 80 |  |  | 400 |
| 03203-3506-K |  |  | CHICO IX 1/30 |  |  | 30 |  |  | 150 |
| 03203-3507-8 |  |  | CHICO X 1/20 |  |  | 20 |  |  | 100 |
| 03203-3522-1 |  |  | CHICO XI 1/40 |  |  | 40 |  |  | 200 |
| 03203-3526-4 |  |  | CHICO 15 1/60 |  |  | 60 |  |  | 300 |
| 03203-3527-2 |  |  | CHICO 16 1/40 |  |  | 40 |  |  | 200 |
| 03203-3529-9 |  |  | CHICO 18 1/120 |  |  | 120 |  |  | 600 |
| 03203-3858-1 |  |  | MARANCEL 1-40 |  |  | 40 |  |  | 190 |
| 03203-3859-K |  |  | MARANCEL 2 1-39 |  |  | 39 |  |  | 195 |
| 03203-3819-0 |  |  | LLANO 3 1/20 |  |  | 20 |  |  | 100 |
| 03203-3853-0 |  |  | VACA8 1/10 |  |  | 10 |  |  | 50 |
| 03203-3854-9 |  |  | VACA 10 1/20 |  |  | 20 |  |  | 100 |
| 03203-3855-7 |  |  | VACA 11 1/80 |  |  | 60 |  |  | 300 |
| Total |  |  |  |  |  | 4,105 |  |  | 19,955 |

The Cerro Casale deposit is entirely within the Nevado 1-840, Cachito 1, and Cachito 3-1298 exploitation concessions. Grasty Quintana & Cia (1997) confirmed CMC’s title to the Nevado and Cachito concessions by means of the ownership of the concessions by Compañía Minera Aldebarán. As part of the mineral patenting process, all claim monuments are surveyed by a licensed Chilean mining surveyor.

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4.2.2 | Surface Rights

There presently are no active agreements for use of surface rights. All surface rights are owned by the government of Chile, which generally assigns mining uses to a high priority.

4.2.3 | Water Rights

The information on water rights herein contained has been obtained from available documents referenced in Section 22 as well as the opinion of Horst Altschwager and Flavio Fuentes of CMC.
CMC reportedly owns water rights in three different areas including Piedra Pomez, Pedernales and Cerro Casale.
The Piedra Pomez area is located approximately 120 km north of Cerro Casale. Applications have been filed for use of groundwater from Piedra Pomez and rights have reportedly been granted and permits secured for a total amount of 1,237 L/s from 17 well sites. Water from Piedra Pomez is destined as the prime source of water for the Cerro Casale Project.
The Pedernales area is located approximately 70 km north of Piedra Pomez. Applications have also been filed to obtain groundwater from the Pedernales area. The submission contained seven applications for groundwater rights from seven production wells. Groundwater rights for a total amount of 510 L/s have been granted and permits have been secured. Reference to these water permits can be found in the Purchase and Sale Agreement between Barrick, Arizona Star and Bema. This water was in the name of PDLA therefore these were part of the sale of Placer Dome (Barrick’s) assets. Water from Pedernales is not destined for use but will rather be kept as a backup source meant to provide additional water to the project should it be required during the mine life.
Surface water rights have reportedly been granted in the immediate Cerro Casale area and a permit obtained for 50 L/s from Río La Gallina. Three other applications (one for 130 L/s on Río La Gallina and two for 180 L/s each on Río Nevado) have also been filed for surface water use in the Cerro Casale area but their status is unknown to AMEC. Groundwater rights were reportedly granted for a total of 33 L/s to be obtained from three production wells identified as PA-18 and M3 located along the Río Nevado Creek and PA-11 located at Pircas Negras. Water right applications for this area were originally denied because DGA (Dirección General de Aguas), the responsible authority, considered the area as headwaters to the Copiapo River, which was subject to prohibition by virtue of DGA N o 193 dated 27/05/93. However, the applications were reconsidered following submission of a legal recourse based on the interpretation of DGA N o 232 of 07/06/94 which provided an exemption for headwaters from sub-basins located more than 35 km away from the Copiapo River. The water rights still have not yet been formally granted by the DGA.

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4.2.4 | Conveyance Rights of Way

There are no existing impediments to obtaining easements for rights of way for access roads, water pipelines, or concentrate pipelines.

4.3 | Royalties

Minera Anglo American Chile Limitada and its affiliates are owed a royalty from production from the Cachito and Nevado mining concessions, which cover all of the Cerro Casale deposit. The royalty is capped at US$3.0 million and varies on the following sliding scale, which is dependent on the gold price:

| · | $425 to $474/oz – 1.0% NSR

| · | $475 to $524/oz – 1.5% NSR

| · | $525 to $599/oz – 2.0% NSR

| · | $600/oz and greater – 3.0% NSR.

AMEC has included a $3 million royalty payment in the sensitivity analyses to account for this royalty.

4.4 | Other Costs

The Purchase and Sales agreement with Barrick requires two payments totaling $80 million at the time a construction decision is made. The entire $80 million is included in AMEC’s sensitivity analyses as an acquisition cost to be paid by Bema and Arizona Star at the construction decision date, as opposed to a royalty to be paid by CMC.

4.5 | Environmental Exposures

Based on AMEC’s review of the project in 2005, five items have been identified as potential environmental exposures. The first two relate to simple administrative matters while the last three will require additional study to confirm to a level necessary to begin operations. These have not changed and are:

| 1. | Environmental approval of power supply infrastructure

| 2. | Environmental approval of port facilities

| 3. | Acid rock drainage (ARD) potential

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| 4. | Impacts on surrounding water systems from water supplies removed from the Piedra Pomez well field

| 5. | Downstream impacts from operation of tailing impoundment and waste rock dump facilities.

4.5.1 | Environmental Approval of Power Supply Infrastructure

Energy supply contracts that include requirements for contractors to have all the necessary permits and approval in place have now expired and will need to be renegotiated. It is reasonable to assume that permits will be granted for power lines, but these cannot be applied for until a power line system is designed.

4.5.2 | Environmental Approval of Port Facilities

The Cerro Casale Project proposes to use existing port facilities currently operated by Compañía Minera Candelaria (Candelaria). Under Chilean Law, responsibility to obtain the necessary environmental approvals and permits resides with the facility owner/operator. As the selected port facility has been operating for a number of years, an environmental approval has already been obtained. However, any modification to the existing port configuration or operation mode will require that a review be conducted by environmental authorities. Supporting documentation will thus have to be filed by Candelaria. Based on that scenario, the timing and ability for CMC to use the existing port facilities will depend on the terms and conditions negotiated with Candelaria. It is reasonable to assume that a contract will be negotiated and that approval for port modifications will be obtained.

4.5.3 | Acid Rock Drainage (ARD) Potential

Information on the potential for ARD at the Cerro Casale Project is presented in a document entitled “Phase 1 and 2 Work on Prediction of Drainage Chemistry” prepared by the Minesite Drainage Assessment Group in October 1999. AMEC reviewed the Executive Summary included in the 2000 Cerro Casale Project Feasibility Study Appendix (PDTS, 2000). AMEC also reviewed an interoffice memorandum prepared by Keith Ferguson on 27 October 1999.
The information presented in Keith Ferguson’s memo indicates that despite “a considerable amount of work on the potential for acid generation/metal release that has been conducted over the past two years”, there is “still significant uncertainty as to whether the wastes will in fact produce ARD/leach metal or even produce any drainage”. The memorandum further stipulates that “any drainage from the Aldebarán (previous name of the current Cerro Casale project) waste will at least contain elevated concentration (over 1,000 mg/L) of sulphate. Whether this would also contain elevated metal concentrations such as copper and zinc is yet to be fully determined but there is certainly a risk.” The memo goes on to summarizing recommendations for priority work for the next phase of environmental studies. These recommendations include conducting additional studies to confirm or refine conclusions reached to date in the ARD assessment work.

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AMEC recommends that further evaluation of ARD potential be performed in order to reduce the level of uncertainty associated with the currently available ARD assessment. Until further information is developed, AMEC considers ARD as a potential environmental exposure for the Cerro Casale Project.

4.5.4 | Impacts on Surrounding Water Systems from Water Supply Operations Conducted at the Piedra Pomez Well Field

A groundwater exploration program was developed in the area of Quebrada Piedra Pomez in order to evaluate the potential for use as a source of water for the Cerro Casale Project. The study was conducted by EDRA (Exploración y Desarrollo de Recursos de Agua S.A.) from 1997 to 1999. The study methodology included pump tests and water quality testing.
The Piedra Pomez basin is identified as an endorreic system and information presented in conclusion to the study supports this classification by indicating that neighboring surface water systems including Río Lamas, Río Qb. Barrancas Blancas and Río Qb. Penas Blancas as well as Salar de Maricunga are not connected based on results of geochemical analysis.
However, information contained in the study report tend to contradict the previous conclusion as it indicates that “the regional geology suggests that the hydrogeologic basin is larger that the hydrographic basin because the characteristics of the volcanic materials that filled the ancient valleys and changed the original landscape indicating that the topographic basin boundary does not represent a boundary for groundwater flow.” A cursory review of pump test results and water chemistry analysis presented in the EDRA report on “Hydrogeology of Quebrada Piedra Pomez” also suggests that the conclusion on the limited influence of the Piedra Pomez aquifer on surrounding water systems requires further evaluation. Nonetheless, The Dirección General de Aguas (DGA) has granted water use permits for 1,237 L/s from 17 well sites.

4.6 | Environmental Approvals and Permits

In accordance with legislative requirements of the Government of Chile described in Law N° 19.300 (Law on the General Basis on the Environment) and its regulations as outlined in Supreme Decree N° 30 (Regulation on the Impact Assessment System), environmental studies were conducted for the Cerro Casale Project and an Environmental Impact Study (EIS) was presented to the Regional Environmental Commission (COREMA) on 12 March 2001. Following a documented review process and presentation of additional support information, approval was granted by COREMA on 1 February 2002 through “Resolución Exenta N° 014.” Through this document, the Cerro Casale Project has thus obtained the main environmental authorization required under Chilean legislative requirements.

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The environmental approval granted to the Cerro Casale Project through “Resolución Exenta N° 014” outlines environmental commitments and requirements applicable to the project as a result of the EIS review process. Among other things, this document considers observations formulated by the public as well as to those expressed by regulatory authorities involved in the project environmental review. The nature and scope of commitments and requirements outlined in the project’s environmental authorization originate from programs and measures described in the EIS document and its addendums. Project development plans and future activities must therefore focus on compliance with specifications outlined in this environmental approval.
The next stage of legislative compliance process is outstanding and will require the project to seek sectorial permits granted by the various agencies that have authority over environmental resources and construction, operation and closure of project infrastructure.
The regional committee contains members of each applicable national Ministries and these members report to their national heads. Once COREMA approves the environmental plan for the project, permits for each operational area must be obtained from the relevant government agencies. These include:

| · | Servicio Nacional de Geología y Minería (SEMAGEOMIN): Mining permit, tailings dam construction and operating permit, waste dump construction and operating permit

| · | Superintencia de Servicios Sanitarios: Permits for water usage and for sewage and liquid industrial residue disposal.

| · | Servicio de Salud Regional: Responsible for worker and community health and safety. Provides operating permit which governs supply of potable water to camps and office, sewage treatment and waste disposal, including inflammable or explosive materials, or specific chemicals, tailing and cyanide handling and storage. Provides permit for operation of kitchen, first aid and medical facilities in both construction and operating stages.

| · | Dirección General de Aguas: Permits for construction and operation of reservoirs, aqueducts and pipelines. Permits for development and production from water wells.

| · | Servicio Agrícola y Ganadero: Permits for construction of site facilities and regulation of atmospheric emissions.

| · | Secretaría Regional del Ministerio de Vivienda y Urbanismo: Permits for construction of camp, administration and mine facilities.

| · | Dirección de Obras Municipales: General construction permits, in cooperation with the Secretaría Regional del Ministerio de Vivienda y Urbanismo.

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| · | President of the Republic: Permits for water purification and industrial waste treatment.

| · | Corporación Nacional Forestal (Conaf): Manages National Reserves. Will need to issue a permit for the water pipeline that crosses the Protected Area Ojos del Salado and the Nevado Tres Cruces National Park.

| · | Ministerio de Bienes Nacionales: Permits for water rights and water pipeline rights of way.

| · | Dirección de Vialidad - Ministerio de Obras Públicas: Permits for modifications of public roads and water crossings.

| · | Comisión Mixta de Agricultura y Urbanismo: Permit for change of surface land use from agriculture (standard use) to non-agricultural use.

| · | Dirección General de Obras Portuarias - Ministerio de Obras Públicas: Permits for construction of port facilities; approval for changes in existing permits.

| · | Armada de Chile: Permits for operation of port facilities and concessions for use of coastline as ports

| · | Superintendencia de Electricidad y Combustibles - Ministerio de Economía Fomento y Construcción: Permits for construction and operation of power and gas distribution lines.

| · | Consejo Nacional de Monumentos: Protection of heritage sites and regulation of relocation of cultural resources. Issues permits for construction of any facility close to heritage sites.

| · | Dirección del Trabajo: Permits for use of labour in construction and routine mining operations.

Although additional study is required for ARD potential from waste rock and potential downstream affects of tailings impoundments and additional costs may be incurred in remediation of any affects, AMEC is not aware of any significant environmental, social or permitting issue that would prevent exploitation of the deposit.

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5.0 | ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

5.1 | Accessibility

The Cerro Casale Project is located in the Maricunga mining district 145 km southeast of Copiapo, northern Chile (Figure 5-1). The project is within the geographic coordinates of 27° 47’ S and 69° 17’ W. The international border separating Chile and Argentina is located approximately 20 km east of the property.
Access to the project is 180 km by road from Copiapo. The initial southbound 25 km is paved highway, which connects to a 155 km gravel road running southeast to the project site. Currently, total driving time from Copiapo to site is approximately 3½ hours. The main dirt road serves as a regional transportation route to Argentina and is being gradually upgraded. A major portion of the route was recently upgraded as part of construction of the Refugio gold project, located north of Cerro Casale.
A regional airport and major supply services are located in Copiapo. Copiapo’s population is about 120,000. Commercial airline flights to Santiago and Antofagasta are available daily.
The terrain surrounding the Cerro Casale deposit is adequate for construction of administration, camp, mine, concentrator, tailings, and waste rock disposal facilities (Figures 5-2 and 5-3).
Surface rights are held by the national government, which normally provides surface use permits for mining operations as a priority use.

5.2 | Climate

The climate at Cerro Casale is typical for the northern Chilean Andes. Precipitation is generally limited to snowfall from April through September and rain is rare. Daytime temperatures in summer months approach 23°C, with night time lows of 5°C. Daytime temperatures in winter are around freezing, with night time temperatures dropping to -15°C.

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Figure 5-1: Location of Cerro Casale Project, Northern Chile
Figure 5-2: Mill Site

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Figure 5-3: Tailings and Waste Rock Site

5.3 | Local Resources

A skilled labor force is available in the Copiapo region and surrounding mining areas of northern Chile.
A source of electric power must be negotiated, but the current concept involves building a temporary power line from Refugio to the site for initial power supply. Long term power will be provided from a generation plant (yet to be permitted and built) near Cardones.
Suitable water supply is available from the presently permitted Piedra Pomez well field, located 121 km north of the project.
Fuel and supplies will be provided from nearby communities such as Copiapo.

5.4 | Infrastructure

Cerro Casale is a green field site. As such, existing site infrastructure is limited to an exploration camp and roads.

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5.5 | Physiography

The Cerro Casale project is in the northern Chilean Andes within an area of high relief. The Río Nevada valley immediately east of the present exploration camp is at an elevation of 3,800 m. The top of Cerro Casale, in the middle of the deposit, is 4,450 m. Other mountains rise to the north and east. The top of Volcan Jotabeche, 10 km north of Cerro Casale, is approximately 5,800 m.
Vegetation is sparse and generally restricted to small plants, mostly along streambeds and river courses.
Wildlife includes guanaco, vicuña, foxes, rabbits, ground squirrels, hawks, condors, and small reptiles.

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6.0 | HISTORY

Anglo American first explored the Aldebarán area in the late 1980’s, drill testing multiple areas of alteration. Anglo American drilled two holes in the Cerro Casale deposit in 1989.
In 1991, Anglo American conveyed its interests in the Cerro Casale property to Compañía Minera Estrella de Oro Limitada (CMEO) and Compañía Minera Aldebarán (CMA), two companies presently owned by Bema and Arizona Star Resource Corporation (Arizona Star), both being members of the legal entity at that time, the Bema Shareholders Group. CMA, on behalf of the Bema Shareholders Group, conducted exploration drilling from 1991 through 1997, targeting both oxide and sulphide gold-copper mineralization. In 1997, Bema completed a feasibility study for development of oxide gold-copper mineralization, a prefeasibility study for an oxide-sulphide operation and a scoping study for development of deep sulphides.
In 1998 PDI through its subsidiary Placer Aldebarán (Cayman) Limited and the Bema Shareholder Group established CMC to continue exploration and development of various gold-copper deposits in an area of interest covering the known gold-copper mineral occurrences in the Cerro Casale area.
Placer Dome Latin America (PDLA) as General Manager of the project continued drilling in 1998 and 1999, leading to completion of a feasibility study in 2000. Work in 1998 included property-wide geological mapping, ground and airborne magnetic surveys and Audio Frequency Magnetic Telluric surveys (AMT). Capital and operating costs were updated by Placer Dome in March 2004.
In 2005, AMEC prepared the 2005 Technical Report documenting the extent to which resource estimation work was performed in accordance with industry standard practices and verifying that mineral resource and mineral reserve estimates in the 2004 Feasibility Study Update were compliant with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000), CIM Definition Standards for Mineral Resources and Mineral Reserves (2005) and Canadian National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators.
In mid-2005, Bema secured the services of MQes to evaluate various mining, material handling, and processing alternatives to support commercial development of the Cerro Casale project.
In early 2006, Barrick Gold Corporation concluded the acquisition of Placer Dome Inc and its subsidiaries. In accordance with a Letter of Agreement dated June 19, 2006 between Bema Gold Corporation (Bema) and Arizona Star Resource Corporation (Arizona Star), Bema and Arizona Star acquired Barrick’s 51% interest in the shares of CMC. Pursuant to this Agreement, Bema and Arizona Star terminated the existing Amended and Restated Shareholder’s Agreement dated June 5, 2003.

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Bema commissioned preparation of this technical report in June 2006 to update the 2005 Technical Report with the results of MQes efforts.

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7.0 | GEOLOGICAL SETTING

7.1 | Regional Geology

The Cerro Casale gold-copper deposit is located in the Aldebarán subdistrict of the Maricunga Volcanic Belt (Figure 7-1). The Maricunga belt is made up of a series of coalescing composite, Miocene andesitic to rhyolitic volcanic centers that extend for 200 km along the western crest of the Andes. The volcanic rocks are host to multiple epithermal gold and porphyry-hosted gold-copper deposits, including Cerro Casale, Refugio, Marte, and La Coipa, as well as numerous other smaller mineral prospects. The volcanic rocks overlie older sedimentary and volcanic rocks of Mesozoic and Paleozoic age.
Reverse faults that strike parallel to the axis of the Andes have uplifted hypabyssal intrusive rocks beneath the extrusive volcanics, exposing porphyry-hosted gold-copper deposits in the Aldebarán area such as Cerro Casale, Eva, Jotabeche, Estrella and Anfiteatro (Figure 7-1). Composite volcanic centers formerly overlying the intrusive complexes are still preserved in the immediate Cerro Casale area at Volcan Jotabeche and Cerro Cadillal.
Structural interpretations from regional geological mapping and Landsat imagery show major fault systems cutting Paleozoic, Mesozoic and Tertiary units. The oldest set of faults strike NW and extend in this direction for 50 km to 60 km. These most likely are extension structures perpendicular to the direction of plate subduction. Major through-going lineaments trend NE and appear to mark boundaries between major lithological domains in basement rocks.
Younger lineaments and faults cut Tertiary and Quaternary volcanic rocks. These strike North, 040°, 310°, and East. Mineralization in individual deposits is generally aligned along one or more of these structural trends.
Major alteration zones, gold and gold-copper mineralization in the Maricunga Volcanic Belt are coincident with subvolcanic intrusive rocks of diorite and granodiorite composition. Intrusives generally occur at the intersection of major structural lineaments. The major alteration zones include La Coipa, Aldebarán (containing Cerro Casale) and Lobo-Amalia.

7.2 | District Geology

The Aldebarán area is underlain by extensive dacitic to andesitic volcanic and volcaniclastic rocks derived from Volcan Jotabeche and Cerro Cadillal. Numerous dioritic to granodioritic subvolcanic plutons related to the volcanic rocks crop out at Cerro Casale, Roman, Eva, Estrella and Anfiteatro (Figure 7-1).

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Figure 7-1: Geology of the Maricunga Volcanic Belt (PDTS, 2000)

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Extensive hydrothermal alteration consisting of quartz-feldspar veinlet stockworks, biotite-potassium feldspar, quartz-sericite, and chlorite occurs in these intrusive centers. Gold-copper mineralization is principally associated with intense quartz-sulphide stockworks, potassic alteration, and phyllic alteration.

7.3 | Cerro Casale Deposit Geology

7.3.1 | Introduction

The Cerro Casale deposit is exposed in a hill of approximate 700 m of vertical relief and 1 km in diameter. Mineralization is related to a series of dacitic to dioritic intrusives, which were emplaced into Miocene andesites and volcaniclastic sedimentary rocks. The Miocene volcanic rocks overlie Oligocene conglomerates, which in turn, overlie Eocene basaltic andesites and rhyolite pyroclastic flows.
Gold-copper mineralization occurs in quartz-sulphide and quartz-magnetite-specularite veinlet stockworks developed in the dioritic to granodioritic intrusives and adjacent volcanic wall rocks. Stockworks are most common in two dioritic intrusive phases, particularly where intrusive and hydrothermal breccias are developed. Mineralization extends at least 1,450 m vertically and 850 m along strike. The strike of mineralization follows WNW (310°) trending fault and fracture zones. The main zone of mineralization pinches and swells in width from 250 m to 700 m along strike and along dip steeply to the southwest. The highest-grade mineralization is coincident with well developed quartz-sulphide stockworks in strongly potassically altered intrusive rocks.

7.3.2 | Lithology

Lithologies important to mineralization and control of resource domaining are dominantly the multi-phase porphyries and related breccias, which intrude the flat-lying volcanic and volcaniclastic rocks. Ten rock units are relevant as ore controls for domaining in resource estimation (Table 7-1). Figures 7-2 and 7-3 show the distribution of these units at surface and in a typical geological section, looking west.
The volcanic-sedimentary sequence is split into four units: conglomerate, felsic air-fall tuff, mafic flow and rhyolite pyroclastic flow (youngest to oldest). The conglomerate is 350 m thick and is made up of red beds with heterolithic cobbles. This unit occurs between the 3750 m and 4100 m elevations. Beneath the conglomerates are well-bedded, felsic air-fall tuffs totaling 100 m. The tuffs overlie amygdaloidal andesite flows present between the 3400 m and 3650 m elevations. The andesites are strongly altered near later dioritic intrusions and are composed mostly of biotite, apatite, and plagioclase.

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Table 7-1: Major Lithological Units at Cerro Casale

| Major Category | Lithological Unit
| Intrusive-Related Breccias | Hydrothermal breccia
|  | Catalina breccia
|  | Microdiorite breccia
| Intrusive Porphyry Units | Biotite porphyry
|  | Granodiorite
|  | Diorite porphyry
| Volcanic-Sedimentary Units | Conglomerate (red beds)
|  | Felsic tuff
|  | Mafic volcanic flows
|  | Rhyolite pyroclastic flows

The oldest unit in the volcanic-sedimentary sequence is a thick section of rhyolite pyroclastic flows showing welded, eutaxitic structures characteristic of pyroclastic flows. This unit extends below the deepest drill holes, which end at an elevation of about 3000 m.
The intrusive porphyry units are dominated by an early-stage, laccolith-shaped body of diorite porphyry which forms the bulk of the Cerro Casale topographic high. The laccolith extends over a circular area of approximately 1 km by 1 km and down to the 3800 m elevation. The porphyry is comprised of approximately 40% plagioclase phenocrysts within in a fine-grained plagioclase matrix. The diorite porphyry is a host to gold-copper mineralization where quartz-sulphide stockworks are developed in around later granodiorite and micro diorite porphyry bodies and breccias.
A near vertical, tabular series of at least three granodiorite bodies cut the diorite porphyry along a WNW trend. The intrusives extend for at least 1 km along strike and are 100 m to 300 m wide. The granodiorite is comprised of 40% crowded phenocrysts of plagioclase, potassium feldspar, hornblende, and biotite. Phenocrysts are subhedral to euhedral. The groundmass is a fine-grained mixture of orthoclase, biotite, and minor quartz. The unit shows a range in alteration from weak sericitization of feldspars and biotite replacement of amphiboles, to intense potassium feldspar flooding of the groundmass with >20% quartz vein stockworks.

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Figure 7-2: Surface Geological Map of Cerro Casale (PDTS, 2000)

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Figure 7-3: Cross Section 850E, Looking Northwest, Cerro Casale Deposit (PDTS, 2000)

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Biotite porphyry is minor by volume but is closely related to mineralization in the upper portion of the deposit. This porphyry is characterized by coarse subhedral to euhedral biotite phenocrysts and may be a potassically altered phase of the granodiorite.
Breccia bodies dip steeply to the south to vertical and are strongly elongated WNW. The breccias are developed principally in the diorite porphyry along the north side of Cerro Casale, but also formed in the granodiorite. The highest gold-copper grades are generally associated with the breccias.
Micro diorite Breccia is a fine-grained, intrusive breccia that contains a variable percentage of angular to subrounded fragments of volcanic rocks. The microdiorite component is finely porphyritic with phenocrysts of plagioclase supported in a fine-grained matrix of orthoclase, biotite, anhydrite, magnetite/specularite and minor quartz. The breccia is strongly altered in all locations and cuts the diorite porphyry along the upper north side of Cerro Casale.
The Catalina Breccia is adjacent to the microdiorite breccia and is thought to be a sulphide-rich phase of the latter. The Catalina Breccia forms a cone-shaped body in the centre of the mineral deposit and is characterized by its matrix of anhydrite, gypsum, barite, tourmaline, rhodochrosite, dolomite, chalcopyrite, pyrite, galena, and sphalerite. In small restricted areas, the breccia contains very high-grade stockworks with up to 13% Cu and 200 g/t Au.
Hydrothermal breccias are common at contacts between diorite porphyry and microdiorite breccia. These occur as porphyry with intense quartz-sulphide stockworks, open spaces and framework-supported rock fragments set in a matrix of quartz-sericite-specularite. The hydrothermal breccias generally occur high in the deposit and grade outward to pebble dikes.
Limited overburden occurs in the immediate area of Cerro Casale, where bedrock is covered by a thin veneer of residual soils. Colluvium and alluvium up to 30 m thick are present in the Río Nevada valley

7.3.3 | Structure

Major fault and fracture zones trend NE and WNW within the Aldebarán district. Cerro Casale and the other mineral occurrences in the Aldebarán area occur at the intersection of these structural zones, showing a structural control to the emplacement of the subvolcanic intrusives and associated mineralization.
Within each deposit and in particular within Cerro Casale, gold-copper bearing quartz-sulphide stockwork zones are strongly elongated along azimuths ranging from 110° to 140° and dip vertically to steeply south. This elongation is coincident with the geometry of the granodiorite intrusives and with the enclosing alteration zone. The alteration zone is up to 1 km wide and 6 km long.

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Topographic lineaments suggest the presence of a third, steeply dipping fault and fracture system on the north side of Cerro Casale that trends 035° to 050°. The Catalina Breccia is located at the intersection of this structure and the WNW stockwork zones.

7.3.4 | Weathering and Oxidation

Oxidation resulting from weathering and/or high oxygen activity in the last phase of hydrothermal alteration overprints sulphide mineralization in the upper portion of the Cerro Casale deposit. Oxidation locally extends deeply along fault zones or within steeply dipping breccia bodies. Placer Dome mapped three types of oxidation states:

1. | zones where ≥90% of the original sulphides are preserved (sulphide)

2. | zones where between 10% and 90% of the original sulphide is preserved (mixed)

3. | zones where less than 10% of the original sulphides remain (oxide).

The depth of oxidation is dependent on the permeability of the altered rock and the presence of high-angle structures. Oxidation generally goes no deeper than 15 m where vertical structures are absent. Oxide is present in linear oxidation zones as deep as 300 m along major fault and fracture zones, or as pendants along the intersection of multiple fault zones (Figure 7-4). Locally there are large blocks of less permeable sulphide material within the oxide zones.

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Figure 7-4: Redox Units, Section 850E (PDTS, 2000)

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8.0 | DEPOSIT TYPES

Gold-copper mineralization at Cerro Casale formed during emplacement of multiple phases of diorite and granodiorite intrusions into a coeval sequence of intermediate to felsic volcanic rocks. Mineralization appears to be most closely related to strong potassic to phyllic alteration of the latest phases of intermediate to felsic intrusives and associated intrusive and hydrothermal breccias. Mineralization is focused in well developed quartz-sulphide stockworks which dip vertically to steeply south and strike WNW. These stockworks and potassic alteration formed during the latest phase of emplacement of the granodiorite as the result of degassing of the intrusion. Fluid pressures broke wall rocks and the upper portion of the granodiorite, forming the microdiorite and hydrothermal breccias. In this regard, the Cerro Casale deposit is a primary gold-copper porphyry with strong affinities to high sulphidation, volcanic-hosted gold systems.

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9.0 | MINERALIZATION

9.1 | Introduction

Gold-copper mineralization associated with Tertiary volcanic rocks and subvolcanic plutons is present in at least eight sites within the Aldebarán district. Cerro Casale is the largest deposit and has been drilled to a detail suitable for estimation of resources and reserves. Mineralization in the district is present where stockworks of quartz-sulphide veins and veinlets have developed in felsic intrusive rocks, intrusive breccias, hydrothermal breccias and volcanic wall rocks. Mineralization is related to degassing of late-stage plutons and development of high-temperature, potassic alteration in the plutons and wall rocks.
Figure 9-1 shows the major gold-copper occurrences on the Aldebarán property and the outline of mining claims that constitute the property. From the northeast, these include Jotabeche, Romancito, Cerro Roman, Eva, Anfiteatro, Cerro Casale, Cerro Catedral, and Estrella.
Exploration drilling is sufficiently advanced at Eva and Cerro Roman to obtain preliminary estimates of resources. Mineral resources for Cerro Roman are classified by Placer Dome as Inferred. Work at Jotabeche, Romancito, Anfiteatro, Cerro Catedral, and Estrella is not sufficient for estimation of gold or copper mineral resources. AMEC did a cursory review of the geology of these satellite deposits but did not verify exploration data and resource estimates.

9.2 | Cerro Casale Deposit

9.2.1 | Alteration

Alteration consists of a zoned, subcircular pattern surrounding the centre of the most pervasively altered diorite porphyry, granodiorite, and intrusive breccias. The outer portion of the system is propylitic alteration in diorite porphyry and volcanic wall rocks characterized by quartz, chlorite, pyrite, sericite, clay, and minor epidote. Mafic minerals are replaced by chlorite and minor magnetite and plagioclase is altered to sericite and clay.
Phyllic alteration is present in most of the diorite porphyry and granodiorite. At least two phases of phyllic alteration may be present. Plagioclase and mafic minerals are replaced with sericite and quartz. Disseminated specularite is locally present. Deep in the deposit there is an early phase of phyllic alteration after which sericitized plagioclase phenocrysts are surrounded with secondary potassium feldspar. In the upper portion of the deposit the phyllic alteration is more extensive, converting most of the diorite porphyry, Catalina Breccia and granodiorite to quartz, sericite, pyrite and tourmaline.

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Figure 9-1: Major Gold-Copper Occurrences in the Aldebarán Property (PDTS, 2000)

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The centre of the alteration system is coincident with gold-copper mineralization and is comprised of intense potassium silicate alteration. Biotite replaces hornblende as aggregates of biotite books and magnetite.
The biotite zone forms a 200 m diameter halo around a core zone of strong potassium feldspar alteration. Potassium feldspar halos in quartz-sulphide veinlets become more frequent towards the center of the system where all plagioclase is totally replaced by secondary orthoclase. Primary textures are obliterated. Argillic alteration is restricted to base-metal veins peripheral to Cerro Casale at Zona de Veta and Cerro Catedral. The argillic alteration forms halos to quartz, alunite, kaolinite, and pyrite veins.
Stockwork vein composition varies. The following types are present:

· | gypsum

· | quartz-limonite/hematite

· | quartz-specularite

· | pyrite (with argillic haloes)

· | anhydrite-gypsum-barite-rhodochrosite-pyrite-chalcopyrite-sphalerite-galena

· | quartz-specularite-pyrite

· | gypsum-pyrite

· | potassium feldspar-quartz ± sulphides

· | quartz-magnetite-chalcopyrite-bornite

· | magnetite-chalcopyrite-bornite ± chlorite

· | biotite + minor magnetite

· | quartz-anhydrite-chalcopyrite.

Gold-copper mineralization is most commonly associated with quartz-limonite/hematite, quartz-specularite-pyrite, potassium feldspar-quartz-sulphide, quartz-magnetite-sulphide and quartz-anhydrite-sulphide veinlets. Veinlets are from 1 mm to 10 mm wide. Sulphides occur disseminated in the vein matrix or along vein margins. Veinlet frequency ranges from none in the latest intrusive phases to more than 35% by volume around the contacts between the granodiorite, microdiorite breccia, and diorite porphyry.

9.2.2 | Mineralization

Gold and copper mineralization is most directly associated with quartz-sulphide-magnetite stock work veins and veinlets in potassically altered rocks. Mineralization extends from the surface of the north side of Cerro Casale at an elevation of 4200 m to the base of existing drilling at 3000 m. Mineralization extends for about 850 m along strike to the WNW, dips vertical to 75° south, and is from 150 m to 700 m wide. The thickest portion of the mineralization is at the 3800 m elevation. Figures 9-2 and 9-3 show typical cross sections of the gold and copper grades across the centre of the deposit. Figures 9-4 and 9-5 show plan views of gold and copper grades in the core of the deposit at the 3800 m elevation.

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Figure 9-2: Measured + Indicated Gold Resources, Section 472200E (MQes, 2006)

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Figure 9-3: Measured + Indicated Copper Resources, Section 472200E (MQes, 2006)

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Figure 9-4: Measured + Indicated Gold Resources, 3832 Elevation (MQes, 2006)

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Figure 9-5: Measured + Indicated Copper Resources, 3832 Elevation (MQes, 2006)

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Figure 9-6: Intensity of Stockwork Veining, Section 850E (PDTS, 2000)

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Figure 9-7: Potassium Feldspar Alteration, Section 850E (PDTS, 2000)

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Gold and copper grades have a high correlation. Cross cutting relationships with host rocks suggest a maximum age of 13.5 Ma (PDTS, 2000, accuracy limits not stated). Fluid inclusion work suggests a temperature of formation close to 500°C. Limited petrographic work suggests that a large portion of the gold is free and present along the margins of pyrite grains. Gold particles found in the Catalina Breccia (the highest grade unit) range from 1 µm to 145 µm, with a mean of 39 µm.
Hypogene copper minerals include chalcopyrite, bornite, and chalcocite-djurleite (Cu 3 S) and minor copper silicate minerals. Secondary copper minerals in the oxide and mixed zones include chalcocite, digenite, covellite, chrysocolla, malachite, and minor copper silicates. Most copper sulphides are in stockwork veinlets rather than disseminated in wall rocks. Locally disseminated chalcopyrite is present in the granodiorite. Disseminated copper zones are low in gold. Bornite increases with depth, corresponding with the highest copper grades below the 3800 m elevation.
Copper is depleted in the oxide zone, being generally less than 0.10% in the upper portion of the deposit. There are sporadic supergene enriched copper zones where chalcocite is present in volcanic rocks and mixed sulphides in intrusive rocks. These rarely persist laterally more than 200 m.
Gold distribution does not appear to be impacted in the oxide zone.
Gold-copper mineralization is strongly related to the presence of diorite, granodiorite, breccia units and the intensity of stockwork veining and potassic alteration. Figures 9-6 and 9-7 show the distribution of stockwork veining and potassic alteration, respectively. Mineralization is related to moderate to strong stockwork veining and moderate to strong potassium feldspar alteration.
The average silver:gold ratio is 3:1. Silver was not obtained for all drilling samples and was not estimated in the resource block models.

9.3 | Eva Deposit

9.3.1 | Geology

Eva is located 5 km northwest of Cerro Casale at a surface elevation of between 4600 and 4900 m. Gold-copper mineralization found to date is in two west-trending zones called Eva Norte and Eva Sur. These zones are 500 m apart. Both extend approximately 800 west and 200 m north.

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Figure 9-8: Geological Map of the Eva Deposit (PDTS, 2000)

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Figure 9-9: Cross Section of Eva Deposit (PDTS, 2000)

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Westward-elongated bodies of quartz monzonite, intruded by later biotite and amphibole-rich dacite porphyry are the focus of alteration and mineralization (Figures 9-8 and 9-9). The quartz monzonite and dacite porphyry intrude relatively flat-lying andesitic to dacitiic flows and volcanic breccias. Hydrothermal breccias occur in the dacite porphyry and are comprised of dacite porphyry fragments and quartz veins set in a fine-grained matrix of quartz, sericite, and chlorite. Pebble dikes are locally present.
The dominant fault and fracture systems strike 290° to 310° and dip approximately 70° south.

9.3.2 | Alteration and Mineralization

Gold and copper values increase where the dacite porphyry, quartz monzonite and volcanic wall rocks are strongly silicified either as replacement of groundmass or as development of quartz-sulphide stockworks. Disseminated magnetite is common. Potassic alteration is generally fine-grained biotite in silicified and sericitized rock and is only rarely present as secondary potassium feldspar.
Gold mineralization generally increases with the frequency of quartz-sulphide stockworks, but can be anomalous in zones with disseminated sulphides.

9.4 | Cerro Roman

9.4.1 | Geology

Figures 9-10 and 9-11 show the surface geology and a typical cross section of the Cerro Roman deposit. Cerro Roman contains porphyries and breccias intruding andesitic to dacitic volcanic rocks in a setting similar to Cerro Casale. The plutons include an early diorite porphyry, followed by quartz diorite porphyry and then dacite porphyry. The plutons are elongated along W and WNW-trending fracture patterns, showing active extensional structures at the time of their emplacement. Late-stage intrusive breccias occur along the margins of the central quartz-diorite porphyry. Hydrothermal brecciation occurs in all intrusive units and in volcanic wall rocks.

9.4.2 | Alteration and Mineralization

Alteration is comprised of a zone of potassic alteration centered on the porphyries, surrounded by a marginal potassic zone and an outer propylitic zone. The entire alteration system is about 500 m by 700 m in plan and extends to the vertical limit of drilling (360 m). The central potassic zone contains well-developed quartz-sulphide veinlets with biotite and potassium feldspar replacement of mafic minerals and plagioclase, respectively. The marginal potassic zone is developed mostly in andesitic wall rocks and is expressed by development of pyroxene, biotite, and magnetite. Propylitic alteration is developed mostly in volcanic wall rocks and is comprised of quartz and chlorite.

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Figure 9-10: Geological Map of the Cerro Roman Deposit (PDTS, 2000)

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Figure 9-11: North-South Cross Section of Cerro Roman Deposit (PDTS, 2000)
Gold-copper mineralization is directly related to the frequency of quartz-magnetite-sulphide veinlet stockworks developed in the intrusive units and adjacent andesite wall rocks. Sulphides include pyrite, chalcopyrite, and bornite. The highest grades occur where dense veinlet stockworks occur along the margins of the central quartz diorite and in breccias. Mineralization occurs within an area 600 m long east-west by 300 m wide north-south. Within this area individual zones of > 0.8 g/t Au are present, separated by envelopes of lower grade mineralization. At least three bodies of the higher-grade mineralization are from 120 to 350 m long and 60 to 150 m wide.
Copper grades are generally low, averaging less than 0.2%.

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9.5 | Estrella Prospect

9.5.1 | Geology

The Estrella area is underlain by relatively flat-lying intermediate volcanic rocks and flow breccias, and by irregular, sill-like porphyry intrusions. The volcanic rocks are andesite and dacite. The subvolcanic sills are coeval with the volcanic rocks and vary from dacite to andesite porphyry. Hydrothermal breccias composed of andesite and dacite fragments set in a matrix of quartz, magnetite and sulphides are developed along high-angle structures that strike NNW. Other hydrothermal breccias are flat-lying and are made up of fragments of andesite and dacite in a matrix of gypsum.
Fault and fracture systems are well developed along four directions. Small-scale faults and fractures strike 350° and 70°. The NNW set appear to influence the development of vertical hydrothermal breccias. More dominant faults trending 50° and 120° cut the smaller features.

9.5.2 | Alteration and Mineralization

Alteration related to gold mineralization consists of pervasive silicification and quartz veining in hydrothermal breccias. Subparallel veins strike NNW and NE. Quartz veins contain magnetite, pyrite, and locally chalcopyrite.
Limited drilling to date suggests gold mineralization is restricted to relatively narrow, sheeted quartz vein systems.

9.6 | Anfiteatro Prospect

9.6.1 | Geology

Flat-lying dacitic to andesitic volcanic flows and flow breccias underlay the Anfiteatro area. The volcanic rocks are intruded by a series of andesitic to dacitic porphyries. The intrusives are composed of plagioclase, quartz and amphibole phenocrysts set in a microcrystalline matrix of plagioclase, secondary biotite, potassium feldspar, amphiboles and quartz. Within the porphyries are intrusive and hydrothermal breccias. Intrusive breccias are comprised of fragments of andesite or dacite porphyry set in a fine-grained matrix altered to chlorite and epidote. Hydrothermal breccias are made up of fragments of porphyry and volcanic rocks in a matrix of quartz, potassium feldspar, pyrite, gypsum, and locally sphalerite.
Fault and fracture systems are dominated by fracture zones and quartz veins that strike 060°.

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9.6.2 | Alteration and Mineralization

Potassic alteration manifested by secondary biotite and local quartz, potassium feldspar and chlorite is present within the porphyries. Gold mineralization is associated with potassic alteration and stockwork veins of quartz, potassium feldspar, biotite, sericite, pyrite, chalcopyrite, and magnetite. The Stockwork Zone within Anfiteatro is an area of stockwork veining 600 m long by 250 m wide in dacitic to andesitic volcanic flows. Veinlets are dominantly quartz, magnetite, and specularite. Mineralization in the Ojo de Buey dacite porphyry is comprised of quartz-magnetite veinlets with limonite and copper oxides.
Soil geochemistry shows average surface gold values of 0.25 g/t and 0.10 g/t in the Stockwork and Ojo de Buey areas, but drilling to date has been relatively negative with the best intercept being 150 m of 0.46 g/t Au in the Stockwork Zone in CMA hole ANF-02. Soil sampling shows up to 0.46 g/t Au in an area 100 m by 150 m at Anfiteatro Zona 10 and up to 0.26 g/t in an area 120 m by 300 m at Anfiteatro Alto. These soil geochemical anomalies have not been drill tested.

9.7 | Romancito Sur

9.7.1 | Geology

An intermediate intrusive porphyry cuts a sequence of intermediate volcanic breccias at Romancito Sur. The volcanic breccias dip 30° to the south. The porphyry strikes west and appears to have followed district-scale fracture zones. Hydrothermal breccias cross-cut the volcanics and porphyry and are composed of fragments of volcanic rocks set in a fine-grained, silicified matrix. Quartz-sulphide veins and stockworks strike ENE, following the trend of the intermediate porphyry.

9.7.2 | Alteration and Mineralization

Porphyry and volcanic rocks are variably silicified, with alteration increasing with proximity to individual quartz veins and stockworks. Silicified rocks also show chloritization of mafic minerals, sericitization of plagioclase and disseminated magnetite and pyrite. Anomalous gold values are associated with the most intensely silicified and veined zones where sulphides are present.
Faults are strongly argillized but this alteration is late and does not appear to be associated with gold mineralization. Potassic alteration is rare.
Gold mineralization >0.5 g/t is associated with a 20 m to 30 m wide zone of quartz-sulphide veins and stockworks that strikes at 70° across the centre of the prospect. Rock chip samples collected from trenches in this area returned gold values up to 2.12 g/t. One third of 247 samples grade greater than 0.5 g/t. Two core holes drilled within this zone; however, returned relatively narrow and discontinuous intercepts.

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9.8 | Other Areas

Surface sampling and drilling at Jotabeche, Zona de Vetas, and Cerro Catedral (Figure 9-1) by Anglo American and Bema Gold Corporation revealed weak zones of gold-copper mineralization that did not warrant additional drilling. Placer Dome did not continue exploration in these areas in 1999 because of negative results.

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10.0 | EXPLORATION

Between the late 1980s and 1999, the Aldebarán area containing Cerro Casale was explored by Anglo American, Bema Gold Corporation, Arizona Star Resources, and Placer Dome. Anglo American drilled core holes at Cerro Casale in the late 1980s following up on alteration anomalies exposed in the rugged terrain. After acquiring the property from Anglo American in 1993, Bema Gold and Arizona Star proceeded in a comprehensive program that included interpretation of Landsat imagery, geological mapping, surface rock-chip sampling, surface geophysical surveys and RC and core drilling. This work continued until Placer Dome entered into an agreement with Bema Gold and Arizona Star in 1998. In the following two years, Placer Dome continued with core drilling at most of the mineralized prospects in the Aldebarán area. This work culminated in a feasibility study on the Cerro Casale deposit in early 2000.

10.1 | Cerro Casale

Anglo American conducted limited geological mapping and drilled two RC holes at Cerro Casale in 1989. The Bema Shareholders Group acquired the property in 1991 and one of its subsidiary companies, Compañía Minera Aldebarán (CMA), began an aggressive program of RC and core drilling. From 1991 to 1997 CMA drilled 224 RC holes totaling 43,317 m and 88 core holes totaling 54,905 m. CMA also undertook geological mapping, surface rock-chip sampling and Bleg soil sampling throughout the district.
Placer Dome continued drilling in 1998 and 1999, leading to completion of a feasibility study in 2000. Work in 1998 included property-wide geological mapping, ground and airborne magnetic surveys and Audio Frequency Magnetic Telluric surveys (AMT). Placer Dome also drilled 30 core holes totaling 23,924 m.

10.2 | Eva

CMA discovered Eva during follow up of Bleg soil and stream sediment sampling in 1993. CMA performed geological mapping, collected 1,200 rock samples, and drilled 37 RC holes totaling 4,574 m from 1993 to 1997. Placer Dome completed airborne magnetic and surface AMT surveys, performed geological mapping, trench, and road-cut sampling and drilled seven core holes in 1998. Placer Dome drilled an additional seven core holes in 1999 for a total of 5,914 m.

10.3 | Cerro Roman

Bema and Arizona Star discovered Cerro Roman in 1993 during reconnaissance geological mapping. From 1994 to 1997, CMA took 1,500 rock-chip samples from surface exposures, performed 1,300 m of trenching, performed surface magnetic and Induced Polarization surveys and drilled 41 RC holes totaling 7,250 m.

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Placer Dome continued exploration in 1998 and drilled 7,207 m of core in 18 holes. Placer Dome also performed geological mapping, trench sampling, rock-chip sampling, and surface AMT surveys.

10.4 | Estrella

Bema/Arizona Star mapped quartz-vein gold-copper mineralization in volcanic rocks and hydrothermal breccias here in 1992, following up on soil geochemical anomalies found by Anglo American in the mid 1980s. In 1997, CMA drilled 24 RC holes totaling 3,378 m. Placer Dome remapped the area in 1998 and trenched obvious areas of alteration. In 1999 Placer Dome drilled four core holes totaling 1,225 m.

10.5 | Anfiteatro

Anglo American performed geological mapping and rock-chip sampling in 1985 and 1986. Between 1992 and 1994, CMA completed detailed geological mapping, surface rock-chip sampling and drilled four RC holes totaling 536 m. Placer Dome drilled three core holes totaling 998 m in 1990.

10.6 | Romancito

Limited exploration work has been completed at Romancito. Regional mapping performed by Placer Dome in 1998 identified the area to be potentially mineralized. Limited rock chip sampling revealed anomalous gold values. In 1999, detailed geological mapping, trenching, rock-chip sampling, and drilling was performed. Two core holes totaling 794 m were drilled.

10.7 | Other Areas

Other areas such as Zona de Vetas and Cerro Catedral have produced few significant results in drilling and sampling.

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11.0 | DRILLING

Reverse-circulation (RC) and core drilling was performed in multiple campaigns since 1989. Anglo American drilled two RC holes in 1989. The Bema Shareholder Group drilled a large number of RC and core holes between 1991 and 1997. Placer Dome drilled additional confirmation, infill, and geotechnical core holes in 1998 and 1999.
Table 11-1 lists drill holes by type, number and total length by year and company. Figure 11-1 shows collar locations and downhole projections of holes, coded by drill campaign.
Table 11-1: Cerro Casale Drilling

| Year |  | Company |  | Type |  | Purpose |  | Holes |  | Meters |
| 1989 |  | Anglo American |  | Core |  | Exploration |  |  | 2 |  |  | 601 |
| 1991 |  | Bema |  | RC |  | Exploration |  |  | 20 |  |  | 1,980 |
| 1992 |  | Bema |  | RC |  | Exploration |  |  | 13 |  |  | 1,670 |
| 1993 |  | Bema |  | RC |  | Exploration |  |  | 22 |  |  | 2,700 |
| 1993 |  | Bema |  | Core |  | Metallurgy |  |  | 6 |  |  | 464 |
| 1994 |  | Bema |  | RC |  | Exploration |  |  | 31 |  |  | 4,517 |
| 1995 |  | Bema |  | RC |  | Feasibility Infill |  |  | 67 |  |  | 13,479 |
| 1995 |  | Bema |  | RC |  | Condemnation |  |  | 11 |  |  | 1,076 |
| 1995 |  | Bema |  | Core |  | Geotechnical, Geostatistical |  |  | 11 |  |  | 2,740 |
| 1996 |  | Bema |  | RC |  | Deep Oxide Exploration |  |  | 20 |  |  | 8,139 |
| 1997 |  | Bema |  | RC |  | Exploration |  |  | 40 |  |  | 9,756 |
| 1997 |  | Bema |  | Core |  | Sulphide Exploration |  |  | 68 |  |  | 51,248 |
| 1997 |  | Bema |  | Core |  | Metallurgy |  |  | 3 |  |  | 453 |
| 1998 |  | Placer Dome |  | Core |  | Exploration, Infill |  |  | 15 |  |  | 12,311 |
| 1998 |  | Placer Dome |  | Core |  | Geotechnical |  |  | 3 |  |  | 2,253 |
| 1999 |  | Placer Dome |  | Core |  | Exploration, Infill |  |  | 8 |  |  | 6,608 |
| 1999 |  | Placer Dome |  | Core |  | Geotechnical |  |  | 4 |  |  | 2,752 |
| Total RC |  |  |  |  |  |  |  |  |  |  |  | 224 |  |  | 43,317 |
| Total Core |  |  |  |  |  |  |  |  |  |  |  | 120 |  |  | 79,430 |
| Total Drilling |  |  |  |  |  |  |  |  |  |  |  | 344 |  |  | 122,747 |

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Figure 11-1: Drill Collar Locations (PDTS, 2000)

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RC drilling was used principally to test the shallow oxide portion of the deposit on the north side of Cerro Casale and to pre-collar deeper core holes. RC holes have a range in depth from 23 m to 414 m and a mode of 100 m. The average RC hole depth is 193 m.
Core drilling was used to test mineralization generally at depths greater than 200 m.
Most RC and core holes were drilled from the southwest to northeast inclined at -60 to -70° to intersect the steeply south-dipping stockwork zones at the largest possible angle. Drill hole spacing varies with depth. Drill hole spacing is shallow oxide mineralization is approximately 45 m (Figure 11-2). Average drill-hole spacing in the core of the deposit in the interval between 3,700 and 4,000 m is about 75 m. Drill-hole spacing increases with depth as the number holes decrease and holes deviate apart. Average spacing at the base of the ultimate reserve pit is about 100 m.
Drilling equipment and methods are documented in several reports by Mineral Resources Development, Inc. (MRDI, 1997a, 1997b, 1997c) and PDTS (2000). In general, drilling equipment and procedures conform to industry standard practices and have produced information suitable to support resource estimates. Sample recovery, to the extent documented, was acceptable. Collar surveying was of suitable accuracy to ensure reliable location of drill holes relative to the mine grid and other drill holes. Downhole surveys of RC and core holes are not complete and locally downgrade the confidence in the position of individual intercepts of deep mineralization. Holes not surveyed are dominated by RC holes testing oxide mineralization less than 200 m deep.

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Figure 11-2: Average and Median Drill Spacing by Elevation (PDTS, 2000)

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11.1 | Drilling Methods

A variety of drilling contractors and drilling equipment have been used on the project since 1991. All equipment was suitable to the desired sample characteristics and hole depths.

11.1.1 | Reverse Circulation Drilling

RC drilling in 1991 was performed by Harris y Cía. using a Schramm 685 drilling rig with face-return hammer bits. This bit style ensures less sample loss and contamination between the more conventional bit and cross-over. Geotec Boyles Brothers did the RC drilling in the following two years using a CSR-1000 drill rig in 1992 and an Ingersoll Rand TH-75 drill rig in 1993. Face-return hammers were also used. Bachy-Franco Chileno drilled RC holes in 1994 using tricone bits. Bachy-Franco Chileno provided one drill with tricone bits in 1995. The rest of the RC drilling in 1995 was performed by Terra Services using two Longyear Drilltech D40K rigs and a combination of hammer and tricone bits. Drills used in 1995 and 1996 were equipped with 5 ¼" (13.3 cm) and 5 1/8" (13.0 cm) bits.
All drilling was done dry unless water injection became necessary to stabilize the hole.
A large number of the RC holes drilled in 1995 and 1996 were pre-collar intervals for deeper core holes. The RC portions of these holes were sampled and assayed where mineralized.

11.1.2 | Diamond Drilling Equipment

Core holes were drilled in 1993 to obtain samples for metallurgical tests of oxide gold mineralization. Geotec Boyles Brothers used a Joy 22 drill rig and NC (61 mm) core tools. Six holes totaling 464 m were drilled. The holes were not properly logged and assays were not obtained separate from the metallurgical results for composites, thus these holes were not used for geological interpretations and resource estimates.
Diamond drilling increased in 1995 with employment of three rigs by Geotech Boyles Brothers. Two Longyear 44 drill rigs and one Boytec Universal 650 drill rig were used. The Longyear 44 rigs used triple-tube HQ-3 (61 mm) and NQ-3 (45 mm) core barrels. The U-650 used a conventional double-tube HX (63 mm) core barrel.
Connors Drilling performed core drilling in 1996 and 1997 with two 40HH drill rigs and one 56A drill rig. The objective of drilling these two years was to test deep sulphide gold-copper mineralization. Holes were collared with HQ tools and reduced as necessary to NQ. This generally occurred at a depth of about 300 m. Holes precollared with RC equipment were set with HQ casing and then drilled to completion with NQ tools.

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Three holes totaling 463 m were drilled in 1997 for metallurgical tests. Assays were not obtained that could be used for resource estimates.
Placer Dome employed Connors Drilling again in 1998 and 1999 using the same drilling equipment. The same practices were observed as in 1997.
Half and one-third core retained after sampling for all holes is presently stored in permanent metal buildings at the project site and are on well organized and well maintained core racks. Cores from metallurgical holes were consumed and are not available for inspection.

11.2 | Geological Logging Practices

Logging of RC drill cuttings and core, followed procedures first introduced by Bema Gold and then modified somewhat by CMA and later by Placer Dome. The basic logging framework of lithologies, alteration, mineralization, and stockwork veining was retained in each campaign. Only parameters to represent intensity of attributes such as alteration and veining were modified. Ultimately, lithology and stockwork veining intensity were used as identification of ore controls for domaining in resource estimation; therefore, the quality of these interpretations is the principal issue material to resource estimates.
CMA used standard logging forms and entered information by hand on paper forms. These were transferred to database technicians in Copiapo where the information was transferred by hand to an electronic database. This practice was followed from 1991 to 1997. Placer Dome geologists used the electronic GEOLOG system and entered logged information directly into a database. The integrity of these entries was investigated by Placer Dome using “Geocheck” software, which examines the database for unique codes, mismatching hole depths in collar files and over lapping “from” and “to” intervals.

11.2.1 | Reverse Circulation Chip Logging

CMA geologists logged cuttings from each 2 m interval at the drill site using a hand lens. Color, silicification argillization, chloritization, limonite, jarosite, manganese oxides, pyrite, stockwork intensity, and magnetite were logged in 1991 through 1995. Potassium feldspar alteration, biotite alteration, chalcopyrite, specularite, copper oxides, and hematite were added in 1995 and 1996. Sericite, bornite, chalcocite, enargite/sulfosalts, dolomite, anhydrite, barite, kaolinite, and igneous textures were added to the logging in 1996 and 1997.
Geologists also logged rock type, grain size, oxide/sulphide ratio, and the estimated percentage of fines and clays in the sample before washing.

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Intensity of alteration and stockwork veining was estimated on a scale of 0-5 (lowest to highest) from 1991 to 1995. This was converted to a scale of 0 to 3 in 1995 (0=0, 1 & 2 = 1, 3 & 4 = 2 and 5 = 3). The intensity scale was 0 = none, 1 = weak, 2 = moderate and 3 = strong. Placer Dome further modified the stockwork intensity scale to signify the estimated volume percent of stockwork veins:

· | 0: 0 to 3%

· | 1: 3% to 7%

· | 2: 7% to 10%

· | 3: >10%

Understandably, the logging of the intensity of attributes is difficult with RC cuttings given that only the most resistant components are retained in a washed sample.
All RC drill cuttings were relogged with a binocular microscope by CMA in 1996 to improve the confidence in logging of oxide/sulphide ratio, oxidation state, rock type, stockwork intensity, and alteration type.

11.2.2 | Core Logging

Between 1993 and 1997, CMA first photographed core at a core shack on site, then logged the core for geotechnical parameters and geology. The scales used for attributes and intensity logged were the same as for RC cuttings.
Placer Dome logged 1998 and 1999 core at site using the electronic GEOLOG Logging System (GLS). Integrity of the data entered was checked by the Geocheck subroutine, which examines the data for improper codes and mismatched intervals. Placer Dome used the same geological codes as CMA. Major intervals of lithology, alteration, and stockwork intensity could not exceed 15 m (but could be repeated). Core was photographed both conventionally and digitally.
Placer Dome modified logging of stockwork intensity in 1998 by excluding gypsum veinlets in the estimation. This was done by selectively relogging core and RC cuttings from the central portion of the deposit and by incorporating results from detailed surface mapping. Veinlet stockwork intensity (minus gypsum veinlets) was combined with lithology to produce the final domains for resource estimation.

11.2.3 | Geotechnical Logging

Geotechnical logging before 1998 was done only on select holes. Vector Engineering logged lithology, core recovery, RQD, joint frequency, joint condition, degree of breakage, degree of weathering and alteration, and hardness for holes CCD007, CCD008, CCD009, CCD011, CCD012 and CCD013. CMA personnel logged RQD, core recovery and fracture frequency for CCD062 to CCD088.

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Placer Dome logged all 1998 and 1999 core for core recovery, degree of breakage, RQD, and magnetic susceptibility. Geotechnical holes GT-001 to GT-006 were also logged for degree of hardness, weathering, and alteration index, fracture conditions, joint conditions, number of fractures, and number of veins. Data were evaluated by Piteau Associates to provide guidance for pit designs.

11.3 | AMEC Review of Logging

AMEC inspected drill core for CCD096, CCD066, CCD067, and CCD068. All core for these holes were cut in half with a diamond core saw. Rock quality is high and few intervals of broken or ground-up core were observed. AMEC found the logging to be professional and representative of the lithology, alteration, and stockwork veining present.
AMEC also randomly inspected approximately 50 boxes of older core in a separate storage facility to for general condition and core recovery. Rock quality was found to be generally high with few intervals of strongly fractured rock and poor core recovery.

11.4 | Core and RC Recovery

Core recovery and RC sample weights are not discussed in the 2000 Feasibility Study by PDTS. Apparently, core recovery values and RC sample weights were not routinely digitized and added to the general drill hole database. Drilling contracts required in excess of 90% recovery for payment. AMEC randomly inspected drill logs and noted general high core recoveries (>95%) in mineralized intervals. Core randomly inspected in both Placer Dome and CMA core storage facilities at the project site showed high recoveries and infrequent intervals of broken core.
MRDI (1997a) reviewed RC sample weights for holes drilled through 1996 and found no relationship between copper grades and recovery. Similarly, gold showed no relationship to recovery in oxide intervals. The average grade of gold in sulphide mineralization, however, increases with recovery below 75%. The number of samples (654) of sulphide mineralization with less than 75% recovery is approximately 3% of the RC sample intervals; therefore, this bias does not materially affect resource estimates.

11.5 | Topography

The most current topography in use was developed by Placer Dome using satellite imagery (PDTS, 2000). AUTOCAD® drawing files were created with 2 m contour intervals in the area of the ultimate pit and at 10 m contours outside the design pit.

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Previous topography was produced by GenCen of Santiago, Chile using 1:8,000 aerial photographs flown in 1994. Topographic contours at 2 m intervals were produced for the pit area after matching contours to drill roads and trenches surveyed by Contreras Topografía Ltda. of Copiapo. A larger map was produced with 5 m contours to cover a 4 km 2 area around the pit area. Quoted vertical and horizontal accuracy is 2 m (MRDI, 1997a).

11.6 | Drill Hole Collar Surveys

Drill-hole collars are clearly marked with rebar or wooden posts cemented in the top of the hole, with metal drill hole identification tags (Figure 11-3). Markers for a moderate number of holes were destroyed by construction of additional drill roads on steep hillsides after the original holes were surveyed. Contreras Topografía Limitada surveyed each hole from 1993 to April 1996 using a theodolite. CMA acquired a Wild T2 theodolite and Wild D13000 laser distance meter in 1996 and surveyed the remaining hole collars. The survey reference datum is the 1956 Preliminary South American Ellipsoid (PSAD56) and the Canoa datum. Control was extended by third-order triangulation from a Chilean military post 15 km south of the project.
Figure 11-3: Drill-Hole Collar Monuments
CMA acquired an Ashtech SCA12, geodetic-grade, global positioning system (GPS) in 1993, and used this to survey drill holes and roads. All holes after CC221 and DD043 were surveyed with this GPS.

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Placer Dome surveyed holes drilled in 1998 and 1999 with a GPS. The Placer Dome report does not clarify if the GPS was a geodetic grade instrument or a less accurate GPS unit.
AMEC checked three drill sites on the surface relative to their plotted position on a detailed drill collar location map and found the positions in the field to be consistent with the map.
MRDI (1997b) checked all drill collar coordinates and elevations against their plotted position on topography and found no drill holes with discrepancies greater than the accuracy of the topographic survey.

11.7 | Downhole Surveys

Holes drilled in 1993 and 1994 were not originally surveyed downhole. In 1995 and 1996, CMA used a Tropari to measure downhole azimuths and dips on 50 m intervals. Few of the previous holes could be re-entered due to caved collars where casing had been removed. Tropari readings showed that some holes deviated significantly downhole from the original collar azimuth and dip setup. CMA hired a contractor to re-survey all accessible holes with a Sperry Sun multi-shot camera. The multi-shot surveys confirmed the deviations obtained by Tropari surveys.
The magnetite content of quartz stockwork vein zones can significantly affect readings of azimuth with a compass tool such as a Tropari or Sperry Sun multi-shot camera. For this reason, Tropari and Sperry Sun multi-shot azimuth readings that deviated significantly (approximately 10° or more) from the adjacent reading up hole were removed from the survey database.
In addition, a large number of Tropari azimuth readings were discarded because it was determined that there was an operator error in reading the instrument.
In 1996 CMA contracted Silver State Surveys of Elko, Nevada to survey all accessible holes using a north-seeking gyroscope. A small drill rig was used to attempt to open previous holes with depths greater than 200 m. Holes were re-surveyed with the gyroscope at 50 m intervals. Forty-six holes were surveyed with a gyroscope at this time.
Most of the 131 holes drilled by CMA in 1996 and 1997 were surveyed by Silver State Surveys or by Comprobe Surveys of Santiago with a north-seeking gyroscope. Approximately 6 holes were surveyed with a Sperry Sun single-shot camera by Connors Drilling.
Placer Dome contracted Comprobe to survey all holes drilled in 1998 and 1999 with a gyroscope.

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A total of 151 drill holes (44%) out of the entire list of 344 drill holes do not have downhole surveys. A majority of these are RC holes less than 200 m deep that were drilled in oxide mineralization. AMEC identified 14 unsurveyed holes (4% of holes) that are greater than 200 m deep. Six are greater than 300 m. Four (CCD009 at 380 m and CCD022 at 591 m, CCC173 at 318 m and CCC182 at 350 m) are in mineralization. The physical positions of intercepts of deep sulphide mineralization in these holes have a low confidence.
AMEC reviewed deviations incurred in holes 200 m deep and less and found that, with two exceptions, the drill holes deviated no more than 10 m from a straight-line projection. Beyond 200 m deviations increased significantly.
AMEC also inspected downhole survey results for anomalous azimuth changes that may have been caused by interference from magnetite in the mineralization. Only holes inclined at less than 80° were inspected because significant changes in azimuth can occur in near vertical holes without any material affect. Four inclined holes were found with changes in azimuth greater than 10° in short distances (10 m to 25 m), which suggest the presence of magnetite and potentially unreliable azimuth measurements. These are CCC098, CCD023, CCD032, and CCD043. Otherwise, downhole surveys appear reasonable and are suitable to support resource estimates.

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12.0 | SAMPLING METHOD AND APPROACH

Sample collection and handling of RC drill cuttings and core was done in accordance with industry standard practices, with procedures to limit sample losses and sampling biases. Drilling in 1991 to 1996 was primarily done with reverse-circulation equipment with hammer or tricone bits. Hammers used face-return bits to limit sample losses from a conventional cross-over. Tricone bits, by their basic design, are centre-return tools.
The majority of RC holes to 1995 are 250 m depth or less. RC holes drilled in 1996 and 1997 targeted deeper oxide mineralization and were as deep as 414 m. Core drilling was used exclusively to test deeper sulphide mineralization and for later infill of shallow mineralization. Core holes are from 30 to 1,473 m deep.

12.1 | Reverse-Circulation Drill Sampling

A variety of sample collection equipment and procedures were used. Drilling was done dry unless water injection for hole conditioning was necessary. From 1991 to 1995, a double cyclone system was used. A primary sample was obtained by running the discharge from the primary cyclone through a Gilson splitter. The discharge from the secondary cyclone was then added to the primary sample using the same Gilson splitter. One discharge hopper on the Gilson splitter was then split again until a final sample from 4 kg to 6 kg was obtained. This sample was placed in a numbered plastic bag and designated for either assay or for a metallurgical split. Metallurgical splits were stored in Copiapo.
RC drilling in 1996 and 1997 used a single cyclone and a Gilson splitter. Final sample weight was 4 kg to 6 kg.
Two-meter sample intervals were used in 1991 to 1994, which resulted in sample intervals crossing rod changes when Imperial 20 ft drill rods were used, or matching intervals when six m drill rods were used. After 1994, 5 ft sample intervals were used with 20 ft drill rods and 2 m intervals were used with 6 m drill rods.
CMA measured weight recovery based on the final sample weight and number of splits.
A rotary wet splitter was used when water injection was required because of perched water zones or hole conditions. The rotary splitter was adjusted to produce a 4 kg to 6 kg final sample, which was discharged into a porous, Olefin bag. According to MRDI (1997b), less than one percent of samples were collected wet. Weight recovery was not measured for wet samples.
All collection, splitting and bagging of samples was performed by CMA personnel.

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12.2 | Drill Core Sampling

Core drilled in 1993 (6 holes) was obtained for metallurgical sampling and was not assayed for resource estimation. Cores drilled in 1995 and early 1996 (11 holes) were placed in covered, wooden boxes at the drill rig by CMA personnel and moved to a covered, secure logging facility at the project camp. Core was logged and marked out into 2 m lengths for sampling. Select samples approximately 5 cm long were removed for density measurements.
Core obtained in 1995 and 1997 by Bema was cut in ⅔ and ⅓ portions with a diamond saw. The ⅔ portion was placed in double plastic bags with a stapled sample number ticket and then sent by truck to Bondar Clegg (now ALS Chemex) in Copiapo for preparation. Samples were delivered to Copiapo two to three times per week. Samples weighed from 12 kg to 14 kg. The ⅓ portion was retained in wood core boxes for reference. AMEC inspected these cores at the campsite and found them to be in good condition on organized core racks and with appropriate, permanent labeling.
These procedures were continued for the remainder of CMA core drilling in 1996 and 1997; with the exception that core was transported in open boxes to the camp logging and cutting facility. All work was done by CMA personnel. Procedures were in accordance with standard industry practices.
Placer Dome used similar procedures for core drilled in 1998 and 1999. Core was delivered to a core and storage facility at the project camp in covered, wooden boxes. The core was marked in 2 m intervals after being photographed and logged, and then cut in half with a diamond saw. One half was sent to Bondar Clegg in Copiapo for sample preparation and assaying. The other half was used as metallurgical samples or retained in the original core box. A majority of second splits of mineralized intervals in 1998 and 1999 core were sent as metallurgical samples and are not available for reference. Sampling by Placer Dome conforms to industry standard practices.
Core transport, sampling, and shipment of samples to Bondar Clegg were done by Placer Dome personnel.

12.3 | List of Significant Assays

Assays exceeding 0.3 g/t Au and used in resource estimates are provided in Appendix A.

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13.0 | SAMPLE PREPARATION, ANALYSES, AND SECURITY

13.1 | Sample Preparation

13.1.1 | Reverse-Circulation Samples

RC samples submitted to analytical facilities (after subsampling) were approximately 4 kg to 6 kg for all drilling campaigns.
RC samples collected in 1991 through 1994 were sent to Bondar Clegg Laboratories in Copiapo for preparation. Bondar Clegg dried each sample, and then crushed the entire sample in a Links mill to between minus 60 and minus 80 mesh. A 150 g split obtained from a riffle splitter was pulverized to 100% passing 150 mesh in a Tema mill.
Assaying of sample pulps were done by Monitor Geochemical Laboratory in Elko, Nevada.
In 1995, RC samples were shipped to Acme Laboratories in Santiago where the entire sample was dried and weighed prior to being crushed to minus 10 mesh. Specifications for the crushing quality are not documented. A 1 kg split was pulverized to minus 150 mesh in a ring-and-puck mill. Specifications for percent passing 150 mesh are not documented. Acme performed the assays in Santiago.
In 1996 and 1997, RC samples were delivered to either Bondar Clegg or SGS Laboratories in Copiapo for preparation. Bondar Clegg was the principal preparation laboratory and SGS handled overflow work, which comprised 39% of the samples. The entire samples were dried and weighed, then crushed in a Rhino jaw crusher to minus 10 mesh. The percent passing this specification is not known. One kilogram of material was pulverized to minus 140 mesh in a ring-and-puck mill. This product was blended and split into four 200 g samples. Three pulps were stored and one was sent to Acme in Santiago for assay.

13.1.2 | Core Samples

CMA and Placer Dome sampled core on nominal 2 m intervals, making a 12 kg to 14 kg sample for the CMA core (⅔ core) and a 9 kg to 12 kg sample for the Placer Dome core (half core).
Core samples from drilling in 1995 and 1996 were shipped to Bondar Clegg in Copiapo. The entire sample was weighed, dried and crushed to minus 10 mesh in a Rhino jaw crusher. The entire sample was then further crushed in 1 kg batches to minus 80 mesh in a 1.5 kg ring-and-puck pulverizer. These were homogenized and then a 250 g split was obtained with a riffle splitter. This split was pulverized to minus 150 mesh in a smaller ring-and-puck mill. Specifications for percent passing each mesh size are not documented.

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Standards and duplicates were prepared by Bondar Clegg personnel and were included in shipments of pulps to Acme Laboratories in Santiago.
In 1996 and 1997, core samples were prepared by Bondar Clegg or SGS in Copiapo. SGS handled overflow comprising about 20% of core samples. Samples were crushed to minus 10 mesh in a Rhino jaw crusher, blended and split to one kilogram. The split was pulverized to minus 140 mesh in a 1.5 kg capacity ring-and-puck mill. Four samples of 200 g each were split from the pulp. One pulp was sent to Acme Laboratories in Santiago for assay. The other three pulps were stored in Copiapo at CMA facilities.
Placer Dome core samples in 1998 were prepared at Bondar Clegg in Copiapo. The entire sample was weighed on an electronic scale and dried at 100°C to 120°C. The entire sample was then crushed to 100% passing 10 mesh in a Rhino jaw crusher. The entire sample was crushed in 1 kg lots to 100% passing 80 mesh in a LM-2 ring-and-puck pulverizer. The samples were homogenized and split to 260 g using a riffle splitter. The final split was pulverized to minus 160 mesh in a LM-2 ring-and-puck mill. Reject was stored. Pulps were sent to Acme Laboratories in Santiago for assay.
In 1999, Bondar Clegg prepared samples in Copiapo and sent pulps for assay at their facility in La Serena. Sample preparation consisted of drying the entire sample at 60°C, then crushing it to 75% passing 10 mesh in a Rhino jaw crusher. A one kg split was then obtained using a Jones riffle splitter. This was pulverized to 95% passing 150 mesh in a LM-2 ring-and-puck mill. Two pulps of approximately 250 g each were split from the pulp. One pulp was sent for assay; the other pulp was stored.
With the exception of core preparation in 1999, the methods for contamination control in sample preparation are not documented. In 1999, supposedly the preparation laboratory cleaned the jaw crusher and ring-and-puck pulverizer with compressed air between each sample and with quartz after every 10 samples. Sieve specifications were checked every 20 th sample. Assays of blanks for the 8 core holes drilled in 1999; however, show evidence of contamination.
Sample preparation protocols generally conform to industry standard practices although the final sample aliquot for RC samples in 1991 to 1994 (150 g) is very small for a gold deposit. A review of assay quality assurance and quality control by MRDI (1997a) shows that in this period the precision was worse than subsequent years when a larger sample pulp was prepared. This affected 86 shallow RC holes. The subsequent protocols of crushing of at least one kg to minus 150 mesh is more appropriate.

13.2 | Assaying

Monitor Geochemical Laboratory in Elko, Nevada performed assays of RC samples in the period of 1991 through 1994. Gold and silver were determined by fire assay with a one-assay ton (29.166 g) sample and gravimetric finish. Copper assays were completed on an unspecified sample weight (possibly 1 g) with atomic absorption spectrometry (AA) after an aqua regia digestion. Detection limits are not documented, although the gold and silver fire assay method should have a lower detection limit of at least 0.02 g/t Au.

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Acme Laboratories in Santiago performed assays in 1995 through 1998. Gold was determined on a one-assay ton sample by fire assay, with an AA finish. Samples exceeding 3 g/t Au were reassayed with a gravimetric finish. Gravimetric results were reported to CMA for samples re-assayed after initial AA analyses. Copper and silver were determined by AA after an aqua regia digestion of a 1 g sample. The lower detection limit for Au was 0.01 g/t.
Bondar Clegg La Serena did the assays in 1999. Gold was determined by fire assay of a one assay-ton sample, with an AA finish. Copper and silver was determined by AA after aqua regia digestion of 1 g of pulp. The lower detection limit for gold was 0.01 g/t.
Assay methods conform to industry standard practices for this type of deposit and for the metals of interest.

13.3 | Assay Quality Assurance and Quality Control (QA/QC)

13.3.1 | On-Site Procedures

Reverse-Circulation Holes
Duplicate samples and geochemical standards have been inserted into the sample series since the inception of CMA’s RC drill programs in 1993. The number of quality control samples and the procedures for submitting them have varied throughout the years. Approximately one in ten samples submitted to laboratories for holes CCC001 to CCC086 were control samples (one standard and one rig duplicate per run of twenty). From 1991 through 1994 (86 holes or 25% of drilling), Monitor Geochemical Laboratories inserted standards internally and CMA submitted RC rig duplicates for second analyses. From 1994 on, standards and duplicates were added to sample shipments at the sample preparation facilities in Copiapo and arrived blind to the analytical laboratory. Holes CCC087 to CCC224 contained one standard or blank and one duplicate per fifteen samples. Preparation and assaying were handled by the same laboratory for holes CCC087 to CCC184. Although Acme ultimately inserted the quality control samples into the sample stream, the laboratory was unaware of which of four standards or blanks was being utilized at any time. Duplicate samples were inserted at site, and therefore were blind to Acme. All standards, duplicates, and blanks were inserted by CMA personnel in Copiapo for holes CCC185 to CCC224, and were therefore blind to Acme. In all cases, the quality control samples were submitted either at random within a specific number of samples, or at specific intervals based on meterage.

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Core Holes
Core holes CCD001 to CCD006 were not assayed, but instead were evaluated as metallurgical samples. All subsequent drill core programs were subject to quality control procedures. Approximately one in ten samples was submitted for quality control for holes CCD007 to CCD017 (one standard and one duplicate per twenty samples). Two sample tags were attached to the sample intended for duplication as a guideline for the preparation facilities, and CMA provided the standard and blank. All quality control samples arrived at the analytical laboratory blind, as they were inserted into the sample stream by the preparation facility in Copiapo. Sample streams for holes CCD018 through CCD088 contained one standard and one duplicate per fifteen samples, and one sample in forty was a field blank. As before, duplicates were identified to the preparation facility by attaching two sample tags to a sample bag. CMA Personnel inserted the field blanks and standards into the sample stream. The blanks were inserted prior to preparation, whereas the standards were inserted after CMA received all prepared samples from the preparation facility. The location of the quality control samples within the sample series remained hidden from the analytical laboratory. In all cases, the quality control samples were submitted either at random within a specific number of samples, or at specific intervals based on meterage. Three quality control samples (one blank, one standard and one duplicate) were inserted on site by Placer Dome personnel in each batch of twenty samples for holes CCD089 to CCD103 and holes GT-001 and GT-002. The control samples were inserted on a random basis within the sample batch. Holes CCDI04 to CCD111 and GT-003 to GT-004 received two standards, two duplicates, and two blanks for each batch of forty samples. As before, the quality control samples were submitted on site in random order by Placer Dome personnel.

13.3.2 | Assay QA/QC - Pre-1995

QA/QC results for the first 86 RC holes were evaluated by MRDI (1994). Internal standards were used, but the recommended values for the standards were not well documented. Rig duplicate samples were collected and analyzed. Overall, the results of these duplicates indicated sampling, preparation, and analytical procedures were adequate for obtaining reproducible (±20%) results for gold and copper. No follow-up work was performed subsequent to that report. Coarse rejects and sample pulps are no longer available for drill holes from that time period (encompassing drill holes CC001 through CC086).
AMEC concurs with MRDI’s conclusions regarding pre-1995 QA/QC and agrees that assays for this period are generally suitable for use in resource estimates.

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13.3.3 | Assay QA/QC - 1995 and 1996

The QA/QC results for diamond holes CCD07 to CCD17 and reverse circulation holes CCC87 to CCC184 were reviewed by MRDI (1997a). This represents a total of 109 holes or 32% of the drilling.
Rig duplicate samples provide the most definitive picture of the overall reproducibility, or precision, of the assay database. These samples include all the sampling variation for the reverse circulation drilling, from the point of the initial sample split, through all the sample preparation stages, and the analysis. Consequently, comparison of the rig duplicates provides the best means of assuring that sampling has been representative and analytical procedures have been adequate. Precision for rig duplicates should be better than ±30% at the 90 th percentile.
Performance of rig duplicates is shown in Figure 13-1. Duplicate pairs with pair means less than 15 times the detection limit were excluded. Excluding very low values is necessary because the precision of measurement is much worse, in percentage terms, at concentrations at or near the analytical detection limit. The selections are such that there is an extremely low probability of excluding any “non-waste” samples.
Figure 13-1: Relative Differences for Rig Duplicates (MRDI, 1997b)
Note: X axis is percentile and Y axis is relative difference
Ninety percent of duplicates have a relative difference of less than ±25% for gold. These data indicate the sample size and preparation methods, combined with the analytical techniques employed by the assay laboratories, are sufficient for obtaining reproducible results within a given batch of samples.

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Results demonstrate that the gold and copper assays in 1995 and 1996 are sufficiently precise to be used in resource estimates.
Standards
CMA prepared standards and blanks and submitted them routinely in the sample stream with an insertion rate of 3.6% to 11.6%. Acme’s performance on inserted standards can be characterized as good; there is no significant drift over time.
Check Assays
Check assays for gold (pulps previously analyzed by Acme Lab were submitted to Chemex Laboratory in Vancouver, BC, Canada) were done on every tenth sample. The agreement between laboratories appears adequate for the needs of a feasibility study, with Acme returning a mean grade 5.3% higher than Chemex (Figure 13-2). Subsequent comparisons to standards revealed that Chemex was biased low relative to standards and therefore the Acme values are more acceptable. Precision for these data are shown in Figure 13-3.
Check assays for copper show an 11% high bias in the Acme results relative to those from Chemex (Figure 13-4). MRDI found in 1996 that Chemex was actually biased low in Cu relative to standards; therefore, the apparent high bias of Acme is not of concern.
Overall, gold and copper assays from the 1995 and 1996 drilling campaigns are suitable to support resource estimates.

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Figure 13-2: Checks of Acme Gold Assays by Chemex (MRDI, 1997b)
Note: X axis is g/t Au for Acme and Y axis is g/t Au for Chemex
Figure 13-3: Precision from Chemex Check Assays of Acme Gold Assays (MRDI, 1997b)
Note: X axis is percentile and Y axis is relative difference between analyses

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Figure 13-4: Chemex Check Assays of Acme Copper Assays (MRDI, 1997b)

13.3.4 | Assay QA/QC – 1996 and 1997

CMA retained Smee & Associates Consulting Ltd. in the fall of 1996 to perform an ongoing independent review and audit of QA/QC procedures (Smee, 1997). MRDI reviewed Smee’s conclusions and recommendations and accepted them (MRDI, 1997a). AMEC reviewed these reports and concurs with the conclusions.
Standards
CMA manufactured 18 geological standards over the life of the Cerro Casale drilling program. Standards were made by sorting -10 mesh reject drill material by grade, and compositing similar grade and mineralogical samples into bulk samples. Standards 1-6 were pulverized to 100% -150 mesh by SGS Labs, Santiago, then homogenized. Standards 7-18 were similarly prepared and homogenized by Bondar-Clegg of Coquimbo, Chile. Numerous splits of each standard were sent to a number of laboratories for round robin analysis. Results of this round robin analysis were used to calculate the accepted mean and standard deviation for each standard. The upper and lower acceptable limits were taken as ±2 standard deviations about the mean concentration for both copper and gold.
Standard results were plotted on time series charts, and out-of-range samples noted. In total, 2,088 submissions of gold standards and 2,065 submissions of copper standards were used with drill core samples of which 8 gold standards (0.38%) and 28 copper standards (1.4%) were out of limits. Batches with standards outside ±2 standard deviations were re-assayed. Two standards (9 and 10) were found to be inhomogeneous.

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Check Assays
Check analyses of Cerro Casale samples were done by Chemex Laboratories of Vancouver, Canada. A total of 3,033 diamond drill core samples were submitted for check analyses for gold and copper, 1,136 reverse circulation samples were analyzed for gold and 711 reverse circulation samples were submitted for check analyses for copper. Table 13-1 lists comparisons of Acme assays and Chemex assays.
Table 13-1: Check Assays by Chemex, 1996 and 1997 (MRDI, 1997b)

|  | DDH Duplicates |  | RC Duplicates |  | DDH Duplicates |  | RC Duplicates |
|  | Gold (g/t) Acme 0.519 |  | Gold (g/t) Chemex 0.509 |  | Gold (g/t) Acme 0.531 |  | Gold (g/t) Chemex 0.492 |  | Copper (%) Acme 0.202 |  | Copper (%) Chemex 0.212 |  | Copper (%) Acme 0.080 |  | Copper (%) Chemex 0.083 |
Difference |  |  | — |  |  | 1.870 | % |  | — |  |  | 7.252 | % |  | — |  |  | -4.722 | % |  | — |  |  | -4.222 | %
Number |  |  | — |  |  | 3033 |  |  | — |  |  | 1136 |  |  | — |  |  | 3033 |  |  | — |  |  | 711 |

On average, Acme analyses for gold are nearly 2% higher in diamond drill core, and 7% higher in reverse circulation samples than Chemex. However, Acme copper analyses are 4.7% lower in core and 4.2% lower in reverse circulation cuttings. These differences are within acceptable tolerances.
Analyses of standards by Acme and Chemex give some guidance in evaluation of the relative bias of each laboratory. Table 13-2 shows results for analyses of standards 8, 10, 11, 12, 14, 15, 16, and 18.
Table 13-2: Acme and Chemex Analyses of Standard, 1996-1997 (MRDI, 1997b)

Standard |  | Acme Average Gold (g/t) |  | Chemex Average Gold (g/t) |  | % Diff |  | Acme Average Copper (%) |  | Chemex Average Copper (%) |  | % Diff. |
8 |  |  | 1.41 |  |  | 1.35 |  |  | 4.39 |  |  | 0.046 |  |  | 0.048 |  |  | -4.45 |
10 |  |  | 0.80 |  |  | 0.74 |  |  | 8.02 |  |  | N/A |  |  | N/A |  |  | N/A |
11 |  |  | 1.32 |  |  | 1.23 |  |  | 6.73 |  |  | 0.787 |  |  | 0.800 |  |  | -1.63 |
12 |  |  | 0.63 |  |  | 0.59 |  |  | 6.28 |  |  | 0.066 |  |  | 0.072 |  |  | -8.45 |
14 |  |  | 0.63 |  |  | 0.58 |  |  | 6.63 |  |  | 0.391 |  |  | 0.406 |  |  | -3.89 |
15 |  |  | 1.27 |  |  | 1.20 |  |  | 5.34 |  |  | 0.453 |  |  | 0.473 |  |  | -4.43 |
16 |  |  | 0.53 |  |  | 0.50 |  |  | 4.64 |  |  | 0.148 |  |  | 0.152 |  |  | -2.53 |
18 |  |  | 0.79 |  |  | 0.74 |  |  | 5.44 |  |  | 0.376 |  |  | 0.390 |  |  | -3.67 |

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The Acme analyses of the gold standards range from 4.39% to 8.02% higher than the Chemex analysis. The Chemex analyses appear to be biased low compared to the Round Robin analysis in Standards 11, 12, 15, and 16. Although the differences in the gold analyses are small, the standard analyses suggest that Acme is closest to the most accepted gold concentration. Similarly, the copper standards show Chemex to be 1.6% to 8.45% higher than Acme, which is consistent with the results from the duplicate analysis. The Chemex standard analyses are higher than the established accepted mean for copper standards 8, 11, 15, and 18. The Acme analyses are therefore considered to be the more appropriate copper values.
Overall Precision for Field or Rig Duplicate Samples
Rig duplicates were obtained on average every 15 samples, or 6.7%. These duplicates should contain the sampling uncertainties introduced by splitting reverse circulation cuttings or core on site, splitting a fraction of crushed sample for pulverization at the preparation laboratory, and selecting a fraction for analysis from the pulp bag.
A total of 2,089 gold and 2,087 copper rig duplicate pairs were obtained from diamond drill core. The data were sorted by increasing mean of the duplicate pairs to facilitate a Thompson-Howarth precision calculation. The Thompson-Howarth bias plot for copper shows an excellent correlation between the two sets of analysis, with few exceptions. The overall precision of sampling and analysis for the Cerro Casale core drilling in 1996 and 1997 is excellent for both copper and gold. This is similar to what was found for the reverse circulation drill samples in other studies by Smee (1997) and MRDI (1997a).
Analysis of Blanks
Field blanks, consisting of coarse gravel-sized, non-mineralized crushed rock were inserted into the sample stream at the Cerro Casale site. These field blanks were blind to the assay laboratory, and were subjected to the entire sample preparation and analytical procedure. Out of 394 field blanks submitted, only five gold analyses (one%) exceeded 0.10 g/t and six copper analyses (1.5%) exceeded 0.03%. Three of the out-of-range blanks were actually a standard erroneously inserted into the sample stream in the position of the coarse blank. This low level of potential contamination is deemed acceptable.
Contamination in the analytical laboratory can occur during a gold fire assay procedure from previously used fusion crucibles, dirty glassware or reagents, or insufficient cleaning of the atomic absorption equipment between sample aspirations. This potential source of contamination was monitored by using a synthetic standard pulp (STD05). A total of 263 gold and 258 copper analyses are reported for STD05 as part of analysis of core. One pulp blank reported greater than 0.10 g/t Au, which was attributed to a data entry error, and only two were reported greater than 0.05 g/t Au. Only two copper blanks were initially reported as exceeding 0.03%, one of which was a data entry error. This low number of failed blanks shows that the sample preparation and analytical techniques were performed in a clean and professional manner.

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13.3.5 | Assay QA/QC – 1998

The quality control and assurance program (QA/QC) for Placer Dome’s 1998 assaying consisted of insertion of control samples into the sample stream prior to preparation and assay. Three types control samples were randomly inserted into every “batch” of 20 samples consisting of one standard, one blank, and one duplicate. This is a 15% control sample split. In addition to these control samples, approximately 10% of the samples with Au assays greater than 0.1 ppm were sent for check assay at Placer Dome’s Research Centre in Vancouver.
Results of the QA/QC program indicate that the assays for the 1998 drilling are of acceptable quality. AMEC understands that no assay jobs from Acme in 1998 had to be repeated.
Standard Samples
Four standards and two blanks were used in the 1998 QA/QC program. The standards used were the same as those employed during the previous drilling campaigns by CMA and are of mineralized material from Cerro Casale. The blanks are of two types. One is a prepared blank and the other is a field blank of unmineralized volcanic rock obtained from exposures south of the project area. Tables 13-3 and 13-4 show the best values and acceptance limits for the standards and blanks.
STD05 is the prepared blank sample and the results of gold analyses of that sample in 1998 are presented in Figure 13-5. With the exception of one sample, all of the results are less than 5 times the detection limit and are considered by AMEC to be within acceptable limits. The sample outside the limits indicates that the sample or batch of samples was contaminated or that the calibration of the instrument was significantly in error. Analyses for copper are presented in Figure 13-6. Two samples fall outside the pass-fail limits. Duplicate pulps should have been prepared and copper reassayed for those two batches.
Results of analyses for gold in STD12 are presented in Figure 13-7. Two samples are significantly below the acceptance limits and indicate a need to reassay the batches that contain those samples. There is also an obvious low bias to the data and an equally obvious downward drift to the data with time. The low bias averages about 3.9%, which is acceptable. Late in the program (batches 55 to 65), the bias is on the order of 6.5%, which is greater than is generally acceptable limits (±5%) and is cause for concern. Figure 13-8 presents the copper results. With the exception of a few samples in batch 65, all of the results are within limits and there is no obvious drift or bias to the data. AMEC suspects that the failing samples are mislabeled standard STD18.

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Table 13-3: 1998 Standards and Blanks Used at Cerro Casale – Gold

| Standard |  | Expected Au ppm |  | Min. Accept ppm |  | Max. Accept ppm |  | Number of Assays |
| STD05 |  |  | Blank |  |  | — |  |  | 0.05 |  |  | 220 |
| STD12 |  |  | 0.62 |  |  | 0.54 |  |  | 0.70 |  |  | 250 |
| STD13 |  |  | 1.51 |  |  | 1.33 |  |  | 1.69 |  |  | 110 |
| STD14 |  |  | 0.62 |  |  | 0.50 |  |  | 0.74 |  |  | 33 |
| STD18 |  |  | 0.74 |  |  | 0.58 |  |  | 0.90 |  |  | 50 |
| STD19 |  |  | Field Blank |  |  | — |  |  | 0.05 |  |  | 406 |

Table 13-4: 1998 Standards and Blanks Used at Cerro Casale – Copper

| Standard |  | Expected Cu % |  | Min. Accept % |  | Max. Accept % |  | Number of Assays |
| STD05 |  |  | Blank |  |  | — |  |  | 0.005 |  |  | 220 |
| STD12 |  |  | 0.066 |  |  | 0.054 |  |  | 0.079 |  |  | 250 |
| STD13 |  |  | 0.140 |  |  | 0.112 |  |  | 0.168 |  |  | 110 |
| STD14 |  |  | 0.400 |  |  | 0.250 |  |  | 0.550 |  |  | 33 |
| STD18 |  |  | 0.380 |  |  | 0.280 |  |  | 0.480 |  |  | 50 |
| STD19 |  |  | Field Blank |  |  | — |  |  | 0.005 |  |  | 406 |

Figure 13-5: 1998 Cerro Casale Standard (Blank) STD05 – Gold

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Figure 13-6: 1998 Cerro Casale Standard (Blank) STD05 – Copper
Figure 13-7: 1998 Cerro Casale Standard STD12 – Gold

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Figure 13-8: 1998 Cerro Casale Standard STD12 – Copper
Figure 13-9 presents gold results for STD13. Two samples are below the pass-fail limits and should have been re-assayed. Batches 0 through 41 exhibit a bias of 4%, which is within acceptable limits, but batches 50 to 64 exhibit a bias of 6.6% low, which is outside limits and is cause for concern. The obvious drift downward with time is also cause for concern. Figure 13-10 presents copper results for STD13. One sample is significantly above the pass-fail limit. The reason for that failure is not obvious and the batch containing that sample should have been reassayed. The results exhibit a very small high bias with no drift with time.
Results of gold analyses for STD14 are summarized in Figure 13-11. All of the gold results are within limits. There is a small, but detectable drift downward with time and an obvious low bias relative to the best value. The average bias is about 2.9% and the bias in batches 55 through 65 is about 3.7%, which is acceptable. Figure 13-12 summarizes the copper results for STD14. All of the samples are within limits and there is no obvious bias in the data. Batches 63 to 65 exhibit a somewhat larger than normal scatter (relative to earlier data) that is not a significant concern, but results such as this should be investigated carefully to see if it is indicative of a problem.

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Figure 13-9: 1998 Cerro Casale Standard STD13 – Gold
Figure 13-10: 1998 Cerro Casale Standard STD13 – Copper

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Figure 13-11: 1998 Cerro Casale Standard STD14 – Gold
Figure 13-12: 1998 Cerro Casale Standard STD14 – Copper

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Almost all of the gold results for SDT18 are within limits and there is a small, but insignificant drift downward with time (Figure 13-13). There is a small high bias relative to the best value. That bias is considered to be insignificant. A single sample in batch 54 is outside the limits and the batch should have been reassayed. Figure 13-14 shows the copper results for STD18. A sample in batch 35 is outside the limits and a sample in batch 6 is nearly out of limits. Batch 35 should have been re-assayed and batch 6 should have been considered for reassay. The pass-fail limits for STD18 appear to be very liberal for both gold and copper and should be re-evaluated.
STD19 is a blank sample collected from near the project area. Gold values (Figure 13-15) show three samples above the pass-fail limit, which is set at 5 times the detection limit for gold. The batches containing those samples (5, 41, and 48) should have been carefully evaluated for problems due to contamination. Copper in STD19 is problematical (Figure 13-16). A significant proportion of the values are above five times the 0.005% detection limit. Those results indicate that the sample either contains more than 0.005% Cu and is not blank or that there is a significant problem with contamination at the sample preparation laboratory. The average grade of the samples (minus a single outlier) is 0.01% Cu. It appears to AMEC that the sample contains approximately 0.01% Cu and should not be considered a copper blank. Scatter in the data also suggest that the detection limit reported by Acme is somewhat low and should be on the order of 0.025% Cu rather than 0.001% Cu.
Figure 13-13: 1998 Cerro Casale Standard STD18 – Gold

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Figure 13-14: 1998 Cerro Casale Standard STD18 – Copper
Figure 13-15: 1998 Cerro Casale Standard (Blank) STD19 – Gold

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Figure 13-16: 1998 Cerro Casale Standard (Blank) STD19 – Copper
Duplicate Samples
Duplicate sample pulps were prepared at Bondar Clegg in Copiapo and submitted to Acme in Santiago. Consecutive sample numbers were given to the original and the duplicate sample pulp. Data received by AMEC contains 416 duplicate samples that have consecutive sample numbers. These samples are useful for determining the analytical precision for the laboratory. Because there is no dependency between the two values in the duplicate pair, AMEC plots the pair maximum against the pair minimum to facilitate visualization of the data and use of the warning line. By doing this, all of the data plot above the X = Y line. The slope of the warning line for gold is 1.15 which approximates a precision level of +15% and the intercept is 0.3g/t, which is 30 times the detection limit (Figure 13-17). For copper, the slope is 1.1 and the intercept is 0.03% (Figure 13-19). Precision is estimated by plotting the relative error against the cumulative frequency of the relative error. This plot provides an estimate of precision which is inversely proportional to the relative error, that is, a relative error of 100% is poor precision, a relative error of 0% is extremely good precision. AMEC standardizes the precision estimate to the relative error at the 90 th percentile. AMEC expects a relative error at the 90 th percentile to be less than 15% for gold and less than 10% for copper.

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Figure 13-17 summarizes the gold duplicate data. The bulk of the data is beneath the warning line. The data above the warning line appears to be sample swaps in some cases, and random differences in other cases. Batches containing the samples above the warning line should have been investigated for possible reassay. Figure 13-18 is a plot of the relative error versus the cumulative frequency of the relative error. At the 90 th percentile, the relative error is about 19%, which is somewhat outside the expected 15%. This is, in part, a result of the samples that fall outside the pass-fail line and are possible bag swaps. The error may also be the result of less than optimum sample preparation.
Figure 13-19 is an X-Y plot of the copper data and shows that most of the data are beneath the warning line. The samples above the warning line should be investigated to determine if any of the batches containing those samples need to be re-assayed. Figure 13-20 is the cumulative frequency of the relative error. At the 90 th percentile, the relative error is about 6%, which is well within expected limits.
Figure 13-17: 1998 Cerro Casale Gold Duplicate Data

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Figure 13-18: 1998 Cerro Casale Gold Duplicate Data
Figure 13-19: 1998 Cerro Casale Copper Duplicate Data

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Figure 13-20: 1998 Cerro Casale Copper Duplicate Data
Check Assays
Samples were randomly selected from the sample database for check assaying at the Placer Dome Research Centre in Vancouver, BC, Canada. A random 10% selection of samples (471 samples) was taken from those samples with a gold assay greater than 0.10 g/t Au. Check assay results are summarized in Table 13-5. Figures 11.1.11 and 11.1.12 of Appendix II of the 2000 Feasibility Study (PDTS, 2000) graphically illustrate the data.
AMEC has not seen the raw data but concur with the Placer Dome conclusion, based on the data summaries, that there is little bias between the two laboratories for either gold or copper.
Table 13-5: 1998 Check Assay Statistics

| Sample (8=471) |  | Mean |  | St. Dev |  | Max |  | 75th Percentile |  | Median |  | 25th Percentile |  | Min. |
| Original Au g/t |  |  | 0.50 |  |  | 0.36 |  |  | 1.97 |  |  | 0.61 |  |  | 0.41 |  |  | 0.23 |  |  | 0.10 |
| Check Au g/t |  |  | 0.52 |  |  | 0.36 |  |  | 2.21 |  |  | 0.68 |  |  | 0.43 |  |  | 0.23 |  |  | 0.08 |
| Original Cu % |  |  | 0.202 |  |  | 0.151 |  |  | 1.250 |  |  | 0.290 |  |  | 0.116 |  |  | 0.090 |  |  | 0.004 |
| Check Cu % |  |  | 0.203 |  |  | 0.152 |  |  | 1.200 |  |  | 0.289 |  |  | 0.118 |  |  | 0.087 |  |  | 0.001 |

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13.3.6 | Assay QA/QC - 1999

The quality control and assurance program for the 1999 assaying consisted of insertion of control samples into the sample stream prior to preparation and assay. As with 1998, three types of control samples were randomly inserted into every “batch” of 20 samples, consisting of one standard, one blank, and one duplicate. In addition to these control samples, approximately 10% of the samples were sent for check assay.
Results of the QA/QC program indicate that the gold assays for the 1999 drilling could be showing a 3% to 10% high bias relative to the standards used and also compared to the Placer Dome Research Centre check assays. The 1999 copper assays are of acceptable quality. A total of 1,026 samples from 26 assay batches required repeat assaying. AMEC did not review the reassayed batches and is not aware of the results of the reassaying.
Standard Samples
The same standards and blanks were used in the 1999 QA/QC program as were used in 1998. The best values and pass-fail limits are presented in Tables 13-3 and 13-4.
STD05 is a prepared blank sample. Results for both gold and copper indicate that there is no contamination occurring during analyses of the samples. The graph for these results is not shown here.
With some exceptions, the gold results for STD12 are within the control lines (Figure 13-21). Three of the exceptions are mislabeled standards and one is unexplained, but which was probably a mislabeled sample. A number of samples fall between the upper warning line and the upper control line. This caused six batches to be reassayed. A small, but obvious high bias relative to the best value is evident and there is an obvious drift to the data with time (green line). The bias is about 3% and is not considered by AMEC to be a problem. The drift is somewhat excessive, but is not corroborated by similar drift in other standards. Copper shows the same four samples outside the control lines and another sample was above the upper control line (Figure 13-22). Two sample batches were reassayed as a result. Otherwise, all of the samples are within the warning lines. There is an obvious high bias relative to the best value that is not corroborated by all of the standards.
For STD13, three sample batches were sent for reassay as a result of one gold result outside the control limits and two gold results between the warning lines and control limits (Figure 13-23). All other samples were within the limits. The data exhibit a very small and probably insignificant high bias relative to the best value and an obvious downward drift with time. This is opposite to the drift observed in STD12. The drift is not considered to be a problem. Copper results for STD13 (Figure 13-24) are all within the control and warning lines and exhibit a small high bias relative to the best value.

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Figure 13-21: 1999 Cerro Casale Standard STD12 – Gold
Figure 13-22: 1999 Cerro Casale Standard STD12 – Copper

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Figure 13-23: 1999 Cerro Casale Standard STD13 – Gold
Figure 13-24: 1999 Cerro Casale Standard STD13 – Copper

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Standard STD14 was not used extensively during the course of the program (Figure 13-25). Gold results, with one exception, are within the warning line. The batch containing the one sample that is on the control line was reassayed. The data are biased high relative to the best value. The average bias is about 6.5%, which is outside generally acceptable limits (5% is generally accepted as the maximum bias between the laboratory and standard). The reason for this bias is not known, but STD12 and STD18 (below) exhibit similar, but on average less high bias. Copper results, with one exception were all well within the warning lines (Figure 13-26). The one sample that was outside the control line resulted in reassay of the batch containing the sample. There is no significant bias to the data.
STD18 exhibits a high bias of about 6% relative to the best value for gold early in the program (Figure 13-27). That bias drifts downward to nil later in the program. All of the samples were within the control lines. A single copper result is outside the control lines (Figure 13-28). The batch containing that sample was re-assayed. Other samples are within the warning lines and there is no discernable bias or drift to the data.
STD19 is a coarse blank collected near the Cerro Casale project that is periodically inserted to test for contamination from the sample preparation equipment. Gold results for this sample show somewhat normal behavior to about batch 135 (Figure 13-29). Results for batches 1 through 134 are more or less reasonable. Six samples are above 5 times the detection limit, which is considered to be a practical upper limit for blank samples. The reasons for those failures are not obvious. From batch 135 through batch 224 (approximately 3,300 samples), however, there are indications of routine and excessive contamination of samples being prepared at the preparation facility. Of the 227 blank samples prepared during that time, 90 fail the five times detection limit test and 42 samples exceed 0.1 g/t Au, containing up to 1.3 g/t Au. In contrast, of the 179 samples analyzed prior to batch 135, 9 exceed 0.05 g/t and three of those results are 0.06 g/t. The failing batches 135 through 224 are mostly samples from holes in prospects other than Cerro Casale, but do include assays for geotechnical holes 99GT003-006 and infill core hole CCD111 at Cerro Casale. Au grades above the 0.4 g/t internal cutoff are present in holes 99GT003, 99GT006 and CCD111. It remains to be determined if the coarse blank actually contained gold or if contamination occurred in sample preparation. The latter is the most likely reason, given the pattern of gold values. Intercepts in these three holes should not be used in resource estimates until the issue of contamination is resolved. Coarse rejects for these holes should be prepared and re-assayed for gold prior to the next resource estimate update. In the meantime, intervals from the subject holes should be considered to be biased high as much as 1.3 g/t.
Copper results for STD19 show an average grade of 0.10 % Cu, which is consistent with the 1998 results. This sample should not be used as a copper blank. Because the sample is coarse, it is subject to contamination during sample preparation, but it is not possible to determine at what level contamination begins, thus, this sample has little value as a monitor for copper contamination.

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Figure 13-25: 1999 Cerro Casale Standard STD14 – Gold
Figure 13-26: 1999 Cerro Casale Standard STD14 – Copper

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Figure 13-27: 1999 Cerro Casale Standard STD18 – Gold
Figure 13-28: 1999 Cerro Casale Standard STD18 – Copper

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Figure 13-29: 1999 Cerro Casale Standard (Blank) STD19 – Gold
Figure 13-30: 1999 Cerro Casale Standard (Blank) STD19 – Copper

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Duplicate Samples
Duplicate sample pulps were prepared at Bondar Clegg. Consecutive sample numbers were given to the original and the duplicate sample pulps.
Gold duplicate results are summarized in Figures 13-31 and 13-32. The X-Y plot shows three samples outside the warning line. The batches containing those samples should be considered for reassay. Figure 13-32 shows the cumulative frequency of the relative error. At the 90 th percentile, the relative error is about 19%, which is somewhat high for this type of project.
Figure 13-33 shows cumulative frequency of the relative error for the early data (pre batch 135) and the late data (batch 135 and higher). This plot used data 20 times the detection limit and above rather than the normal 30 times the detection limit in order to have enough data to investigate. The results clearly show that at the 90 th percentile, the relative error of the late data is much higher (40%) than the relative error of the early data (27%). This may be, in part due to the small number of data, but may also be due to sample contamination by the sample preparation equipment that is indicated by the results of STD19.
Figure 13-34 is the X-Y plot for copper duplicate samples. All but two samples are under the warning line. Batches containing those samples should have been investigated for possible reassay. The cumulative frequency of the relative error at the 90 th percentile is approximately 7%, which is within the normal range for this type of project.
Figure 13-31: 1999 Cerro Casale Gold Duplicate Data

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Figure 13-32: 1999 Cerro Casale Gold Precision Estimate
Figure 13-33: 1999 Cerro Casale Precision Estimate by Data Date

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Figure 13-34: 1999 Cerro Casale Duplicate Copper Data
Figure 13-35: 1999 Cerro Casale Copper Precision Estimate

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Check Assays
Samples were randomly selected from the sample pulps for check assaying at the Placer Dome Research Centre in Vancouver, BC, Canada. A random 10% selection of samples (359 samples) was taken from the assay database. Check assay results are shown in Table 13-6 and presented graphically in Figures 11.1.27 and 11.1.28 of Appendix II of the 2000 Placer Dome Feasibility Study (PDTS, 2000). The data suggests a 5% to 10% high bias for the Bondar Clegg gold assays in comparison to the Placer Dome Research Centre gold assays.
Copper check assays show good agreement with little bias.
AMEC has not reviewed these data, but based on the summary statistics, concurs with the Placer Dome assessment that Bondar-Clegg exhibits a high gold bias and little or no copper bias for the 1999 drilling program.
Table 13-6: 1999 Check Assay Statistics

| Sample (8=359) |  | Mean |  | Standard Deviation |  | Max |  | 75th Percentile |  | Median |  | 25th Percentile |  | Min. |
| Original Au (g/t) |  |  | 0.300 |  |  | 0.340 |  |  | 2.100 |  |  | 0.420 |  |  | 0.160 |  |  | 0.060 |  |  | 0.010 |
| Check Au (g/t) |  |  | 0.270 |  |  | 0.310 |  |  | 1.860 |  |  | 0.390 |  |  | 0.160 |  |  | 0.060 |  |  | 0.010 |
| Original Cu % |  |  | 0.097 |  |  | 0.115 |  |  | 0.973 |  |  | 0.126 |  |  | 0.068 |  |  | 0.015 |  |  | 0.002 |
| Check Cu % |  |  | 0.095 |  |  | 0.109 |  |  | 0.905 |  |  | 0.124 |  |  | 0.064 |  |  | 0.017 |  |  | 0.002 |

With the exception of contaminated batches 135 to 224 in 1999, all assaying is of suitable accuracy and precision to support resource estimates.

13.4 | Density

Measurements of bulk density were performed during the 1995 and 1996 core drilling campaign by E.C. Rowe and Associates (MRDI 1997a), by Kappes, Cassiday and Associates (KCA) during the 1996 and 1997 deep sulphide core drilling campaign, and by Placer Dome in 1998. A total of 877 density measurements were obtained from drill core of mineralized and waste units in these three drilling periods.
E.C. Rowe and Associates obtained bulk density measurements for 55 samples of oxide and sulphide mineralization using American Standard Testing Materials (ASTM) Method C97. This method involves weighing a dried sample of core, immersing it in water to fill pore spaces, and then reweighing the core in both air and water. This can overestimate bulk density when the rock is porous. MRDI (1997a) checked the method for 30 oxide samples by using a wax-coating, water immersion method (ASTM C914) performed by Rock Tech Laboratories in Salt Lake City, Utah, and found the initial measurements to be reliable.

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Another 117 core samples of deep sulphide mineralization were measured for bulk density by KCA in Reno, Nevada using a natural density method on non-sealed samples. Forty of these samples were checked by MRDI using the wax-coated, water immersion technique (ASTM C914-95), the results for which did not compare well with KCA’s measurements. The remaining 77 samples were measured with the ASTM C914-95 method and values obtained by KCA were not used. An additional 22 samples of mineralized granodiorite porphyry were measured by Lakefield Laboratories in Santiago using the ASTM C914 procedure. The 1995-1997 density data are summarized in Tables 13-7 and 13-8.
Table 13-7: Summary Statistics for Bulk Density Determinations, by Rock Type, All Sulphides

|  | Diorite Porphyry Sulphide |  | Microdiorite Breccia Sulphide |  | G. Diorite Porphyry Sulphide |  | Catalina Breccia Sulphide |  | Mafic Volcanics Sulphide |  | Pyroclastic Rocks Sulphide |  | Volcaniclastic Rocks Sulphide |
Mean (t/m³) |  |  | 2.63 |  |  | 2.66 |  |  | 2.61 |  |  | 2.64 |  |  | 2.87 |  |  | 2.59 |  |  | 2.72 |
Median (t/m³) |  |  | 2.64 |  |  | 2.67 |  |  | 2.61 |  |  | 2.61 |  |  | 2.87 |  |  | 2.61 |  |  | 2.72 |
Mode (t/m³) |  |  | 2.67 |  |  | 2.65 |  |  | 2.67 |  |  | NA |  |  | 2.84 |  |  | NA |  |  | NA |
Standard Deviation |  |  | 0.064 |  |  | 0.068 |  |  | 0.069 |  |  | 0.189 |  |  | 0.045 |  |  | 0.285 |  |  | 0.044 |
Minimum (t/m³) |  |  | 2.48 |  |  | 2.44 |  |  | 2.49 |  |  | 2.39 |  |  | 2.81 |  |  | 2.22 |  |  | 2.68 |
Maximum (t/m³) |  |  | 2.74 |  |  | 2.81 |  |  | 2.74 |  |  | 2.99 |  |  | 2.95 |  |  | 2.91 |  |  | 2.77 |
Number |  |  | 57 |  |  | 41 |  |  | 22 |  |  | 7 |  |  | 10 |  |  | 4 |  |  | 4 |

Table 13-8: Summary Statistics for Bulk Density Determinations, by Oxidation State, All Rock Types

|  |  | Oxide |  | Sulphide |  | Mixed |
| Mean (t/m³) |  |  | 2.42 |  |  | 2.65 |  |  | 2.44 |
| Median (t/m³) |  |  | 2.44 |  |  | 2.66 |  |  | 2.42 |
| Mode (t/m³) |  |  | 2.33 |  |  | 2.67 |  |  | 2.40 |
| Standard Deviation |  |  | 0.123 |  |  | 0.105 |  |  | 0.109 |
| Minimum (t/m³) |  |  | 2.02 |  |  | 2.22 |  |  | 2.30 |
| Maximum (t/m³) |  |  | 2.65 |  |  | 2.99 |  |  | 2.63 |
| Count |  |  | 52 |  |  | 145 |  |  | 6 |

Placer Dome selected 673 core samples from 1998 holes for bulk density measurements. A 10 cm sample of un-split core was taken at 20 m intervals downhole in drill holes 98CCD090 to 98GT02a. Dried core was weighed in air on a balance, and then weighed in water. The difference in weight between the two measurements represents the water volume of the sample. The dry weight divided by the volume is the density. Samples were considered to be non-porous so they were not coated with wax. This was generally confirmed by MRDI tests of E.C. Rowe and Associates measurements in 1997.

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Good agreement was found between earlier density measurements and those obtained in 1998. Placer Dome performed a statistical evaluation of the bulk density by lithology, alteration, stockwork intensity, and degree of oxidation. Of these parameters, degree of oxidation appears to be the main control to bulk density followed by lithology (Figure 13-34). Densities increase with depth; however, this is essentially measuring the change of the degree of oxidation. Density values used for tonnage calculations are presented in Table 13-9.
Table 13-9: Specific Gravity for Mineralization Domains

| Rock and Mineralization Type
| C01 (Oxide and Mixed) | C02+C03+C04+C05 (Sulphide) | C06 (Catalina Breccia) | C15 (Undefined)
| 2.40 | 2.65 | 2.58 | 2.61

Figure 13-36: Boxplot of All Density Measurements by Oxidation Categories

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This approach to categorizing density assignments is appropriate because it incorporates differences between key rock types (intrusives, breccias and non-intrusives; oxidation state) and differences between non-mineralized and mineralized rock (stockwork intensity).
Density measurement methods are suitable to support mineral resource and mineral reserve estimates and were performed with protocols conforming to industry standard practices. AMEC agrees with the assignment of densities by oxidation domain.

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14.0 | DATA VERIFICATION

14.1 | Database Development and Integrity Checks

Geological, geotechnical and analytical information were developed over a period of multiple exploration programs between 1991 and 1999, involving Bema Gold, CMA, MRDI, and Placer Dome staff. Entry of information into databases utilized a variety of techniques and procedures to check the integrity of the data entered. During the 1991 to 1993 period, geological data were entered into spreadsheets in a single pass by CMA personnel in Copiapo. The 1994 geological information were entered twice and corrected by MRDI in San Mateo, California. CMA staff in Copiapo used dual entry of data in 1995 to 1997. Placer Dome converted all databases to GEOLOG® format and then entered all geological logs directly into this system without a paper log step.
With the exception of one period of drilling, assays were received electronically from the laboratories and imported electronically into drill hole database spreadsheets.
Historical databases include detailed geological and geotechnical logging, assays and density measurements. The entire database includes 23 fields for geological attributes and 5 fields for assays (gold, silver, total copper ppm, total copper percent and sample weight). MRDI (1997a, 1997b) audited all geological and assay databases for CMA drilling from 1991 to 1997. Placer Dome data for drilling in 1998 and 1999 have not been previously audited.
For this technical report, AMEC was supplied a database including assays (hole ID, from, to, Au assay, Cu assay, lithology code, oxidation code, stockwork intensity code and sample number), drill hole collars (hole ID, grid coordinate, total depth and elevation) and drill hole surveys (hole ID, depth, azimuth, dip).

14.1.1 | Data for 1991 to Early 1996 Drilling Campaigns

As part of the 1996 oxide exploration program, data entered into Quattro Pro® spreadsheets for the 1991-1993, 1994, 1995 and a portion of the 1996 drilling were converted by CMA to dBASE® files. Changes in logged attributes were also incorporated. Dual entries of geological logs for 1994, 1995, and 1996 were compared by MRDI, and mismatched entries were corrected using original logs.
Assays performed by Monitor Geochemical Laboratories in 1991 to 1993 were downloaded from Monitor’s electronic bulletin board and imported directly into Quattro Pro® spreadsheets and then the database. In 1994, assays were entered from faxed certificates twice, once at CMA in Copiapo and again at MRDI in San Mateo. These were converted to dBASE® files, compared and corrected. Assays for 1995 and 1996 from Acme in Santiago were downloaded from a bulletin board and imported directly into spreadsheets and then the database.

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MRDI (1997a) audited 5% of entries for geological attributes and assays against original logs and certificates for the 1991 to early 1996 drilling campaigns and found an error rate of 0.2%. This is considered to be an acceptable error rate for data used to support resource estimates.

14.1.2 | Data for Late 1996 through 1997 Drilling Campaign

Geological logs were entered into Quattro Pro® spreadsheets by CMA personnel twice, then converted to dBASE® files. The files were compared and discrepancies fixed by comparing the information to original logs. Assays were imported directly into spreadsheets and then the dBASE® database as text files from Acme Santiago’s electronic bulletin board.
Data from all periods up to the completion of the oxide-sulphide prefeasibility study in late 1997 were combined by MRDI in San Mateo and audited. MRDI (1997b) checked 5% of the data added in 1996 and 1997 and found an error rate of 0.294%. Data are suitable to support resource estimates.

14.1.3 | Data for 1998 and 1999 Drilling by Placer Dome

The geological database by Placer Dome contains 16 separate fields covering rock type, rock code, texture, oxidation state, stockwork characteristics, and mineralogy. Assays include gold, copper, and silver. Available documentation suggests that this information was entered directly into GEOLOG® at the core logging facilities, then imported into an Access® database. Assays were downloaded as text files from Acme Santiago’s bulletin board and imported directly into Access.

14.2 | AMEC Data Verification

14.2.1 | Database

AMEC checked geological entries for seven pre-1998 RC holes and six pre-1988 core holes against GEOLOG® outputs to confirm that transformation of the data from the original formats was error free. In addition, all geological codes for one 1998 (98CCD089) and one 1999 (99CCD110) Placer Dome drill hole were checked against original GEOLOG® prints. No errors were found in a total of 3,393 entries.
Assays for CMA drilling in 1991 to 1997 were audited in detail by MRDI (1997a, 1997b and 1997c). Low error rates were verified. For the 2005 technical report, AMEC checked all gold and copper assays for holes 98CCD089 and 99CCD110 and found no errors for these 1,558 entries (4.5% of total 1998-1999 database).

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AMEC checked downhole survey records for gyroscope surveys of 1998 and 1999 holes and found database entries to agree with these documents. Survey files for pre-1998 holes were not available for review.

14.2.2 | Geological Interpretations

AMEC was provided original cross sections and plans used to develop outlines of rock types, alteration, stockwork intensity, and oxidation state for the deposit. These included:

| · | oxidation state, 1997, sections 250 to 1,200, 50 m intervals

| · | oxidation state, 1998, sections 250 to 1,200, 50 m intervals

| · | lithology, 1998, plans on 30 m intervals

| · | stockwork intensity, sections 250 to 1,200, 50 m intervals

| · | stockwork intensity, plans on 30 m intervals

| · | lithology, 1997, cross sections 250 to 1,200, 50 m intervals

| · | stockwork intensity with gold composites, cross sections 250 to 1,200, 50 m intervals

| · | resource blocks, 1998, measured + indicated resources.

AMEC inspected sections and plans of outlines of geological attributes to determine if the interpretations obeyed attributes posted on drill hole traces and if the interpretations were reasonable. In general, interpretations were reasonable with smoothed outlines that ignored minor anomalies in contacts. The result was interpretations that could be used for resource estimation without creating artifacts of interpolation along irregular contacts. The contacts between oxide, mixed oxide-sulphide and sulphide material follow topography and structures as expected. Interpretations of contacts between intrusive, breccia and volcanic units are reasonable relative to the model of a diorite porphyry laccolith, high-angle granodiorite intrusive and high-angle breccias. Stockwork intensity is subjective, given the variability of the logging of intensity of this feature. The relationship between gold and copper grades and the highest stockwork intensity is evident.

14.2.3 | Sampling and Assaying

AMEC did not independently sample drill core and obtain commercial assays of check samples. This was not considered to be necessary given the extent of historical blind QA/QC undertaken by CMA and Placer Dome (see Section 13.3 of this report) and the level of independent auditing of sampling and assaying by MRDI in 1994 through 1997.

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15.0 | ADJACENT PROPERTIES

There are no properties immediately outside the Aldebarán area claims that are pertinent to the Cerro Casale project.

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16.0 | MINERAL PROCESSING AND METALLURGICAL TESTING

The mineral reserves of the Cerro Casale project consist of copper and gold mineralization in roughly equal, but low-grade, economic quantities. The geologically complex, billion tonne resource includes at least nine distinct and recognized rock types. Average metal grades range from 0.12% to 0.40% copper and from 0.4 to 0.8 g/t gold. A small zone of high grade mineralization that contains 1.12% copper and 4.1 g/t gold is also reported. The ores are relatively hard with metric work indices of up to 22 (for comparison, an ore of average hardness will have a metric work index of about 14). Comparatively fine grinds are necessary to achieve adequate liberation of the economic minerals in sulphide mineralization.
The principal copper mineral is chalcopyrite with minor amounts of bornite reported in some areas. Gold tends to report with chalcopyrite in predominantly sulphide ores. A concentrate of 25% copper that contains up to 55 g/t gold and 110 g/t silver has been produced in bench-scale flotation tests.
There is a comprehensive, feasibility level, database, which supports technical and economic evaluations completed to date. The principal documents are a comprehensive feasibility study prepared for the project by PDTS in January 2000, along with updates to capital and operating costs for this document prepared by PDTS in 2004 and 2005. The 2000 Feasibility Study report includes summaries of the relevant test bases including sample analyses and various bench-scale and pilot test programs.
In its 2000 Feasibility Study, PDTS interpreted the test information and prepared a design basis for the project including design criteria, conceptual process flowsheets, major equipment selections, and conceptual level plan and elevation drawings. An equipment-factored capital cost estimate was developed that reflected Placer’s extensive experience in developing both gold and copper operations. An operating cost estimate, using PDTS internal cost data, was also prepared.
Subsequently, the original feasibility study results were reviewed and updated by PDTS, first in March 2004 and again in mid-2005. Bema has also contributed to this work, in particular, in commissioning several crushing and grinding studies. This work, also including reviews of flotation test work and gold leaching investigations, is presented in a report by MQes dated May 2006.
The metallurgical database and the project history were reviewed previously in detail by AMEC as part of a National Instrument 43-101 Technical Report (AMEC March 2005). This update to that technical report considers the reasonableness of salient issues regarding the initial test work and the feasibility study. This work also considers issues where conditions may have changed since then or where significant new assumptions or strategies were made.

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16.1 | Scope of Facilities

The design concepts and operating philosophies applied to the development of the Cerro Casale mine, metallurgical plant, and infrastructure are consistent with current industry standard practices. Technical emphasis is placed on the utilization of two lines of large, state-of-the-art, mineral processing process equipment. The operations will also be highly automated with comprehensive “design for maintenance” provisions.
The original, 2000 PDTS study basis included a nominal 150,000 t/d mining and processing operation. Plant facilities included a concentrator, flotation tailings leaching plant, concentrate pipeline, and infrastructure. Concentrator operations included three lines of SAG mills, “tank style” flotation cells, vertical stirred ball mills for regrinding, conventional thickeners for concentrates, and high-rate thickeners for mill tailings.
A trade-off study by MQes evaluated the feasibility of reducing the number of grinding lines from three to two. The study was based on simulations performed by Contract Support Services (CSS) using the JKSimMet evaluation software, evaluation techniques, and database. The study also included reviews of the recovery implications of increasing the final grind size from 120 microns (based on the grind optimization studies completed for the PDTS 2000 report) to 150 microns (a more typical grind size for large porphyry copper mineral treatment operations).
MQes concluded that a two-line plant is technically feasible if the hard and soft ores are blended to reduce the overall average plant feed hardness. Increasing the target grind size also substantially reduces the necessary installed grinding power. Sample calculations by AMEC indicate that slightly lower mineral recovery at a moderately coarser grind is offset by slightly reduced grinding costs. Therefore, the two-line grinding plant can also be economically justified. Possible risks are continued escalations of power costs and consumables, and of reduced operating flexibility.
Metallurgical operations will be conducted with the following plant facilities and unit operations:

| · | Primary Crushing and Coarse Ore Stockpile

| · | Two-Line SAG, Ball Mill, Crusher Circuit

| · | Flotation, Regrinding, and Concentrate Cleaning and Upgrading Circuits

| · | Concentrate Handling Unit Processes (Thickening, Slurry Transport, Filter Plant)

| · | Cleaner Tailing Leaching and Carbon-in-Pulp (CIP) Gold Recovery Circuit

| · | Gold Refinery

| · | Cyanide Destruction Circuit

| · | Dump Leaching Pad and Gold Recovery Plant

The design capacities for the operations include heap leaching of oxide ores at 75,000 t/d and milling (SAG grinding and flotation) of 150,000 t/d of mixed and sulphide ores. Tailings (at approximately 12,000 t/d) from the cleaner flotation operation are treated by cyanidation for gold recovery.

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A sulphide concentrate is produced and transported via slurry pipeline to a coastal site for shipping to offshore smelting and refining facilities. The gold contained in concentrate sold is projected to average 840,000 oz/yr. Copper contained in the concentrate sold will average 133,000 t/yr over the life-of-mine.
Doré bars are produced on site from heap leaching of oxide ores and CIP treatment of flotation cleaner tails. Up to an additional 439,000 ounces of gold will be produced from the dump leach operation in the early years of the project. Cleaner tails leaching will produce and additional 75,000 oz/yr of gold.

16.2 | Design Criteria

The PDTS study completed in 2000 included a four page summary of general project and process design criteria. These initial design criteria were used to guide the development of process flowsheets, equipment selections, and plant arrangements that provide the basis for estimating capital and operating costs. These criteria provide the base case for the Cerro Casale Project. Updated plant design criteria now include the following:

| · | Run-of-Mine Crushing Rate, Operating - 8,900 t/h

| · | Coarse Ore Stockpile Capacity, Live Storage - 135,000 t

| · | Grinding Rate, Calendar - 150,000 t/d

| · | Grinding Rate Operating - 162,500 t/d

| · | Range of Ball Mill Grinding Work Index, Bond - 16.5 kWh/t to 18.3 kWh/t

| · | Primary Grind, P80 - 150 μm

| · | Rougher Flotation Retention Time - 30 min

| · | Cleaner Tailing Leach Retention Time - 24 h

| · | Valley Fill Run-of-Mine Dump Leaching - 75,000 t/d

The complete list of preliminary design criteria also catalogue rock and mineral densities, general flow rates and slurry densities, mineral recoveries, settling and filtration rates, overall operating schedules, and other relevant design information. However, this information has not been presented in a single document of project criteria that describes all relevant process data. AMEC believes this document should be maintained and updated as additional design criteria are obtained..
The level of detail presently available provides a sufficient basis for defining project concepts and evaluating the metallurgical and economic performance.

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16.3 | Metallurgical Recoveries

As part of the original scoping work developed and managed by PDTS, a mapping program was performed to determine the variation in response amongst the various rock types. The program was used as a guide in preparing composites for the bench scale program. Regression analyses were developed to establish the relationship between head grade and tailing assay for each composite. The relationships were generally confirmed by pilot tests and provide a useful database for continued monitoring of development programs.
Mineral recovery is also influenced by the primary grind size. Finer grinds tend to yield higher recoveries. From a series of grind-recovery tests a pr imary grind size of 120 microns (μm) was selected as the optimum grind size. As grinding investigations proceeded, it was concluded that a coarser grind was possible and that the revenue-cost relationships would not materially change.
Subsequent confirmatory flotation test work by PDTS in 2005 returned lower than expected recoveries, especially of copper. Bema and MQes believe the samples used had “aged” to a point where they were no longer representative of the original core material. As a result, Bema Gold and Placer reviewed the results of the 2005 test work and the trends in grade and recovery from all previous work and calculated recovery targets for the project as 74.7% for gold and 86.4% for copper. The evolution of these targets is summarized in Table 16.1.
Table 16-1: Metal Recoveries into Concentrate

| Study | Au Rec, % | Cu Rec, %
| Original PDTS 2000, 120 μ grind | 75.03 | 87.29
| New PDTS 2005, 150 μ grind | 75.13 | 82.87
| Calculated Recovery, Adjusted to 150 μ | 74.70 | 86.40

Source: MQes Project Update, May 2006
The various regression formulas are consistent with copper recoveries throughout the test programs. Gold recoveries tend to be slightly overstated. Also, the mapping tests trended toward coarser grinds. The grind sizes used for the initial feasibility evaluations tend to mitigate this slight bias. The “Calculated Recoveries” shown above are considered appropriate for feasibility level economic reviews. Although AMEC considers these adjustments to expected recoveries to be reasonable, confirmation through follow-up test work is recommended.

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16.4 | Supporting Data and Test Work

16.4.1 | Ore Classifications and Rock Types

The rock types identified in the Cerro Casale mineral deposit are principally hard, mostly un-oxidized granites, diorites, and fine-grained volcanics. The predominant economic minerals identified in the various rock types are chalcopyrite and bornite. Gold is typically carried in the copper minerals in the sulphide zones and as free gold in the oxidized zones.
Rock types are further classified into three categories, according to the degree of weathering. Material exhibiting a loss of less than 10% of its original sulphide content is classified as a sulphide (DSL, DSU, GS, and VS). At the other extreme, material with less than 10% of its original sulphide content remaining is classified as an oxide (AO). All other material is considered mixed (MDBX, CBX).
The principal rock types, proportions, and associated grades are presented in Table 16-2.
Table 16-2: Mineral Reserves by Metallurgical Rock Type (MQes, 2006)

|  |  | Tonnage |  | Average Grades
| Rock Type |  | (Mt) |  | (%) |  | (% Cu) |  | (g/t Au)
| Upper Diorite Sulphide (DSU) |  | 112.8 |  | 10.9 |  | 0.26 |  | 0.58
| Lower Diorite Sulphide (DSL) |  | 71.6 |  | 6.9 |  | 0.31 |  | 0.74
| Granodiorite and Biotite Porphyry Sulphides (GS) |  | 201.0 |  | 19.4 |  | 0.31 |  | 0.78
| Volcanic Sulphide (VS – VB, MVF, and VPF) |  | 357.0 |  | 34.5 |  | 0.26 |  | 0.66
| Microdiorite and Hydrothermal Breccia (MDBX) |  | 105.3 |  | 10.2 |  | 0.34 |  | 0.81
| Oxide and Mixed – Cu > 0.10% (AO) |  | 41.0 |  | 4.0 |  | 0.24 |  | 0.71
| Catalina Breccia (CBX) |  | 3.2 |  | 0.3 |  | 1.15 |  | 4.11
| Oxide – Cu <= 0.10% |  | 117.8 |  | 11.4 |  | 0.04 |  | 0.54
| Undefined/Others (UD) |  | 25.7 |  | 2.5 |  | 0.24 |  | 0.42
| Totals |  | 1,035.4 |  | 100.0 |  | 0.26 |  | 0.69

Note: Errors in totals are due to rounding.
Five rock types comprise 86.4% of the overall reserves. Characterization of the metallurgical behavior of these five ore types, with respect to establishing the plant design criteria is generally adequate.

16.4.2 | Mineralogy

Mineralization is associated with quartz vein stockworks containing sulphides and magnetite, as well as a potassic-feldspar alteration. Scans from x-ray diffraction indicated that the most common minerals are, in decreasing order: quartz, feldspar, mica, chlorite, gypsum, pyrite, chalcopyrite, and bornite.

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Copper and gold are strongly correlated. Gold content has a tendency to follow copper content, as long as copper is present in stockwork-controlled chalcopyrite or bornite. The correlation does not hold for disseminated copper occurrences. Silver is reported to be present at ratios of 2:1 to 3:1 with respect to gold. Modal analysis demonstrated that (except for GS) up to 85% of the gold content is associated specifically with chalcopyrite. Less than 1% occurs with pyrite.
Copper is found mainly in chalcopyrite but bornite and, to a lesser extent, chalcocite, digenite, covellite, chrysocolla and malachite occur occasionally. The bornite-to-chalcopyrite ratio increases with depth. The average copper grade also increases by up to 25% at depth. Chalcocite, covellite, chrysocolla, and malachite are found at the oxide/sulphide boundary.

16.4.3 | Comminution

Two phases of comminution parameter investigations were completed. Altogether, the tests included impact, SAG, rod, and ball mill grindability, and abrasion index measurements. The first test work phase was completed for Bema Gold Corp. at McClelland Laboratories in Reno, Nevada, and Hazen Research (Hazen) in Golden, Colorado in 1997. The second phase was commissioned by Placer Dome in 1998 and performed at Hazen. Results from the drop weight tests conducted by Hazen were then interpreted by CSS for simulation work utilizing the JKSimMet software in 1999. Follow-up simulations by CSS investigated conditions using pre-crushing, larger pebble ports, open-circuit SAG milling, and roll crushing.

16.4.4 | Selection of Optimum Grind Size

A series of batch rougher tests were conducted at different primary grind sizes with composites of four of the principal sulphide rock types. An economic evaluation model was developed from the resulting metallurgical responses, with relative revenue levels and associated operating cost estimates calculated to compare the different primary grind target scenarios. The trade-off analyses also considered the impact of lower mining cut-off grades.
The optimum grind size is influenced by the metal price scenarios assumed. Higher metal prices often warrant the incremental recoveries gained from finer grinds and improved mineral liberation. Higher costs, typically from escalation in power costs, tend to indicate a coarser optimum grind size at the cost of reduced recovery.

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Increasing grind size from 120 microns to 150 microns reduces copper and gold recoveries by about one percent. At up to $450 per ounce for gold and $1.50 per pound for copper, the saving in grinding cost exceeds the potentially “lost” revenue. All grinding investigations by Placer and Bema have been guided by these price-cost trade-offs.

16.4.5 | Flotation

The flotation response of the Cerro Casale composites was defined by a bench-scale program followed by pilot plant testing at the Placer Dome Research Centre (PDRC). Confirmation bench scale tests were conducted at G&T Metallurgical (G&T) in Kamloops, British Columbia.
An initial phase of bench-scale investigations consisted of 94 mapping tests on assay coarse rejects taken from 19 drill holes. Specific samples, 88 in all, representing single rock types and/or alteration patterns, were tested individually in order to establish the variability of the flotation response and to decide how many discrete rock types should be defined for further flowsheet development.
The results for these samples, regrouped into four different rock types, were used to establish by regression analysis the expected ore response versus feed grade equations later used for the economic model. This work remains as an important source of reference information for assessing the metallurgical behavior under variable feed grade conditions.
Samples from the initial bench-scale mapping tests were recombined into seven new composites representing DSL, DSU, GS, MDBX, VS, VB and AO rock types to optimize flotation procedures and primary grind size. Emphasis was placed on tests involving the DSL and DSU materials which represent nearly half of the orebody.
A pilot plant campaign was then completed at the PDRC facilities using six rock type composites (DSU, DSL, GS, MDBX, VB+VS, AO). Various combinations were tested using ratios intended to simulate expected mine output over the life of the project.
Confirmation bench-scale tests were completed in parallel by G&T on five composites (DSL, DSU, GS, MDBX, VB), at similar grinds as those used by PDRC. A series of batch tests, in open and locked cycle, were performed by G&T just before the 2000 Feasibility Study Report was completed.
These programs established proper reagent doses, pH levels and the tolerance of the process to additions of oxidized ores. The conditions for use of collectors and depressants in the cleaner circuits were also clarified.
The early bench scale work was mostly exploratory in nature but later metallurgical objectives became more focused on defining the optimized process conditions for the feasibility report. The final definition of the expected metallurgical response including flotation and leach kinetics and reagent consumptions is based upon the latter confirmatory test work.

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AMEC believes the flotation test work and process database is appropriate for project definition and feasibility level investigation.

16.4.6 | Concentrate Quality

Samples of concentrate from two flotation pilot plant runs were sent to SGS Canada Inc. for chemical analysis (Table 16-3), with average composite head grades shown for reference.
Table 16-3: Indicative Concentrate Quality Analyses

| Item | Units | DSU #16-17 | DSL #25
| Gold | oz/t | 1.011 | 0.91
| Silver | oz/t | 5.76 | 6.44
| Copper | % | 27.16 | 20.28
| Arsenic | % | 0.25 | 0.21
| Mercury | ppm | 30 | 6
| Lead + Zinc | % | 1.84 | 1.78
| Composite Head Grade |  |  |
| Copper | % | 0.22 | 0.4
| Gold | g/t | 0.54 | 1.1

Source: PDTS, Feasibility Study 2000
The concentrations of common penalty elements are included in Table 16-3. Only arsenic and mercury have potential penalty issues. AMEC considers these to be minor and not of material concern.

16.4.7 | Cyanidation of Cleaner Flotation Tails

Cyanidation of tailings from the flotation circuit was investigated in parallel with the flotation test work to recover additional gold. Recovery of gold from the cleaner tails represents a potential recovery increase of 6 to 7%, which equates to an average of 72,000 ounces of gold annually.
Cleaner tailing leach work was completed by G&T in December 1999. For the cyanidation/CIL trials, both cyanide concentrations of 250 ppm and 500 ppm reached gold extraction nearing 91% after 24 hours, with a third of the copper content dissolved as well. Both concentrations required higher consumptions of cyanide and lime than the bottle roll tests (1.5 kg/t of cyanide and 3.5 kg/t of lime), likely due to increased demand from leached copper.

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A reprise of the 2000 tailings leach test program was performed by Placer in late 2004, and the results indicated potentially high cyanide consumption due to elevated levels of cyanide-soluble copper in the cleaner tailings. As noted previously, Bema believes the samples used for this test program are no longer representative of the original composites. Oxidation of copper sulphides, particularly chalcopyrite, may have increased the cyanide-soluble copper content to unacceptable levels. The present MQes report assumes that the metallurgical results of the original PDTS work remain representative of leaching performance. AMEC considers this to be reasonable - subject, as noted, to follow-up testing and analysis.

16.4.8 | Thickening

Pocock Industrial Inc. (Pocock) conducted standard and high-rate thickening tests on products of the 1999 pilot plant trials, but these tests targeted a finer grind than ultimately selected. Further settling test work was conducted with scavenger tailings samples of the various rock types tested during the last phase of flotation work (G&T, January 2000 report).
This work indicated that conventional thickeners should be employed for concentrates and that high-rate thickeners may be used for tailings. The present solid/liquid data base is considered adequate for feasibility level process evaluations.

16.4.9 | Filtration and Transportable Moisture Limits

Leaf and pressure filtration tests were conducted by Pocock on samples of concentrate produced during the pilot plant trials at PDRC. Vacuum filtration tests could not dewater concentrates to less that 18.8%. The equivalent cake moistures achieved by pressure filtration were 11% and 12%. AMEC considers this level to be high. Additional trials will be needed when new concentrate samples become available.

16.4.10 | Slurry Rheology

Viscosity measurements were reportedly made by Pocock on thickened concentrate and tailings products obtained from the pilot plant trials realized by PDRC. The results were not available for this review.

16.4.11 | Cyanide Destruction and Water Treatment

The introduction of a leach circuit on the first cleaner tailings requires a cyanide destruction circuit. The circuit design may need to deal with elevated amounts of dissolved copper as well as cyanide. Cyanide destruction will be performed using the INCO/SO 2 destruction process. This process, in fact, requires some soluble copper to catalyze the reaction of cyanide to cyanate.

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The tailings pond site is likely to show a negative water balance, with seasonal excess inflows accumulated and evaporated over the rest of the year. If any tailings pond water is to be discharged, the current assumption is that no further treatment of the effluent will be required, neither for removal of heavy metals nor for pH adjustment.

16.5 | Supporting Data and Test Work

16.5.1 | Comminution

A substantial base of information has been compiled regarding the grindability of various rock types and composites. Work began with the initial bench testing programs by PDTS in 1996 and 1997. The test work included the various Bond methods, MacPherson Autogenous tests, and the JKSimMet Drop Tests. Similar work has been commissioned by Bema Gold on composites prepared by Bema.
The 2000 PDTS feasibility study included data from a variety of recognized grindability tests. For comparison, the grinding characteristics of the four ore types representing the majority of the Cerro Casale ores are shown against two notable large and recently constructed Australian operations (Table 16-4).
Table 16-4: Comparative Grindability Parameters

|  | Cerro Casale | Cadia | Fimiston
|  | DS | MDBX | GS | VS | Monzonite | Volcanics |
| A | 65 | 65 | 65 | 65 | 65 | 65 | 42 - 50
| B | 0.51 | 0.42 | 0.56 | 0.43 | 0.58 | 0.494 | 0.61 - 0.65
| Ta | 0.38 | 0.38 | 0.47 | 0.4 | 0.494 | 0.21 | 0.26 - 0.40
| A*B | 33 | 27 | 36 | 28 | 38 | 32 | 26 - 33
| Ai | 0.28 | 0.43 | 0.35 | 0.33 | 0.26 - 0.34 | 0.26 - 0.34 | 0.13 - 0.40
| Impact Wi | 10.4 | 9.2 | 14.6 | 10.5 |  |  |
| Bond Rod Wi | 19.3 | 22.1 | 18.8 | 19.3 | 16 - 19.1 | 23.8 | 18.9 - 21.1
| Bond Ball Wi | 16.9 | 18.3 | 16.5 | 16.7 | 11.9 - 21.4 | 22.3 | 13.0 - 15.9
| Auto Wi | 16.5 | 18.5 | 18.1 | 17.0 |  |  |

Source: Bechtel Data Review, 12 February, 04 and PDTS 2000 Report
Most all of the development work to date is based on SAG technology as being the most appropriate means for treating the hard Cerro Casale ores. Large grinding units are now featured in most modern high-tonnage copper concentrators. State-of-the-art unit power draw capabilities of up to 26,000 kW for SAG mills and 14,000 kW for ball mills are now available. The technology of mill drives is equally advanced and well established.

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The major process challenges are in matching and balancing the number of mill lines, the power required for large grinding mills, and the physical requirements for grinding hard and relatively fine-grained rock types.
Cerro Casale ores are unusually and uniformly hard in comparison to other porphyry-hosted orebodies. In one sense, the data is consistent with geological observations of limited penetration of oxidation into the depths of the deposit. However, weathering (oxidation) can proceed rapidly as rock faces are exposed during mining operations reducing overall resistance to breakage. It is reasonable to expect, therefore, that the grindability information available represents a “worst case” condition.
In May 2005, a series of simulation studies were completed using JKSimMet process modeling software to evaluate earlier grinding test work and simulations. The first objective was to identify the conditions where throughput and grinds can be achieved with two grinding lines and comparatively larger equipment units. A second objective was to estimate the potential capital cost savings.
The studies suggested that a two-line plant with available power of about 112,000 kW is technically possible and can process 150,000 tonnes per day. A range of Bond Ball Mill Work indices of 16.5 to 18.3, corresponding to the values shown in Table 16.4, is used in this analysis. It is also considered that a grind of 150 microns will not materially affect metal recoveries. For confirmation of projected grinding capacities, AMEC reviewed the mill sizing concepts with simpler, but broader, calculation routines (Table 16-5).
Table 16-5: Power Draw Estimates for Cerro Casale Comminution Circuits

|  |  | Three Grinding Lines
|  |  | Units, ea |  | Diam, ft |  | Length, ft |  | kW per Unit |  | Total kW
| Primary Crushers* |  | 2 |  |  |  |  |  | 746 |  | 1,492
| SAG Mills |  | 3 |  | 40.0 |  | 22.0 |  | 19,600 |  | 58,800
| Pebble Crushers** |  | 3 |  |  |  |  |  | 746 |  | 2,238
| Ball Mills |  | 6 |  | 24.0 |  | 34.0 |  | 10,500 |  | 63,000
|  |  | Total |  | 123,530
|  |  | Two Grinding Lines
|  |  | Units, ea |  | Diam, ft |  | Length, ft |  | kW per Unit |  | Total kW
| Primary Crushers* |  | 2 |  |  |  |  |  | 746 |  | 1,492
| SAG Mills |  | 2 |  | 42.0 |  | 25.5 |  | 25,600 |  | 51,200
| Pebble Crushers** |  | 3 |  |  |  |  |  | 746 |  | 2,238
| Ball Mills |  | 4 |  | 26.0 |  | 38.0 |  | 14,000 |  | 56,000
|  |  | Total |  | 110,930

Notes: * 60 x 110 Gyratory Crushers. **MP-1000 Cone Crushers.
A two-line plant supplies about 10% less power. This reflects the limits imposed by the largest commercially demonstrated mill sizes that are presently available. This places a two-line plant closer to available power limits and introduces a lower operating flexibility. It is possible that the mine will deliver harder ore over several day periods, temporarily increasing power loads and reducing throughput However, this is not an unusual situation in modern, high-tonnage copper mining operations. Table 16-6 illustrates the nature of these limits.

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Table 16-6: Estimated Power Required vs. Ore Hardness (Work Index)

| Work Index, Wi | Target Grind, μ | Required Power, kW | % of Capacity
| 3-Lines | 2-Lines
| 16.5 | 120 | 108,900 | 87 | 98
|  | 150 | 97,100 | 77 | 88
| 18.3 | 120 | 120,800 | 96 | 109
|  | 150 | 107,700 | 86 | 97

MQes and PDTS updated the optimized grind calculations from the January 2000 feasibility study and considered the general trends to account for increased power, new grind-recovery relationships, consumable costs, and present commodity prices. Grind size, recovery, and mass pull balances as reported in the 2000 study were used to establish new material balances in flotation. The results were used, in turn, to support a conclusion that a grind size target of 150 microns is appropriate for the present economic conditions.
MQes also evaluated the economics of utilizing High Pressure Grinding Rolls (HPGR) in place of the SAG mills to produce ball mill feed. Roll crusher technology was widely used in the early 20th century in the “tonnage” mills of the day. It was recognized that crushing is more power efficient than grinding but the wear surfaces on the rolls were considered to be more expensive and difficult to maintain. The modern high pressure rolls are said to have overcome these challenges.
Two large roll crusher operations are currently being constructed and early results will be of interest for potential benefit to the Cerro Casale Project. While the power savings may be significant, multiple crusher and screening lines will increase capex and operating complexity. AMEC cannot otherwise comment on these developments at this time.
MQes concluded the study work by updating and adjusting the capital cost estimates to include the larger mills with selected factors adjusted in consideration of a two-line plant.

16.5.2 | Flotation

The flotation responses of Cerro Casale composites were defined in a progressive series of bench and pilot-scale test programs at the PDRC. Bench-scale programs proceeded through mapping composite tests, rougher/scavenger flotation tests, kinetic tests, regrind size optimization, three-stage cleaning in open and closed circuit, and recovery/head grade/concentrate grade tests.

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The generalized process flowsheet for these tests is shown in Figure 16-1. This flow scheme is also representative of the intended scaled-up commercial process.
Figure 16-1: Generalized Process Flowsheet
The programs have proceeded through optimization definition tests and compositing for pilot testing. Important design criteria developed from the work included appropriate and effective reagents and dosage rates, pH levels, flotation times, concentrate production rates and concentration ratios, plus characterization of expected concentrate quality.
A typical mass balance for this flowsheet is shown as Figure 16-7. These figures provide the design basis for scale-up from lab and pilot work to a commercial operation.
Stream #5 represents the cleaner flotation tailing that is treated by cyanidation. An estimate of this mass flow would be mill feed (150,000 t/d) times the weight percent of cleaner tailing (6.7%), which equals 10,500 t/d calculated or 12,000 t/d design.

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Table 16-7: Average Metallurgical Performance for Combined MDBX, VBX, and GDS Ore

| Stream | 1 | 2 | 3 | 4 | 5 | 6 | 7 | 8 | 9
| Weight, % | 100 | 92.3 | 7.7 | 8.9 | 6.7 | 2.2 | 1.2 | 1.0 | 99
| Assays
| Copper, % | 0.31 | 0.02 | 3.7 | 3.5 | 0.12 | 13.5 | 1.6 | 26.7 | 0.03
| Iron, % | 4.3 | 3.9 | 8.2 | 8.4 | 5.2 | 18.3 | 10.2 | 27.2 | 4.0
| Gold, g/t | 0.83 | 0.21 | 8.2 | 7.9 | 0.8 | 29.2 | 5.8 | 55.1 | 0.25
| Distribution
| Copper, % | 100 | 7 | 93 | 99 | 3 | 96 | 6 | 90 | 10
| Iron, % | 100 | 85 | 15 | 15 | 8 | 9 | 3 | 7 | 93
| Gold, % | 100 | 24 | 76 | 76 | 6 | 78 | 8 | 93 | 30
| From PDTS Feasibility Report, January 2000, Volume 3 - Metallurgy, Appendix 3.2

The mass flow rates are used to calculate the required volumes and number of flotation cells required for commercial operation (refer to Table 16.8 next page). Sample calculations indicate that the selection process is appropriate and that installed flotation capacity is adequate.
Table 0-8: Flotation Cell Selection Criteria

|  | Lab & Pilot Results, min | Scale-up Factor | Commercial Residence Time, min | Installed Cell Volume, m3 | Configuration
| Roughers | 15 | 2 | 30 | 9600 | 6-lines of 10-160 m 3 cells
| 1st Cleaners | 7 | 1.7 | 12 | 840 | 3-lines of 7-40 m 3 cells
| 2nd Cleaners | 4 | 1.5 | 6 | 120 | 3 lines of 4-10 m 3 cells
| 3rd Cleaners | 2 | 2.5 | 5 | 90 | 3 lines of 3-10 m 3 cells

Source: AMEC Technical report 2005
The flotation tests also defined the appropriate reagents and physical and chemical conditions for efficient mineral recovery and concentration. These consumption rates are used in the original PDTS 2000 study, and through the updates by Placer, Bechtel and MQes. Typically, the laboratory dosage rates are higher than commercial experience. A benchmarking review by Bechtel indicates the collector dosages projected for Cerro Casale are substantially higher that other comparable operations.

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16.5.3 | Regrinding

Six vertical, stirred, ball mills are included in the original PDTS study. The Bechtel review indicated that the installed power and number of units is high compared to other typical copper concentrators. Conversely, AMEC is concerned that there may be insufficient regrind power and capacity.
A screen analysis of the rougher concentrate indicated a P80 of 100 microns. The target grind is 30 microns. The measured (and presumably required) grinds were on the order of 25 microns. This is relatively fine compared to other copper operations.
AMEC recommends performing additional regrinding tests, including jar tests, vendor tests, and simulations. The immediate affect on project economics is not material.

16.5.4 | Cyanidation of Cleaner Tailings

The recovery of gold by cyanidation and the Carbon-in-Pulp (CIP) processes are well-understood technologies. The finely-ground cleaner tails exhibit rapid leaching and extraction kinetics with recoveries approaching 90 percent in less than 24 hours. The potential complication of soluble copper has been noted and various mitigation strategies are being considered.
The process will include thickening of cleaner tails to 40% solids and leaching for 24 hours in five stirred tanks in series. The dissolved gold and silver is then adsorbed by activated carbon in a CIP circuit of six stirred tanks in series. Some copper is also adsorbed.
Carbon is retained in each tank with discharge screens and is periodically pumped forward from tank to tank. CIP tail slurry is passed over safety screens prior to delivery to the cyanide destruction process. Loaded carbon is delivered to the elution plant approximately once per day.
In the elution plant, the loaded carbon is stripped of its adsorbed metal content. The carbon is first rinsed and then washed with hydrochloric acid. Copper is removed from the carbon by a cold water rinse. Gold and silver are then eluted from the carbon with a hot, and high strength, cyanide solution. Barren carbon is treated in a reactivation furnace, screened to remove degraded fines, and returned to the leaching process.
Gold and silver are recovered from the eluate by electrowinning. The electrowon metal is melted and poured into Doré bars.
The design criteria prepared by PDTS in the original 2000 feasibility study are reasonable and appropriate for the intended processes. The equipment selected for the process is capable of treating up to 12,000 tonnes per day of cleaner tailings and producing some 72,000 ounces of gold.

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While considered to be conventional and conservative, the design criteria for the tailings leach plant are based on a limited amount of test work. Since both copper and gold (and, most likely, small amounts of mercury) are present, additional test work is recommended to assure that the process will efficiently and economically handle the expected metal contents.

16.5.5 | Solid-Liquid Separation

Thickener selections are based on test work by Pocock, which defined achievable settling rates, flocculant types and dosages, overflow clarities, and final underflow densities.
The tailings thickeners are high-capacity units (i.e. deeper tanks, re-circulating feed-wells, and relatively small diameters). Test work indicated a sizing factor range of 3.6 to 4.2 square meters of tank surface area per cubic meter per hour of slurry feed. PDTS chose the low range for thickener selection, which resulted in the sizing of a 91 m diameter high rate thickener. Five high-rate thickeners are required.
Mineral concentrates typically do not respond as well to the application principles of high-rate thickening. A thickener sizing factor of 0.4 m 2/t/d was applied according to reported test results. This falls within a normal range of settling rates (0.21 to 0.68 m 2/m 3 per hr) for other concentrate applications. A single, 30-meter diameter conventional thickener is needed for concentrate. A single 80-meter diameter conventional thickener is appropriately sized for cleaner tailings leach feed.
Pressure filtration was selected to dewater the concentrates delivered by slurry pipeline to the Punta Padrones site near Caldera. The final moisture of 12% is higher than expected but it is likely that application parameters can be adjusted to a more typical range of 8 to 9% moisture.
AMEC considers these selection criteria to be reasonable at this conceptual stage. Using a “worst case” selection factor will have no material effect on plant economics. Additional testing as the project is developed is recommended.

16.5.6 | Concentrate Pumping and Slurry Pipeline

An average of 1,300 metric tonnes per day of copper concentrate will be produced. The concentrate will be thickened to approximately 55% solids for pumping from the mine site to the port. The technology has been successfully applied to transporting copper concentrates at two large copper operations in Chile under very similar conditions.
PSI Consulting has been involved in both of the Chilean projects. The details of the PSI report are not available. AMEC believes the capital cost estimate for the pipeline is a capacity-factored estimate and will need to be confirmed by a more detailed engineering study.

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16.5.7 | Heap Leaching

One conclusion from the 2006 study by MQes is that oxide ores should be treated separately from the mixed and sulphide ores. It was also concluded that the costs of crushing the oxide ores could not be economically justified. Therefore, a dump leaching scenario was developed to treat, uncrushed, run-of-mine oxide ore. The potential problems associated with cyanide soluble copper would be mitigated by excluding ores to the heap that have cyanide-soluble copper values of 0.1% or more.
The proposed process will leach run-of-mine ore in a valley-fill dump. The low-grade leach solutions will be collected at the toe of the dump and contacted with activated carbon in a cascade of columns. Gold and silver will be recovered from loaded carbon in the same manner as for the gold circuit in the cleaner tailings leaching operations.
Trade-off studies and examination of the area available for the leach dumps indicate that a mining and stacking rate of 75,000 tonnes per day is appropriate. Earlier column tests indicated possible heap leach recoveries in excess of 70% in one year’s time. The tests also indicated a less-than-usual sensitivity of recovery to particle size, even at relatively coarse sizes (12 mm plus). The study presumed that the column test results might be extrapolated to a run-of-mine situation by downgrading overall recovery to 65%.
Capital costs for the heap leach facility are estimated to be $90 million or approximately 5% of project capital. Assuming an average grade of 0.5 g/t and an average recovery of 55%, an additional 250,000 ounces of gold can be produced by an oxide heap leach program. An additional benefit may be that the dump material that would otherwise have been considered waste and, therefore only an operating expense, now becomes a revenue source.
The underlying assumptions for the program are considered to be reasonable but difficult to confirm. The logistics of organizing and conducting a small dump leach test are somewhat daunting. It may be necessary to make a project decision base upon a more carefully considered trade-off study. AMEC considers that this study should be commissioned before a final decision is made regarding dump leaching of oxides.

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17.0 | MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1 | Mineral Resource Estimates

The mineral resource estimates in the 2000 Feasibility Study for the Cerro Casale project were calculated under the direction of Marc Jutras, P. Eng. of Placer Dome. The estimates were prepared in 1999, and were made from 3-dimensional block models utilizing Placer Dome’s in-house mine planning software. Project limits are 4,000 to 4,405 East, 23,900 to 27,905 North, and 2,506 m to 5,005 m elevation. Projects limits are in truncated UTM coordinates with 470,000 subtracted from easting coordinates and 6,900,000 subtracted from northing coordinates. Cell size was 15 m east x 15 m north x 17 m high.
June 2006 mineral resource and mineral reserve estimates are developed from the same block model; however, mine plans were developed using a model which MQes re-blocked into 15 m high blocks. Global and local variances resulting from re-blocking are of negligible consequence with respect to mineral resource and mineral reserve estimates. AMEC recommends revising the block model to use 15 m high blocks (the selected mining bench height) for future engineering and economic evaluations.

17.1.1 | Geologic al Models and Data Analysis

Various geological aspects were modeled in order to assess their extent in controlling gold and copper mineralization at Cerro Casale. Geologic models were created for lithology, structure, oxidation, stockwork intensity, K-feldspar alteration, and silicification. AMEC’s reviews of these models are discussed in previous sections. Based on field observations and initial review of the completed geologic models, PDTS concluded that the Cerro Casale gold model would be best represented by a combined lithologic-stockwork intensity model, whereas the copper model should be a combination of lithology-oxidation level-stockwork intensity parameters. The combined models, along with their percent of the total project model volume, are shown in Table 17-1. AMEC concurs with this philosophy for development of geologic models or domains for use in grade interpolation at Cerro Casale.
These mineralized domains were reviewed through exploratory data analysis to determine appropriate estimation or grade interpolation parameters. The data analysis involved X-Y scatterplots, generation of histograms and cumulative frequency or probability plots, boxplot diagrams and contact plots. The data analysis was done on composited assay data. Assays were composited into 2 m down-hole composites. A composite length of 2 m was chosen because most of the assay lengths were taken at 2 m intervals. While AMEC agrees with the philosophy of this composite length choice, AMEC also believes that a larger composite length (5 m for example) may have been more appropriate considering the model block size and style of mineralization. Impact on the global estimate for this model, however, would likely be minimal.

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Table 17-1: Gold and Copper Geologic Models or Domains

| Model Code | Rock Code | Description | Volume % of Total Model
| Gold |  |  |
| G01 | DGB_0 | Intrusives; Stockwork Intensity = none | 0.65
| G02 | DGB_1 | Intrusives; Stockwork Intensity = low | 0.85
| G03 | DGB_2 | Intrusives; Stockwork Intensity = medium | 0.93
| G04 | DGB_3 | Intrusives; Stockwork Intensity = high | 0.32
| G05 | MDHB_0 | Breccias; Stockwork Intensity = none | 0.02
| G06 | MDHB_1 | Breccias; Stockwork Intensity = low | 0.09
| G07 | MDHB_2 | Breccias; Stockwork Intensity = medium | 0.06
| G08 | MDHB_3 | Breccias; Stockwork Intensity = high | 0.007
| G09 | CBX | Catalina Breccia | 0.003
| G10 | VMR_0 | Volcanics; Stockwork Intensity = none | 3.75
| G11 | VMR_1 | Volcanics; Stockwork Intensity = low | 0.98
| G12 | VMR_2 | Volcanics; Stockwork Intensity = medium | 0.73
| G13 | VMR_3 | Volcanics; Stockwork Intensity = high | 0.08
| G15 | UNDEF | Colluvium, Dikes, Faults, remaining lithologies | 91.55
| Copper |  |  |
| C01 | OXMX | Oxide + Mixed (oxide and sulphide) | 0.39
| C02 | SUL_0 | Sulphide; Stockwork Intensity = none | 4.22
| C03 | SUL_1 | Sulphide; Stockwork Intensity = low | 1.80
| C04 | SUL_2 | Sulphide; Stockwork Intensity = medium | 1.67
| C05 | SUL_3 | Sulphide; Stockwork Intensity = high | 0.41
| C06 | CBX | Catalina Breccia | 0.003
| C15 | UNDEF | Undefined | 91.50

17.1.2 | Histograms, Cumulative Frequency Plots, and Boxplots

Histograms and cumulative probability or probability plots display the frequency distribution of a given variable and demonstrate graphically how that frequency changes with increasing grade. Boxplots show the frequency distribution of the composite data by means of a graphical summary. These plots are useful for characterizing grade distributions, and identifying multiple populations within a data set.
Gold and copper display similar patterns in boxplots and scatterplots. Both show positively skewed lognormal distributions, mostly showing the presence of only a single population. The exception is the mixed oxide + sulphide domain for copper where the cumulative probability plot clearly shows at least two populations. Coefficient of variation (CV) values for gold range from 0.52 to 1.40, except for domain G03, which has a CV of 2.59 (Figure 17-1). Copper CV values range from 0.58 to 1.86 (Figure 17-3). Results are summarized in boxplots shown in Figures 17-1 to 17-4.

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Figure 17-1: Boxplot Summary of Gold Composite Data, Un-cut (PDTS, 2000)

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Figure 17-2: Boxplot Summary of Gold Composite Data, Cut Grades (PDTS, 2000)

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Figure 17-3: Boxplot Summary of Copper Composite Data, Un-Cut (PDTS, 2000)

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Generally, these analyses show fairly homogeneous gold and copper grades within each domain. Higher grades on average mimic the stockwork intensity level within each lithology. The Catalina Breccia contains the highest average gold and copper grades.
Figure 17-4: Boxplot Summary of Copper Composite Data, Cut Grades (PDTS, 2000)

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17.1.3 | Grade Scatterplots

Copper versus gold scatterplots were used to determine what degree of correlation exists between the two grades and if trends are evident. The plot for all composite data is shown in Figure 17-5. A certain degree of relation between the two metals exists, with a correlation coefficient of 0.5.
Grade versus X, Y, or Z coordinates scatterplots were also constructed. Results show higher gold grades generally located in the center of the deposit. Higher copper grades are generally found at depth.
Figure 17-5: Gold vs. Copper Scatterplot (PDTS, 2000)

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17.1.4 | Contact Profile Analysis

Contact plots were generated to explore the relationship between stockwork intensity levels within the same lithology, and between similar stockwork intensity level and different lithologies. The plots are constructed with software that searches for data with a given code, and then searches for data with another specified code and bins the grades according to the distance between the two points. This allows for a graphical representation of the grade trends away from a “contact.” If average grades are reasonably similar near a boundary and then diverge as the distance from the contact increases, the particular boundary should probably not be used as a grade constraint. If there is a distinct difference in the averages across a boundary, there is evidence that the boundary may be important in constraining the grade estimation.
The contact plots for gold values show similar grades between like stockwork intensity levels. Between differing intensity levels (same lithology) the trends are gently transitional, from lower to higher grades between lower to higher intensity domains. This trend becomes more acute between differing stockwork intensity levels in different lithologies. Contact plots for copper values show similar to slightly transitional trends across the copper domain boundaries.

17.1.5 | Estimation Domains

The data analyses demonstrated that most of the domains should be treated as soft boundaries with respect to gold and copper. PDTS chose a “semi-soft” philosophy to reflect the transitional nature commonly found between stockwork intensity domains of the same lithology.
The Catalina Breccia, due to its distinctly higher grades, was treated as its own interpolation domain with hard boundaries to adjacent domains with respect to gold and copper. Also the oxide and mixed unit (C01) contact was treated as a hard boundary with respect to copper. AMEC concurs with this philosophy.
The boundary philosophy for different lithologies with the same stockwork intensity was to be a transparent (i.e., no constraints on composite selection other than what is defined by the search ellipse of the particular domain). AMEC generally agrees with this, but a limited boundary sharing approach may be a better choice for some of these type of boundaries. Again, the relatively small differences in grades between lithologies means that implementing a “semi-soft” method here would not likely change the global estimate, but may result in better local estimates.

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17.2 | Evaluation of Extreme Grades

Extreme grades of copper and gold were examined using histograms, CDF plots, and decile analysis prepared by Placer Dome. Results of these analyses yielded cutting thresholds for each domain for gold and copper. These are shown in Table 17-2, along with the number of composites affected and percent metal cut. Generally, the distributions do not indicate a problem with extreme grades for copper or gold (for most domains). Selected capping levels remove about 0.5% of metal. Notable exceptions are G03 for gold, which lost 4% metal, and the high-grade Catalina Breccia domain (G09 and C06) in which 3% Au and 2% Cu metal were cut. The capped grades were applied to composited assays.
Table 17-2: Cutting Thresholds or Cap Grades for Gold and Copper Composite Data, (PDTS, 2000)

| Model Code | Cutting Value | Number of Composites Cut | Metal Content Cut
| Gold | g/t |  | %
| G01 | 2.00 | 28 | 0.5
| G02 | 3.50 | 19 | 1.0
| G03 | 0.00 | 12 | 4.0
| G04 | 3.00 | 4 | 0.5
| G05 | 10.00 | 2 | 0.5
| G06 | 3.50 | 6 | 0.5
| G07 | 4.50 | 4 | 1.0
| G08 | 2.80 | 2 | 0.5
| G09 | 30.00 | 3 | 3.0
| G10 | 1.80 | 8 | 1.0
| G11 | 2.50 | 5 | 0.5
| G12 | 2.50 | 4 | 0.5
| G13 | 3.50 | 2 | 0.5
| G15 | 2.00 | 31 | 3.5
| Copper | % |  | %
| C01 | 1.50 | 9 | 0.5
| C02 | 1.50 | 4 | 0.5
| C03 | 1.25 | 10 | 0.5
| C04 | 1.50 | 17 | 1.0
| C05 | 1.50 | 2 | 0.5
| C06 | 6.00 | 4 | 2.0
| C15 | 2.00 | 6 | 2.0

Statistical summaries for the cut composite data are shown as boxplot summary plots in Figures 17-2 and 17-4 for gold and copper, respectively. Gold CV values decreased slightly (range of 0.49 to 1.24) with the previously high G03 domain now having a CV of 0.91. Copper CV values are only slightly lower ranging from 0.57 to 1.81.

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AMEC agrees with the results and implementation by Placer Dome of the extreme grade analysis for gold and copper grades at Cerro Casale.

17.3 | Variography

Variography, a continuation of data analysis, is the study of the spatial variability of an attribute. Variography was developed by PDTS on gold grades and copper grades for each domain. The experimental variograms used in this analysis were relative pairwise variograms. For every domain, a set of variogram maps, down-the-hole variograms, omni-directional variograms, and directional variograms were calculated. The sequence and type of variograms utilized were to first investigate the presence of any strong preferred direction of grade continuity with the variogram maps in the X-Y, X-Z, and Y-Z planes. The down-the-hole variograms gave a better determination of the nugget effect and short-range continuity, while the omni-directional variogram gave a general perception of the sill and continuity range. Finally, the directional variograms gave the final directions of continuity. These were determined by doing a set of variograms at azimuth increments of 10° in the X-Y plane. After selecting the best direction of continuity in that plane, two other sets of variograms were calculated at increments of 10° in the vertical plane of that direction and in the vertical plane perpendicular to that direction. The best direction of continuity in those two planes was selected and the final and third direction of continuity was automatically defined by being perpendicular to the two previous ones. The down-hole and omni-directional variograms were good for both metals, showing well-structured variograms, while directional variograms were generally only fair.
The final three experimental variograms were modeled with double structured spherical variograms for each rock type and normalized (re-scaled) to a sill of 1.00. The modeled variogram parameters are similar for gold and copper, with copper having a slightly more prominent first structure and second range, on average. The parameters are shown in Table 17-3.
For the gold variographic analysis, rock types G03 and G04, G07 and G08, and G12 and G13 were grouped due to a lack of samples in these individual units. The main directions of continuity were found to the south-east at an azimuth ranging from 115° to 140°, and down dip at angles varying from -70° to -90°. Northeast trends for gold were also observed but in the intrusive units only (rock types G01 to G04). The second ranges for the two directions varied from 59 m to 179 m. The nugget effect is usually low and represents about 16% of the sill on average, while the first and second structures are 32% and 52% respectively on average.
In regards to the copper variographic analysis, the main directions of grade continuity are found to be from the east to southeast, with azimuths ranging from 90° to 130°, and down dip at angles varying from -70° to -90°. The second ranges along these directions are from 45 m to 179 m. The nugget effect is also low, representing about 13% of the sill, while the first and second structures represent about 41% and 46% respectively, on average.

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For both metals, the undefined domains (G15 and C15) were estimated using the respective variogram parameters of the volcanic, no stockwork intensity domain.
Table 17-3: Gold and Copper Variogram Parameters for Estimation Domains (PDTS, 2000)

| Nugget | Sills | Axis Directions (azimuth / dip) | First Structure Ranges | Second Structure Ranges
| Co | C1 | C2 | Principal (P) | Minor (M) | Vertical (V) | P1 | M1 | V1’ | P2 | M2 | V2
Gold Domains
G01 | 0.141 | 0.258 | 0.601 | 130 / 0 | 220 / -80 | 220 / 10 | 13.4 | 22.2 | 23.5 | 143.0 | 84.2 | 72.4
G02 | 0.176 | 0.327 | 0.497 | 140 / 0 | 230 / -80 | 230 / 10 | 11.9 | 25.7 | 9.7 | 128.0 | 140.0 | 94.6
G03+G04 | 0.185 | 0.224 | 0.591 | 140 / 0 | 230 / 80 | 230 / 10 | 33.6 | 17.8 | 25.7 | 82.8 | 74.9 | 74.9
G05 | 0.165 | 0.270 | 0.565 | 130 / 0 | 40 / -90 | 40 / 0 | 42.8 | 34.0 | 17.8 | 143.0 | 127.0 | 75.2
G06 | 0.156 | 0.420 | 0.424 | 135 / 0 | 225 / -70 | 225 / 20 | 13.4 | 31.1 | 17.8 | 123.0 | 138.0 | 69.6
G07+G08 | 0.212 | 0.316 | 0.472 | 115 / 0 | 205 / -85 | 205 / 5 | 33.7 | 19.9 | 14.0 | 92.9 | 85.0 | 59.3
G09 | 0.142 | 0.541 | 0.317 | 120 / 0 | 210 / -90 | 210 / 0 | 49.8 | 14.1 | 10.4 | 77.1 | 59.2 | 34.8
G10 | 0.164 | 0.364 | 0.472 | 125 / 0 | 215 / -90 | 215 / 0 | 55.3 | 31.7 | 39.6 | 138.0 | 94.7 | 65.2
G11 | 0.101 | 0.286 | 0.613 | 120 / 0 | 210 / -90 | 210 / 0 | 65.1 | 10.4 | 8.9 | 143.0 | 138.0 | 81.4
G12+G13 | 0.125 | 0.180 | 0.695 | 125 / 0 | 215 / 90 | 215 / 0 | 82.9 | 21.8 | 15.9 | 179.0 | 120.0 | 80.9
Copper Domains
C01 | 0.098 | 0.512 | 0.390 | 130 / 0 | 220 / -90 | 220 / 0 | 7.9 | 24.7 | 7.9 | 81.7 | 64.0 | 64.0
C02 | 0.117 | 0.394 | 0.489 | 120 / 0 | 210 / -90 | 210 / 0 | 87.7 | 33.5 | 22.7 | 165.0 | 93.6 | 84.7
C03 | 0.075 | 0.412 | 0.513 | 100 / 0 | 190 / -70 | 190 / 20 | 45.4 | 19.8 | 19.8 | 179.0 | 109.0 | 76.9
C04 | 0.108 | 0.432 | 0.460 | 105 / 0 | 195 / -80 | 195 / 10 | 34.5 | 34.5 | 32.5 | 175.0 | 124.0 | 96.4
C05 | 0.170 | 0.318 | 0.512 | 90 / 0 | 0 / -90 | 0 / 0 | 57.2 | 25.7 | 18.8 | 140.0 | 102.0 | 86.7
C06 | 0.234 | 0.416 | 0.350 | 125 / 0 | 35 / -90 | 35 / 0 | 33.9 | 18.8 | 18.8 | 65.8 | 45.1 | 31.1

17.4 | Estimation

Modeling for gold and copper grades consisted of grade interpolation by ordinary kriging (OK). Only capped grades were interpolated. Nearest-neighbor (NN) grades were also interpolated for validation purposes. The radii of the search ellipsoids were oriented to correspond to the variogram directions and second range distances (Table 17-3). Block discretization was 3 x 3 x 3.
A two pass approach was instituted each for gold and copper grade interpolation. The first and main interpolation was set-up so that a single hole could place a grade estimate in a block within a sparsely drilled region yet multiple holes would be used in areas of more dense drilling. Blocks needed a minimum of 6 composites in order for a block to receive an estimated grade. Maximum composite limits were set to 20. Because usage of data from multiple drill holes was not forced during the interpolation runs, AMEC and PDTS checked the model in areas likely to be Measured (i.e., areas of higher density drilling). Almost all of these blocks used the maximum number of composites, which meant, that because of the search ellipsoids used, multiple holes must have been used.

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A second pass, mimicking all parameters of the first, was run strictly for Inferred mineral resources and used 1.5 times the first pass search ellipse size.
Bulk density values were assigned into the resource model by means of the copper domains. The assigned values were: 2.40 (C01 domain), 2.65 (C02, C03, C04 and C05 domains), 2.58 (Catalina Breccia or C06 domain) and 2.61 (C15 or undefined domain).
The block model was edited to the topographic surface.

17.4.1 | Validation

Inspection of Estimation Run Files
Interpolation scripts were printed, examined, and compared to the interpolation plan and variogram parameters. No errors were found.
Visual Inspection
AMEC completed a visual validation of the Cerro Casale deposit block model. Grade interpolation was examined relative to drill hole composite values by inspecting sections and plans. The checks showed good agreement between drill hole composite values and model cell values.
Grade Variability
Placer Dome checked the smoothing in the estimates by applying a correction to the variance of the declustered composite data to reflect the change of support from core grades to block grades, and then comparing its coefficient of variation (CV) to that of the resource block estimates. This correction for change of support was accomplished with the Indirect Lognormal Correction (ILC) method. Results show that the coefficient of variability of the gold estimates is 4.3% lower than that of the corrected gold composites, while the coefficient of variation of the copper estimates is 13.8% lower than that of the corrected copper composites. In general, the amount of smoothing anticipated, given by the relative difference in coefficients of variation, varies between 10% and 30%. In this case, the gold estimates appear slightly more variable while the copper estimates have an adequate amount of smoothing. AMEC concurs with this analysis.

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Model Checks for Bias
AMEC checked the block model estimates for global bias by comparing the average metal grades (with no cutoff) from the ordinary kriged model (OK) with means from nearest-neighbor estimates. (The nearest-neighbor estimator declusters the data and produces a theoretically unbiased estimate of the average value when no cutoff grade is imposed and is a good basis for checking the performance of different estimation methods.) Results (only for blocks classified as Measured and Indicated) are displayed in Table 17-4. Results show no apparent global bias, except for an apparent low bias in kriged estimates of domain G09.
AMEC and Placer Dome also checked for local trends in the grade estimates (grade slice or swath checks). This was done by plotting the mean values from the nearest-neighbor estimate (AMEC) or declustered composite data (Placer Dome) versus the kriged results for benches, northings, and eastings swaths. The kriged estimate should be smoother than the nearest-neighbor estimate or declustered composite data, thus the nearest-neighbor estimate and declustered composite data should fluctuate around the kriged estimate on the plots. Results for gold and copper showed the two trends behaving as predicted and demonstrating no significant trends of gold or copper in the estimates.

17.5 | Mineral Resource Classification

The Mineral Resources of the Cerro Casale project were classified into Measured, Indicated, and Inferred Mineral Resources by PDTS. Parameters were chosen based on the gold variogram models. Measured Mineral Resources were set by a search ellipse defined by the first ranges of the variogram; Indicated Mineral Resources used a search ellipse defined by the second variogram ranges; and Inferred Mineral Resources were set using a search ellipse that was 1.5 times the second ranges of the respective variogram models. Only blocks that contained interpolated gold values were used in the Inferred category.

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Table 17-4: Global Model Mean Grade Values by Domain

|  | Nearest-Neighbor Estimate | Kriged Estimate | % Difference
| Gold (g/t) |  |  |
| G01 | 0.140 | 0.142 | 1.4
| G02 | 0.435 | 0.444 | 2.0
| G03 | 0.659 | 0.664 | 0.8
| G04 | 0.699 | 0.730 | 4.2
| G05 | 0.462 | 0.443 | -4.3
| G06 | 0.611 | 0.596 | -2.5
| G07 | 0.813 | 0.814 | 0.1
| G08 | 0.962 | 0.974 | 1.2
| G09 | 4.702 | 4.109 | -14.4
| G10 | 0.148 | 0.148 | 0
| G11 | 0.405 | 0.411 | 1.5
| G12 | 0.556 | 0.568 | 2.1
| G13 | 0.651 | 0.661 | 1.5
| Copper (%) |  |  |
| C01 | 0.069 | 0.067 | -3.0
| C02 | 0.080 | 0.081 | 1.2
| C03 | 0.193 | 0.194 | 0.5
| C04 | 0.279 | 0.276 | -1.1
| C05 | 0.285 | 0.295 | 3.4
| C06 | 1.181 | 1.148 | -2.9

Inspection of the model and drill hole data on plans and sections combined with spatial statistical work and validation results done by PDTS and reviewed by AMEC support this classification scheme. AMEC would recommend that in future work PDTS should directly ensure that multiple holes located within the respective search ellipse will be used in estimating Measured and Indicated Mineral Resources rather than their current indirect method. Nonetheless, AMEC finds that the Cerro Casale Mineral Resources were estimated and categorized using logic consistent with the CIM definitions referred to in National Instrument 43-101.

17.6 | Mineral Resource s

The Mineral Resources of the Cerro Casale project include material within an optimistic ultimate pit shell, which was developed by MQes using a Lerchs-Grossman algorithm and the following parameters:

| · | Gold price – $550/oz

| · | Copper prices – $1.75/lb

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| · | Mining cost - $0.80/t mined

| · | Stockpile re-handling cost - $0.29/t re-handled

| · | Processing cost - $3.31/t milled

| · | Heap leach cost - $1.85/t leached

| · | G&A cost - $0.47/t milled.

Cerro Casale Mineral Resources were last estimated in January 2000 using different metal prices and operating costs.
The current Mineral Resources for Cerro Casale are summarized in Table 17-5. Mineral resources are incremental to the mineral reserves. The current mineral resource estimate is compliant with Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000), CIM Definition Standards for Mineral Resources and Mineral Reserves (2005) and Canadian National Instrument 43-101.
Table 17-5: Cerro Casale Mineral Resources (MQes, 24 June 2006)

|  |  |  |  |  |  | Grades |  |  | Contained Metal |
| Mineral Resource Category |  |  | Tonnage (Mt) |  |  | Gold (g/t) |  |  | Copper (%) |  |  | Gold (K oz) |  |  | Copper (M lb) |
| Measured |  |  | 34 |  |  | 0.40 |  |  | 0.22 |  |  | 436 |  |  | 164 |
| Indicated |  |  | 347 |  |  | 0.40 |  |  | 0.24 |  |  | 4,460 |  |  | 1,835 |
| Measured + Indicated |  |  | 381 |  |  | 0.40 |  |  | 0.24 |  |  | 4,896 |  |  | 1,835 |
| Inferred |  |  | 301 |  |  | 0.35 |  |  | 0.25 |  |  | 3,385 |  |  | 1,657 |

Notes: 1. Mineral Resources are defined with a Lerchs-Grossman pit design based on metal prices of $550/oz Au and $1.75/lb Cu, and average G&A costs of $0.47/t milled, mining costs of $0.80/t mined, stockpile re-handling costs of $0.29/t re-handled, heap leach costs of $1.85/t leached, and plant operating costs of $3.31/t milled. 2. Mineral Resources are exclusive of Mineral Reserves. 3. Summation errors are due to rounding.
The resulting pit shell fulfills the expectation of reasonable extraction test in declaring mineral resources at Cerro Casale. AMEC agrees with this logic, implementation, and the resource estimate.

17.7 | Mineral Reserves

Reserves are calculated using metal prices of $450/oz gold, $1.50/lb copper, a total processing cost of $6.29/t for mill ore and $2.30/t for heap leaching. Mill processing recovers copper and gold, whereas oxide ore heap leaching recovers gold only. AMEC notes that the copper price used is 20 to 36% higher than the long-term (ten-year) copper prices used for reserves by AMEC and most metal price forecasters. If a long-term copper price of $1.50/lb is not achieved this could materially impact the project.

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Mill ore economic cutoff grades vary by ore types based on the different recoveries. Gold economic cutoff grades range from 0.52 to 0.70 g/t. Copper economic cutoff grades range from 0.22 to 0.23% copper.
Heap leach oxide economic cutoff grade is 0.25 g/t gold with only soluble copper grades less than or equal to 0.10 % copper considered as heap leach feed.
Cerro Casale Mineral Reserves are summarized in Table 17-6.
Table 17-6: Mineral Reserves (MQes, 24 June 2006)

|  |  |  |  |  |  | Grades |  |  | Contained Metal |
| Mineral Reserve Category |  |  | Tonnage (Mt) |  |  | Gold (g/t) |  |  | Copper (%) |  |  | Gold (K oz) |  |  | Copper (M lb) |
| Proven |  |  | 205 |  |  | 0.71 |  |  | 0.24 |  |  | 4,706 |  |  | 1,099 |
| Probable |  |  | 830 |  |  | 0.68 |  |  | 0.26 |  |  | 18,228 |  |  | 4,706 |
| Proven + Probable |  |  | 1,035 |  |  | 0.69 |  |  | 0.25 |  |  | 22,934 |  |  | 5,805 |

Notes: 1. US$450/oz Au and US$1.50/lb Cu prices used. 2. Metallurgical recovery equations are noted in Table 16-3 of this report. 3. The life-of-mine waste-to-ore strip ratio is 2.9:1. 4. Summation errors are due to rounding. 5. Mineral Reserves exclude Mineral Resources.
This reserve has a life-of-mine waste-to-ore strip ratio of 2.9:1.
Mine designs and production planning is suitable to support reserve estimates and are compliant with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000), CIM Definition Standards for Mineral Resource and Mineral Reserves (2005) and Canadian National Instrument 43-101 (2005) of the Canadian Securities Administrators. Sensitivities to variations in metal prices, operating costs and capital costs have been assessed in economic analyses.

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18.0 | OTHER RELEVANT DATA AND INFORMATION

There is no other data or information relevant to the project that is not covered in other sections of this report.

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19.0 | REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION AND DEVELOPMENT PROPERTIES

19.1 | Mining Operations

Of the two alternatives studied by MQes, the most favorable involves mining and processing of 117.8 Mt of oxide ore by heap leaching for gold, 917.6 Mt of sulphide ore by milling for gold and copper, and 3,005 Mt of waste. The mine life is comprised of 2 pre-production years and 17 production years. Oxide ore is processed at approximately 75,000 t/d and sulphide ore is milled at approximately 150,000 t/d. The ultimate pit will measure over 2,600 m from rim to rim and the highwall will have a vertical extent of over 1,500 m, ranking the proposed final pit wall among the worlds highest.
The primary crusher is located 500 m south of the ultimate pit limit. Waste dumps and low-grade stockpiles are located within 500 m of the pit entrance. The Río Nevado valley is used to store waste rock. The northern edge of the waste rock dump forms the buttress for the tailings dam. The dumps and stockpiles are built from the 4,087 m pit entrance elevation from the onset of mining. Figure 19-1 shows the site layout.
Mine plans were developed using Howard Steidtmann’s proprietary “Pit Optimization Package” or “POP!” software. AMEC considers this mine planning software to be robust, accepted by the mining industry, and appropriate for assessing the mining potential of the Cerro Casale deposit. AMEC checked the MQes results for both raw cone generation and scheduling within the ultimate pit and phases using NPV Scheduler, and obtained similar reserve estimates and production schedules.
The following sections provide summary descriptions of the key mine planning steps.

19.1.1 | Economic Modeling

Block values are calculated using costs and recoveries outlined elsewhere in this report.
Stockpiling of low grade sulphide material is used as an optimization strategy allowing higher grade mill feed to be processed earlier in the mine life. Mill ore entering the stockpile must also cover a re-handle cost of $0.294/t or it will not be stockpiled.
AMEC has valued blocks in NPV scheduler and produced a cone giving similar oxide, mill, and waste tons as reported by MQes. AMEC is confident that the blocks have been valued using the economic parameters supplied by MQes and presented in this report.

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Figure 19-1: Site Plan (PDTS, 2000)
Note: The site plan shows general location for key facilities, as conceived in the 2000 PDTS study. The pit, plant, waste dumps, and tailings impoundment are in the same locations in the current study; although, the pit and dumps are slightly larger. Heap leaching occurs in the stockpile area.

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Only Measured and Indicated Resources are treated as ore. Inferred Resources are treated as waste in the Profit Model.
AMEC reviewed the economic modelling methodology and parameters applied. They are considered to meet standard practices and appropriate for this deposit. A spreadsheet model was built to replicate the Profit Model calculation and used to check selected block values from different process groups and spatial areas within the ultimate pit. The spreadsheet calculated values corroborated the Profit Model values.
The revenue and cost parameters used in the Profit Model are as follows.
Metallurgical Recoveries
The eight rock types in the resource model were treated as distinct process groups based on their general lithology and metallurgical characteristics. Gold and copper recovery formulas were developed for each process group, as a function of the head grades (Section 16.3 and Table 16-3). The formulas were applied to the resource model gold and copper grades for blocks classified as Measured and Indicated Resources only. Inferred material was treated as waste.
Processing Cost
The costs applied by process group and downstream product costs are listed in Table 19-1. The cost of processing, administration, and plant services is estimated on a dry tonne basis.
Table 19-1: Process Costs

| Mill Processing | $3.72/t milled
| Heap Leaching | $2.30/t leached
| Freight | $54.00/t conc.
| Smelting | $78.00/t conc.
| Refining | $0.075/lb Cu
| Deduction | 1% of conc.
| Participation | 10% Cu price variance from $0.90/lb Cu
| Marketing & Insurance | 0.03% of product value
| Moisture | 10% of shipped conc.
| Losses | 0.2% of conc.
| Administration and Services | $0.65/t ore
| Plant Services | —

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Mining Cost
The mining cost and incremental mining cost per bench below the 4051 m elevation used in the Profit Model are listed in Table 19-2. All mining is performed on 15 m benches, with pit stage wall berms at 30 m intervals.
Table 19-2: Base and Incremental Mining Costs

| Item, Pit Rim 4051 | Cost $/t
| Ore delivered to Mill Base | 0.59400
| Incremental above pit rim | 0.00646
| Incremental below pit rim | 0.01089
| Ore delivered to Heap Leach Base | 0.65700
| Incremental above pit rim | 0.00646
| Incremental below pit rim | 0.01089
| Waste Base | 0.61900
| Incremental above pit rim | 0.00452
| Incremental below pit rim | 0.01089

AMEC believes the base and incremental cost to be reasonable and has verified that cone costs are similar to mining cost used in economic spread sheet.
Mine Dewatering
The pit de-watering requirement is largely unknown, although standing water is encountered in the exploration drill holes approximately 200 m to 250 m from surface. The ultimate pit will bottom at 3339 m elevation, 748 m below pit entrance elevation, and some 500 m below the elevation of the Río Nevado river valley and the base of the saturated tailing basin. The faults and fracture systems intersecting the pit walls are expected to be water bearing, and it is expected that a mine de-watering system will be required which will include a system of perimeter wells plus in-pit wells and sump systems.
De-watering requirements have an average operating cost of $0.001/t over the life of the mine, and are included in the mine operating cost estimate.
Metal Price
An average gold price of $450/oz and copper price of $1.50/lb was used in the Profit Model calculations.

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19.1.2 | Pit Shell Optimization

Wall Slopes
Piteau Associates provided pit slope recommendations for ultimate and internal phase pit design. AMEC has found that simplified versions of the recommended angles were used in the design. AMEC recommends that the ultimate pit and phase designs be reviewed by Piteau.
Optimization Methodology
The ultimate pit limits were defined using POP!, which utilizes the industry standard Lerchs-Grossman algorithm for economic pit limit definition. The input for this process consists of the Profit Model, the highwall slope constraints and the current topographical surface. The output from this process is an optimized or unsmoothed pit shell, which honors the economic and geotechnical constraints, but does not accommodate ramp access or minimum mining widths. This unsmoothed pit shell is then used a guideline for creating mineable or ‘smoothed’ pit design that includes ramps.
The ultimate pit was subdivided into ten pit phases increasing the number of phases generally increases NPV by making target scheduling constraints easier to achieve; however, there is a limit to the number of phases that can be added to a pit. This occurs when the phases become to narrow to be mined with equipment selected.
AMEC believes an opportunity exists to improve projects economics by excluding high cost incremental material from the design by increasing the bench discount rate used in developing Lerchs-Grossman pit cones. For example, AMEC suggests increasing the bench discount rate from the current calculated rate of 0.63% (5% annual rate divided by the average advance rate of 8 benches per year) to say 2.0%. Incorporating this elevated bench discount rate into the design parameters used by AMEC to check the MQes cone produces a cone with 86 Mt less ore and 439 Mt less waste. The increment lost has a strip ratio of over 5 to 1. AMEC believes that eliminating this high cost increment from the final pit design could improve the projects economics.
AMEC considers the optimization methodology to be appropriate for this deposit also feels that in a few places the phases have become to narrow to be mined and should be modified.

19.1.3 | Pit Access and Phase Design

Access to the pit from the topographic entrance level is via a 10% decline ramp. Ramps are designed with a width of 35 m to allow for a traveled road surface of 31 m (three truck widths), a containment berm of 3 m on the outside edge of the ramp, and a 1 m drainage ditch. AMEC considers this road width to be sufficient for ramps on which trolley assist is not used. A wider ramp (say 44 m) would be required if trolley assist haulage were to be considered.

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AMEC reviewed the phase designs and found for the most part all of the phases have access, although not all. AMEC recommends using the current phase designs as input for another iteration of smoothing to ensure all of the phases have access.
Figure 19-6 shows a generalized north looking sectional view of the ten phases.
Figure 19-2: North Looking Section through Pit Phases (MQes, 2006)

19.1.4 | Stockpile and Dump Design

Low-grade Ore Stockpiles
Low-grade sulphide materials will be stockpiled for reclamation and processed either later in the mine life or as necessary to maintain mill feed tonnage. The sulphide stockpile is located to the east of the plant site and constructed from the 4087 m elevation on top of a portion of the waste dump.

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Parameters used in stockpile design include the following:

| · | Density of stockpiled rock – 1.95 t/m 3

| · | Constructed overall dump slope – 37°

| · | Constructed overall stockpile height – 40 m

| · | Approximate basal area – 110 ha.

Maximum capacity of the sulphide stockpile is approximately 80 Mt.
Waste Dumps
Approximately 3,005 Mt of waste, exclusive of stockpiled ore, will be mined over the mine life. Of this total, some 80 Mt of waste rock is scheduled for tailing dam construction.
The Río Nevado valley to the east and southeast of the open pit is well situated and able to contain the waste. Valley floor elevations range from 3820 m in the north to 3725 m in the south end of the dump area. The waste is contained between the eastern and western sides of the valley. The north face of the waste dump forms the downstream backing for the tailings dam. Only the south side of the waste dump is open through a relatively narrow throat in the valley walls through which the Río Nevado flows. The waste dump area contains the requisite waste tonnage with a finished top elevation of 4160 m. The south side of the dump body is terraced to provide for a finished slope of 220 m.
Ground stability problems are not anticipated. Small berm failures at or near the edges of active dump areas are expected, but are not considered to be an impediment to dump development. Once the waste has advanced to the east wall of the river valley, the dump will be essentially contained and stable.
The parameters used in the design of the waste dumps include the following:

| · | Density of rock fill in dump – 1.95 t/m 3

| · | Angle of repose of material – 37°

| · | Constructed dump slope at south end – 22°

| · | Constructed dump slope at north end – 22°.

Acid Rock Drainage (ARD)
ARD assessment work has shown that most of the sulphur in Cerro Casale waste rock occurs as sulphate minerals which readily dissolve in water, potentially resulting in drainage waters with over 1,000 mg/L of sulphate. Some of the rock and tailings materials also have potential to release acidic drainage and associated elevated metal concentrations. Due to the relatively dry climate, minimizing contaminant transport will be the key to controlling potential ARD.

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Preliminary modelling of infiltration into the waste rock dump suggests that there will be no net infiltration for periods with average annual precipitation, and very low infiltration (10 to 15 mm/a) during years with higher than average precipitation. Compaction of the surface of the waste rock dump will reduce the infiltration values by a factor of 10. Good compaction can be accomplished by rubber-tired vehicle traffic, so haul-truck traffic patterns will be directed with this in mind.
No costs have been included for waste characterization, special handling, or segregating of waste types in this study.
Mine Access Roads
All roads are crowned and ditched to enhance drainage. Road dimensions and characteristics include the following:

| · | Width - 35 m

| · | Gradient - 10% maximum

| · | Berm - 2.0 m high on outside edge of ramp.

Dust suppression is provided by three 90 t water trucks.
Tailing s Dam Construction
The tailings dam is to be constructed by the mine using a combination of locally available materials and run-of-mine waste. Dam construction entails a downstream construction technique using run-of-mine waste rock which is hauled to location with the mine haul trucks. The downstream slope of the tailing dam initially abuts the north toe of the waste dump and is eventually covered by waste. The tailings dam and the north end of the waste dump are constructed concurrently, on an as-needed basis, to provide adequate volume in the tailings pond. Approximate tonnages and distances are included in the haulage cycle calculations to account for tailings dam construction.
The pit design is larger than previous designs, reducing the buffer zone between the tails impoundment and the pit. According to Bema, the opportunity exists to move the dam away from the pit to re-establish an appropriate buffer zone. AMEC believes this issue is of minor consequence with respect to resources and reserves, but suggests that the appropriate design revisions are included in future engineering studies.

19.1.5 | Production Schedule

AMEC reviewed the detailed production schedule and finds that it conforms to the pit design. AMEC scheduled the deposit within the MQes phases using NPV Scheduler and was able to match the MQes production schedule reasonably well. Of concern to AMEC are the bench advance rates, which peak at 12 benches per year in a few periods. Advance rates this high may be obtainable but certainly, but are not sustainable, and should be reviewed.

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AMEC recommends that another pass be completed at smoothing the ten phases and that another pass at scheduling be completed within these new phases to reduce the bench advance rates.

19.1.6 | Equipment

Cerro Casale pre-stripping and stockpiling operations in Year -2 and Year -1 yield 51 Mt (141,000 t/d) and 131 Mt (359,000 t/d), respectively. Production in Years 1 through 5 totals 215 Mt/a (589,000 t/d), which increases to an average of 314 Mt/a (860,000 t/d) for the subsequent 8 years. The total mining rate decreases annually to 52 Mt/a (144,000 t/d) from Years 14 through 16, as mining approaches the ultimate pit limits.
Open pit material movement is performed with haul trucks and a combination of electric shovels, hydraulic excavators, and large front-end loaders. The initial production fleet consists of eight 381 mm blasthole drills, thirty seven 308 t trucks, four 1,200 t class electric shovels, and four 25 m 3 rubber-tired loaders.
MQes fleet requirements are assumed to be identical to those assessed by Metalica in 2005; however, the revised production schedule has a different production profile and an overall tonnage increase of approximately 6%. As such, the Metailica equipment fleet exceeds initial requirements by 25-30% from Year -2 through Year 5. Conversely, fleet requirements are underestimated from Years 6 through 15. AMEC believes the variances described are within the level of accuracy of the cost estimates and are likely offsetting. AMEC recommends optimizing capital equipment expenditures by defining annual production requirements, haulage profiles, and mobile equipment requirements to suit the revised mine plan.
MQes capital equipment expenditures are accelerated by one year from those assessed by Metalica in 2005. Given the current state of the equipment market, this approach is probably slightly conservative in that some more common units such as dozers are likely more readily available, but long lead items such as trucks and shovels will require advanced expenditure.
Further mobile equipment expenditure opportunities may exist in considering fewer/larger front end loaders and haul trucks.

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19.2 | Recover ability

Metallurgical recovery information is discussed in detail in Section 16.3 of this report. Metallurgical recovery functions were developed for gold and copper for each major metallurgical unit. These functions are regression functions dependent on grade.
Recovery functions were used in conjunction with anticipated smelting contract conditions for the sale of the copper concentrate and doré gold, to derive the net smelter value (NSR) of the expected metal production.
AMEC verified these conditions and generally found them to be reflecting usual terms for such type of contracts. A total average smelting penalty of $3.00/t of dry concentrate is indicated but the details of which ore types, which minor elements and which scales of application and penalty rates used to derive this number are not indicated. From Section 16.1.12 of this report, the most likely element to incur smelting penalties will be mercury.

19.3 | Markets

AMEC does not envision any concerns related to marketing concentrates or doré.

19.4 | Contracts

Smelting, refining, transportation, handling, rates or charges appear to be reasonable and within industry standards.

19.5 | Environmental

According to Flavio Fuentes of Placer Dome Latin America (AMEC 2005), there are no requirements for bond posting in Chile.
Information on remediation and reclamation requirements was not available for review. A total of $16 million is budgeted for mine closure.

19.6 | Taxes

The economic analysis of the Cerro Casale project is based on a discounted cash flow analysis on a pre-tax basis.

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19.7 | Operating Cost Estimates

19.7.1 | Operating Cost Summary

Unit operating costs, updated in June 2006 are shown in Table 19-5.
Table 19-3: Unit Operating Costs

| Area | Cost ($/t) |
| Mining | 0.80 | /t ore treated (oxide, mixed & sulphide)
| Mining | 0.29 | /t stockpile re-handled
| Heap Leaching | 1.85 | /t ore leached (oxide)
| Milling | 3.31 | /t ore milled (mixed and sulphide)
| General & Administrative | 0.47 | /t ore treated (oxide, mixed & sulphide)
| Offsite Costs (concentrate) | 1.89 | /t ore milled (mixed & sulphide)
| Total | 3.12 | /t leached
| Total | 6.47 | /t milled
| Total | 6.76 | /t re-handled and milled

Notes: 1. Life of mine averages. 2. Unit costs exclude waste mining costs. 3. Total rehandle and mill cost of $6.76 applies to only the stockpiled tonnes.

19.7.2 | Mine Operating Costs

Mine operating costs are estimated on a yearly basis by determining major and support equipment requirements, including supplies, consumables, and manpower. Cost information is derived from manufacturer’s information or extrapolated from existing Placer Dome operations.
The following major cost centers are included:

| · | operating labor

| · | maintenance labor

| · | engineering and geology

| · | mine operating costs

| · | drilling

| · | blasting

| · | loading

| · | hauling

| · | roads and dumps

| · | general services

| · | pit de-watering.

Mine operating costs were further revised in 2005 by Metalica. However, the revision is based on the Placer Dome pit, production schedule, and dumps. With increased pit and dump sizes in the current study, the haulage cycles will be marginally longer; although, AMEC believes the current cycles, productivities, and costs are within the level of accuracy for the study and comparable to those of similar operations.

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Two key areas of concern in the mine operating costs were fuel and tire costs.
AMEC investigated the tire costs and found that the prices used in the cost model are equivalent to the mean of two recent quotations solicited for similar projects, and are therefore reasonable.
The fuel price in the MQes economic model is $0.50/l, which is lower than the long term prices expected for a remote job site. However, given the uncertainty in long term fuel prices and AMEC recommends evaluating sensitivity to fuel price.

19.7.3 | Processing Plant and Heap Operating Costs

Processing costs include:

| · | primary crushing and coarse ore conveying

| · | concentrator and thickening for tailings

| · | concentrate pipeline

| · | concentrate filtration and load out

| · | leach, elution and gold refining

| · | water supply systems, water reclaim and tailings

| · | camp and road maintenance, water wells.

Labor and consumable costs were revised in the June 2006 appraisals by MQes. The overall processing costs were revised to $3.31/t ore processed. Heap costs are estimated to be $1.85/t leached.
Manpower
The staffing of the plant includes metallurgical staff, operations, and maintenance personnel for the mill, heap leach, tailings and water systems, filtration plant and concentrate loading. AMEC reviewed staffing in terms of numbers in each area and found the staffing to be adequate.
Grinding Media and Liners
The costs for grinding balls and liner wear appear to be higher than expected. A credit equivalent to $0.25/t could be realized to reflect the better wear rates achieved with the improved steel metallurgy of modern grinding media and current pricing. AMEC recommends testing sensitivity to grinding media and liner costs.

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Electricity
The plant electrical costs are calculated from an equipment list with operating loads, which are adjusted for utilization. The electrical load assessment is sufficiently detailed for this type of study, but the unit cost for electrical power is of concern. Long term electrical power costs are estimated to be $0.0502/kWh, delivered from a yet-to-be-built power plant in Cardones. AMEC recommends assess the impact of an electrical power price increase to determine the potential impact to resources and reserves should the assumed price not be achievable.
Maintenance Supplies
The calculation should have been based on percentages of the indicated capital expenses per operating area and category (structural, architectural, mechanical, piping, electrical, instrumentation, etc.). This is an acceptable method of evaluating the likely requirements for maintenance parts, which would equate to a unit cost of approximately $0.90/t. Instead, MQes included a value of $0.25/t under the direction of Bema based on actual operating information and current estimates from other large scale mining projects. AMEC recommends testing sensitivity to increasing this value by $0.25/t to determine the potential impact to resources and reserves if actual maintenance supplies costs are higher than projected.

19.7.4 | General and Administr ative Operating Costs

General and administrative (G&A) operating costs include personnel, accounting, warehousing, transport of employees, human resources, insurance, and head-office allocations. Cerro Casale G&A cost estimates are reasonable.

19.8 | Capital Cost Estimates

Total capital costs by facility are provided in Table 19-6, as referenced in the Volume 4 of the June 2006 MQes study provided to AMEC.
Table 19-4: Total Project Capital Costs (MQes, 2006)

| Area |  | Cost ($ million) |
| Pre-stripping |  |  | 0.0 |
| Power Line |  |  | 35.7 |
| Heap Leach |  |  | 90.1 |
| Mine Fleet |  |  | 500.3 |
| Process Plant & Water System |  |  | 1,581.0 |
| Mine Closure |  |  | 16.0 |
| Total |  |  | 2,223.1 |

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Pre-production capital costs are shown in Table 19-7.
Table 19-5: Total Pre-production Capital Costs (MQes, 2006)

| Area |  | Cost ($ million) |
| Direct Costs |  |  |
| Plantsite & Roads |  |  | 51.8 |
| Primary Crusher |  |  | 32.9 |
| Coarse Ore Stockpile |  |  | 13.8 |
| Conveying |  |  | 37.9 |
| Grinding Facilities |  |  | 263.0 |
| Flotation Facilities |  |  | 135.0 |
| Water Supply |  |  | 115.5 |
| Shops & Warehouses |  |  | 26.2 |
| First Aid Building |  |  | 0.4 |
| General Office |  |  | 10.1 |
| Assay Laboratory |  |  | 3.7 |
| Portside Facility |  |  | 26.7 |
| Open Pit - Pre-production Stripping & Mining Equipment |  |  | 253.3 |
| Power Supply |  |  | 30.4 |
| Tailings Disposal |  |  | 45.1 |
| Concentrate Handling |  |  | 14.2 |
| Concentrate Pipeline |  |  | 59.0 |
| Accommodations |  |  | 22.5 |
| Sub-total Direct Costs |  |  | 1,141.6 |
| Indirect Costs |  |  |  |
| Vendors |  |  | 7.3 |
| Construction Overheads |  |  | 98.3 |
| Operations Overheads |  |  | 33.8 |
| Warehouse Inventory |  |  | 32.5 |
| Freight & Duties |  |  | 43.5 |
| Taxes & Duties |  |  | 12.7 |
| Project Management |  |  | 114.1 |
| Design & Engineering |  |  | 92.4 |
| Commissioning |  |  | 29.3 |
| Sub-total Indirect Costs |  |  | 463.9 |
| Total Construction Cost |  |  | 1,605.5 |
| Heap Leaching Facilities (incl. Contingency) |  |  | 90.1 |
| Cleaner Tails Leaching Facilities (incl. Contingency) |  |  | 38.6 |
| Refinery (incl. Contingency) |  |  | 9.3 |
| Cyanide Destruction (incl. Contingency) |  |  | 14.1 |
| Power Line (incl. Contingency) |  |  | 35.7 |
| Contingency |  |  | 167.6 |
| Total Capital Cost |  |  | 1,960.9 |

Notes: 1. Summation errors are due to rounding.
Sustaining capital costs are summarized in Table 19-6.

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Table 19-6: Sustaining Capital Costs (MQes, 2006)

| Area |  | Cost ($ million) |
| Mining Equipment |  |  | 247.0 |
| Mine Closure |  |  | 16.0 |
| Total |  |  | 263.0 |

19.8.1 | Capital Cost Review

AMEC reviewed capital costs for mine facilities and infrastructure using current spreadsheets, and process flowsheets and drawings from previous studies. Civil, concrete, steel, and piping drawings were not available. AMEC reviewed the estimating methods used by MQes, and compared the totals against the 2005 cost estimate and similar projects. Emphasis was given to major capital items and unit prices for each. Following are comments related to AMEC’s capital cost review.
Direct Costs
Quantities for civil works were estimated based on the general arrangement drawings developed for the project using historical unit prices available in Placer Dome database. AMEC suggests increasing civil and earthworks excavation costs by $1.0 million, to account for rock that will likely be encountered while excavating. The cost increase is based on the assumption that 10% of the detailed excavation is rock.
AMEC believes current installed concrete prices are higher than those estimated by MQes, increasing costs by $10.2 million.
AMEC believes current structural steel costs are higher than those estimated by MQes, increasing costs by $9.4 million.
Architectural costs appear to be reasonable.
AMEC believes mechanical equipment costs for the plant and mine are underestimated by $15.7 and $24.8 million, respectively. These variances are supported by recent quotes for similar plant equipment.
The concentrate pipeline costs appear to be $15.6 million low, in AMEC’s opinion.
AMEC believes electrical equipment costs are underestimated by $19.7 million, of which most is attributed to cost increases for wrap around motors.
In AMEC’s opinion, the direct costs are underestimated by a total of $96.3 million.

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Indirect Costs
Cost assessments for EPCM services, construction camp, road maintenance, property acquisition, metallurgical testing, insurance, vendor representatives, freight, assay lab spares, spare parts, import duties, owner’s costs, and commissioning are, in AMEC’s opinion, reasonable. However, AMEC recommends reducing construction equipment rental costs by $2.0 million to account for what appears to be an overestimate.
The capital cost estimate is comprised of a combination of scaled and escalated cost estimates from previous studies, some dating back to 1997. AMEC believes the contingency should be increased by $83.5 million to a total of approximately 15% of the total construction cost to account for expenditures we believe will be incurred as the mine and plant are designed and detailed cost estimates are prepared.
In AMEC s opinion, the indirect costs are underestimated by a total of $81.5 million.

19.8.2 | Sustaining Capital Cost Review

AMEC reviewed estimates for sustaining capital required over the life of the mine. These consist of mining equipment replacements and mine closure costs. Equipment selections and qualities appear reasonable relative to AMEC’s experience with similar scale projects.

19.9 | Economic Analysis

Economic analysis of the Cerro Casale project is based on a discounted cash flow analysis on a pre-tax basis, using Proven and Probable Mineral Reserves and the production plan described in Section 19. Annual revenues are calculated from the production data, plant performance data, minus capital and operating costs. Discounted cash flow analysis indicates that the project offers a positive return. The MQes cashflow model is presented in Table 19-7.

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Table 19-7: Summary Cashflow (MQes, 2006)

| Year -3 |  | Year -2 |  | Year -1 |  | Year 1 |  | Year 2 |  | Year 3 |  | Year 4 |  | Year 5 |  | Year 6 |  | Year 7
MINING DATA |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Ore Mined (Dry Tonnes) |  |  | 4,840,134 |  | 40,579,965 |  | 89,226,524 |  | 93,052,754 |  | 57,899,432 |  | 68,118,967 |  | 74,167,783 |  | 30,267,980 |  | 60,425,140
Waste Mined (Dry Tonnes) |  |  | 46,552,388 |  | 90,494,181 |  | 125,664,353 |  | 121,568,803 |  | 156,995,326 |  | 146,847,568 |  | 138,796,030 |  | 287,072,524 |  | 258,668,454
HEAP LEACH DATA |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Ore Leached (Dry Tonnes) |  |  | 4,600,000 |  | 27,400,000 |  | 38,299,999 |  | 36,040,120 |  | 7,806,240 |  | 3,331,800 |  | 149,040 |  | 156,060 |  | 18,360
Gold Recovered (kg) |  |  | 887 |  | 6,966 |  | 12,290 |  | 13,646 |  | 5,661 |  | 1,735 |  | 383 |  | 35 |  | 11
Gold Recovered (Ounces) |  |  | 28,511 |  | 223,981 |  | 395,140 |  | 438,721 |  | 182,006 |  | 55,789 |  | 12,303 |  | 1,123 |  | 364
MILLING DATA |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Ore Milled (Dry Tonnes) |  |  |  |  | 4,100,000 |  | 52,000,000 |  | 54,799,998 |  | 54,800,000 |  | 54,800,001 |  | 54,800,002 |  | 54,800,000 |  | 54,799,999
Contained Gold (kg) |  |  |  |  | 3,677 |  | 32,988 |  | 43,071 |  | 37,712 |  | 38,498 |  | 49,382 |  | 34,687 |  | 39,383
Copper (%) |  |  |  |  | 0.23% |  | 0.24% |  | 0.36% |  | 0.29% |  | 0.29% |  | 0.34% |  | 0.29% |  | 0.28%
Gold (g/t) |  |  |  |  | 0.90 |  | 0.63 |  | 0.79 |  | 0.69 |  | 0.70 |  | 0.90 |  | 0.63 |  | 0.72
Copper Recovered (t) |  |  |  |  | 8,003 |  | 104,527 |  | 165,323 |  | 137,574 |  | 136,780 |  | 158,463 |  | 137,715 |  | 130,574
Gold Recovered in Concentrate (Ounces) |  |  |  |  | 77,162 |  | 673,847 |  | 911,255 |  | 788,774 |  | 814,002 |  | 1,090,255 |  | 730,996 |  | 853,521
Silver Recovered in Concentrate (Ounces) |  |  |  |  | 154,325 |  | 1,347,695 |  | 1,822,511 |  | 1,577,549 |  | 1,628,003 |  | 2,180,511 |  | 1,461,992 |  | 1,707,041
Concentrate Grade (Cu%) |  |  |  |  | 23.86% |  | 23.92% |  | 24.86% |  | 25.86% |  | 25.81% |  | 26.34% |  | 25.44% |  | 26.45%
Concentrate, Grade (Au g/t) |  |  |  |  | 72 |  | 48 |  | 43 |  | 46 |  | 48 |  | 56 |  | 42 |  | 54
Cleaner Tails Leach |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Gold Recovered (oz) |  |  |  |  | 5,639 |  | 71,517 |  | 75,368 |  | 75,368 |  | 75,368 |  | 75,368 |  | 75,368 |  | 75,368
GROSS ANNUAL REVENUES (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Total |  |  | 12,817,318 |  | 162,594,338 |  | 834,721,894 |  | 1,150,089,320 |  | 894,320,845 |  | 846,498,312 |  | 1,018,214,589 |  | 787,341,037 |  | 820,385,274
UNIT OPERATING COSTS - On Site (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
G&A - Mill |  |  | 0 |  | 0 |  | 25,768,000 |  | 25,768,000 |  | 25,768,000 |  | 25,768,000 |  | 25,768,000 |  | 25,768,000 |  | 25,768,000
G&A - Heap Leach |  |  | 25,768,000 |  | 25,768,000 |  | 0 |  | 0 |  | 0 |  | 0 |  | 0 |  | 0 |  | 0
Mining |  |  | 42,734,349 |  | 98,612,051 |  | 147,422,357 |  | 144,453,628 |  | 139,801,203 |  | 146,524,676 |  | 153,749,864 |  | 243,051,678 |  | 236,872,895
Heap Leach |  |  | 8,510,000 |  | 50,690,000 |  | 70,854,998 |  | 66,674,222 |  | 14,441,544 |  | 6,163,830 |  | 275,724 |  | 288,711 |  | 33,966
Milling |  |  | 0 |  | 13,571,000 |  | 172,120,000 |  | 181,387,993 |  | 181,388,000 |  | 181,388,003 |  | 181,388,007 |  | 181,388,000 |  | 181,387,997
Total |  |  | 77,012,349 |  | 188,641,051 |  | 416,165,355 |  | 412,283,844 |  | 361,398,747 |  | 359,844,509 |  | 361,181,595 |  | 450,496,389 |  | 444,062,858
ANNUAL OPERATING COSTS - Off Site (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Total |  |  | 17,898 |  | 7,728,010 |  | 97,571,187 |  | 149,441,657 |  | 121,163,892 |  | 120,665,759 |  | 138,618,571 |  | 122,213,969 |  | 113,681,634
TOTAL OPERATING COST (US$) |  |  | 77,030,247 |  | 196,369,060 |  | 513,736,542 |  | 567,725,500 |  | 482,562,639 |  | 480,510,268 |  | 499,800,166 |  | 572,710,357 |  | 557,744,492
CAPITAL COSTS (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Net Cashflow (US$) | -622,859,000 |  | -632,912,928 |  | -802,274,722 |  | 298,585,443 |  | 541,563,819 |  | 374,058,206 |  | 355,288,044 |  | 518,414,423 |  | 207,030,680 |  | 251,140,782
Cum. Net Cashflow (US$) | -622,859,000 |  | -1,255,771,928 |  | -2,058,046,651 |  | -1,759,461,208 |  | -1,217,897,389 |  | -843,839,183 |  | -488,551,139 |  | 29,863,284 |  | 236,893,964 |  | 488,034,746
ECONOMIC INDICATORS |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
NPV @ 0% (US$) |  |  | 3,361,209,994 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
NPV @ 5% (US$) |  |  | 1,348,329,428 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
IRR |  |  | 13.1% |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Cash Cost Au/oz (with Cu and Ag credits) |  |  | $106.83 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Payback |  |  | 4.9 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |

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Table 19-8: Summary Cashflow (MQes, 2006)

| Year 8 | Year 9 | Year 10 | Year 11 | Year 12 |  | Year 13 |  | Year 14 |  | Year 15 |  | Year 16 |  | Year 17 |  | Total
MINING DATA |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Ore Mined (Dry Tonnes) | 64,507,908 | 49,484,909 | 55,141,222 | 69,806,624 | 26,483,395 |  | 56,820,118 |  | 87,373,559 |  | 63,401,459 |  | 43,772,672 |  | 0 |  | 1,035,370,545
Waste Mined (Dry Tonnes) | 265,474,870 | 21,2968,470 | 274,837,195 | 260,109,987 | 30,346,1633 |  | 232,822,950 |  | 60,509,295 |  | 13,874,738 |  | 8,637,875 |  | 0 |  | 3,005,356,640
HEAP LEACH DATA |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Ore Leached (Dry Tonnes) | 2,160 | 0 | 0 | 0 | 0 |  | 0 |  | 0 |  | 0 |  | 0 |  | 0 |  | 17,803,779
Gold Recovered (kg) | 1 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  | 41,615
Gold Recovered (Ounces) | 38 |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  | 1337,977
MILLING DATA |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Ore Milled (Dry Tonnes) | 54,799,999 | 54,800,001 | 54,799,999 | 54,799,999 | 54,800,000 |  | 54,799,999 |  | 54,799,999 |  | 54,800,000 |  | 54,800,001 |  | 39,466,769 |  | 917,566,766
Contained Gold (kg) | 46,496 | 41,419 | 44,498 | 43,627 | 25,431 |  | 30,657 |  | 45,488 |  | 42628 |  | 30301 |  | 17,370 |  | 647,314
Copper (%) | 0.36% | 0.34% | 0.21% | 0.31% | 0.25% |  | 0.25% |  | 0.28% |  | 0.27% |  | 0.29% |  | 0.19% |  | 0.29%
Gold (g/t) | 0.85 | 0.76 | 0.81 | 0.80 | 0.46 |  | 0.56 |  | 0.83 |  | 0.78 |  | 0.55 |  | 0.44 |  | 0.71
Copper Recovered (t) | 173,105 | 163,273 | 101,685 | 149,993 | 119,940 |  | 118,252 |  | 131,402 |  | 128,377 |  | 137,743 |  | 64,987 |  | 2,267,717
Gold Recovered in Concentrate (Ounces) | 1,070,417 | 940,387 | 1,008,477 | 1,004,634 | 534,315 |  | 678,909 |  | 1,035,047 |  | 994,885 |  | 698,335 |  | 368,457 |  | 14,273,677
Silver Recovered in Concentrate (Ounces) | 2,140,835 | 1,880,774 | 2,016,955 | 2,009,269 | 1,068,630 |  | 1,357,818 |  | 2,070,094 |  | 1,989,771 |  | 1,396,670 |  | 736,913 |  | 28,547,354
Concentrate Grade (Cu%) | 26.97% | 26.56% | 26.90% | 26.94% | 26.45% |  | 26.90% |  | 27.13% |  | 26.72% |  | 26.41% |  | 27.18% |  | 26.23%
Concentrate, Grade (Au g/t) | 52 | 48 | 83 | 56 | 37 |  | 48 |  | 66 |  | 64 |  | 42 |  | 48 |  | 57
Cleaner Tails Leach |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Gold Recovered (oz) | 75,368 | 75,368 | 75,368 | 75,368 | 75,368 |  | 75,368 |  | 75,368 |  | 75,368 |  | 75,368 |  | 54,280 |  | 1,261,960
GROSS ANNUAL REVENUES (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Total | 1,048,920,192 | 959,989,896 | 801,206,821 | 948,134,832 | 643,671,719 |  | 703,684,593 |  | 904,654,473 |  | 877,195,077 |  | 772,346,927 |  | 390,423,393 |  | 14,577,210,939
UNIT OPERATING COSTS - On Site (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
G&A - Mill | 25,768,000 | 25,768,000 | 25,768,000 | 25,768,000 | 25,768,000 |  | 25,768,000 |  | 25,768,000 |  | 25,768,000 |  | 25,768,000 |  | 25,768,000 |  | 438,056,000
G&A - Heap Leach | 0 | 0 | 0 | 0 | 0 |  | 0 |  | 0 |  | 0 |  | 0 |  | 0 |  | 51,536,000
Mining | 259,164,188 | 218,183,849 | 247,703,081 | 282,731,904 | 315,578,471 |  | 265,593,727 |  | 147,777,013 |  | 93,238,673 |  | 70,382,742 |  | 11,445,363 |  | 3,265,021,711
Heap Leach | 3,996 | 0 | 0 | 0 | 0 |  | 0 |  | 0 |  | 0 |  | 0 |  | 0 |  | 217,936,991
Milling | 181,387,997 | 181,388,003 | 181,387,997 | 181,387,997 | 181,388,000 |  | 181,387,997 |  | 181,387,997 |  | 181,388,000 |  | 181,388,003 |  | 130,635,005 |  | 3,037,145,995
Total | 466,324,181 | 425,339,852 | 45,485,078 | 489,887,900 | 522,734,471 |  | 472,749,724 |  | 354,933,010 |  | 300,394,673 |  | 277,538,745 |  | 167,848,368 |  | 7,009,696,698
ANNUAL OPERATING COSTS - Off Site (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Total | 1,484,73,783 | 141,151,169 | 89,242,964 | 129,149,328 | 103,219,409 |  | 101,358,874 |  | 113,379,516 |  | 111,794,400 |  | 11,9051,239 |  | 55,321,989 |  | 1,983,245,247
TOTAL OPERATING COST (US$) | 614,797,964 | 566,491,022 | 544,102,041 | 619,037,229 | 625,953,880 |  | 574,108,597 |  | 468,312,526 |  | 412,189,072 |  | 396,589,985 |  | 223,170,357 |  | 8,992,941,945
CAPITAL COSTS (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Net Cashflow (US$) | 406,022,228 | 321,898,875 | 256,604,780 | 324,797,603 | 17,317,838 |  | 124,075,996 |  | 433,541,947 |  | 459,906,004 |  | 368,756,942 |  | 160,253,035 |  |
Cum. Net Cashflow (US$) | 894,056,973 | 1,215,955,848 | 1,472,560,628 | 1,797,358,231 | 1,814,676,069 |  | 1,938,752,065 |  | 2,372,294,012 |  | 2,832,200,016 |  | 3,200,956,959 |  | 3,361,209,994 |  |
ECONOMIC INDICATORS |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
NPV @ 0% (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
NPV @ 5% (US$) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
IRR |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Cash Cost Au/oz (with Cu and Ag credits) |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |
Payback |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |  |

Project No.: 152187 | Page 19-18 |
6 August 2006 |  |

Table 19-9: Metal Price Sensitivity Analysis (MQes, 2006)

| Copper | Economic |  | Gold Price ($/oz-Au) |  |  |
| ($/pound) | Factors | Units | 400 | 425 | 450 | 475 | 500
| 1.00 | NPV5% | M$ | ($297) | ($51) | $194 | $439 | $684
| IRR | % | 2.8% | 4.6% | 6.4% | 8.0% | 9.5%
| Cash Cost | $/oz-Au | $224 | $224 | $224 | $224 | $224
| Payback | Years | 14.1 | 10.2 | 8.6 | 7.6 | 6.9
| 1.25 | NPV5% | M$ | $281 | $526 | $771 | $1,016 | $1,262
| IRR | % | 6.9% | 8.5% | 10.0% | 11.4% | 12.8%
| Cash Cost | $/oz-Au | $166 | $166 | $166 | $166 | $166
| Payback | Years | 8.2 | 7.4 | 6.7 | 5.8 | 5.0
| 1.50 | NPV5% | M$ | $858 | $1,103 | $1,348 | $1,594 | $1,839
| IRR | % | 10.4% | 11.8% | 13.1% | 14.4% | 15.7%
| Cash Cost | $/oz-Au | $107 | $107 | $107 | $107 | $107
| Payback | Years | 6.5 | 5.6 | 4.9 | 4.7 | 4.4
| 1.75 | NPV5% | M$ | $1,435 | $1,680 | $1,926 | $2,171 | $2,416
| IRR | % | 13.4% | 14.7% | 15.9% | 17.1% | 18.3%
| Cash Cost | $/oz-Au | $48 | $48 | $48 | $48 | $48
| Payback | Years | 4.9 | 4.7 | 4.4 | 4.2 | 4.0
| 2.00 | NPV5% | M$ | $2,012 | $2,257 | $2,503 | $2,748 | $2,993
| IRR | % | 16.2% | 17.4% | 18.5% | 19.7% | 20.8%
| Cash Cost | $/oz-Au | ($11) | ($11) | ($11) | ($11) | ($11)
| Payback | Years | 4.4 | 4.2 | 4.0 | 3.8 | 3.6

The MQes economic model does not include an allocation for working capital; however, when AMEC applied standard estimates for working capital to its sensitivity analysis, the internal rate of return remained positive. All other inputs are appropriate and, apart from the first few years of development, all future annual cash flows are positive.
After incorporating some potential cost increases identified by AMEC in a sensitivity analysis, the IRR and cumulative cash flows remain positive.
As with many projects of this type, the Cerro Casale project is most sensitive to changes in metal price and rather less so to changes in operating cost and capital expenditures.
In addition to computing the IRR, the NPV at several discount rates was computed. An appropriate discount rate to use is the owner’s cost of capital. Excluding financing charges implies that the project is financed with all equity.
In AMEC’s opinion, the level of detail used in the economic analysis is appropriate for a feasibility study.

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19.10 | Payback

MQes financial analyses indicate the base case mine plan has a positive return, with a payback period of 4.9 years. AMEC agrees with this assessment, but feels the payback could increase to 8 years with incorporation of all the recommended sensitivity items.

19.11 | Mine Life

The mine life is currently estimated to be approximately 17 years, excluding the pre-production period, which AMEC believes is reasonable.

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20.0 | INTERPRETATIONS AND CONCLUSIONS

20.1 | Technical Basis

A large part of the technical support for mineral resource estimates, mineral reserve estimates, metallurgy, project design, operating cost estimates, capital cost estimates, environmental studies, and permitting are documented in a 2000 Feasibility Study by Placer Dome. Capital cost estimates were updated by PDTS in February 2004. Capital and operating cost estimates were again updated by PDTS and Metalica in 2005, and by MQes in 2006. AMEC reviewed the June 2006 capital and operation costs and provide our opinions as of that date. The technical basis for mineral resources and mineral reserves meet the requirements of Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000) and Canadian National Instrument 43-101.

20.2 | Mineralization and Alteration

Gold-copper mineralization occurs in quartz-sulphide and quartz-magnetite-specularite veinlet stockworks developed in the dioritic to granodioritic intrusives and adjacent volcanic wall rocks. The geology is well understood and is documented with appropriate geological mapping and drill hole logging. Modelling of ore controls is suitable to support resource estimates.

20.3 | Drilling

In general, drilling equipment and procedures conform to industry standard practices and have produced information suitable to support resource estimates. Sample recovery, to the extent documented, was acceptable. Collar surveying was of suitable accuracy to ensure reliable location of drill holes relative to the mine grid and other drill holes. Downhole surveys of RC and core holes are not complete and locally downgrade the confidence in the position of individual intercepts of deep mineralization. Holes not surveyed are dominated by RC holes testing oxide mineralization less than 200 m deep.

20.4 | Sampling, Sample Preparation, and Assaying

Logging of RC drill cuttings and core followed procedures suitable for recording lithology, alteration, and mineralization in a porphyry deposit. AMEC found the quality of logging to be generally professional and interpretations of lithology and stockwork veining intensity to honor original logs. Geological data and interpretations are suitable to support resource estimates.

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Sample collection and handling of RC drill cuttings and core was done in accordance with industry standard practices, with procedures to limit sample losses and sampling biases.
Sample preparation and assay protocols generally met industry standard practices for gold and copper, although the 150 g split for pulverization in 1991 through 1994 is substandard for gold analyses and resulted in poorer precision compared to subsequent years.
Gold was determined on a one assay-ton aliquot (29.116 g) by fire assay with either a gravimetric or atomic absorption finish. Copper and silver were obtained from a 2 g sample aliquot by atomic absorption after an aqua regia digestion. Assay methods conform to industry standard practices.

20.5 | Assay QA/QC

Assay QA/QC protocols were observed throughout all drilling campaigns, with blind SRMs, blanks and duplicates being inserted into the sample series since the inception of CMA’s RC drill programs in 1993. Monitor Geochemical Laboratories used internal quality control procedures for assays in 1991 through 1994.
Acceptable assay accuracy and precision are indicated for drilling programs from 1991 to 1997 based on detailed audits by MRDI and Smee and Associates.
AMEC independently evaluated QA/QC data for 1998 and 1999 drilling campaigns. Assays of SRMs show suitable accuracy. Assays of pulp duplicates indicate a precision for gold of ±19% and ±6% for copper at the 90 th percentile, which is marginally acceptable for gold. Assays of SRMs in 1999 show erratic patterns, but pulp duplicates indicate a preparation and assay precision for gold and copper the same as 1998. Analyses of blanks show contamination of up to 1.3 g/t Au during sample preparation for batches 135 to 234. These are mostly for holes in prospects other than Cerro Casale, but do include assays for Cerro Casale core hole CCD111 and geotechnical holes 99GT003-006. Gold grades above the 0.4 g/t internal cutoff are present in holes 99GT003, 99GT006 and CCD111. These should be considered to be suspect until verified by re-assaying. Coarse reject material should be reassayed for these holes prior to the next resource estimate update.
Check assays by PDRC, Vancouver suggests that Bondar Cleggs’ Au assays are biased 5% to 10% high, depending on the sample batch. This is more than generally acceptable, but can be used provisionally used in a feasibility study.
AMEC reviewed all previous analyses of QA/QC data by MRDI and Smee and Associates and agrees with their conclusions. With the exception of some remedial work required for holes CCD111 and geotechnical holes 99GT003 and 99GT006 (representing a small percentage of resource blocks), assays are of sufficient accuracy and precision to support resource estimates.

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AMEC did not independently sample drill core and obtain commercial assays of check samples. This was not considered to be necessary given the extent of historical blind QA/QC undertaken by CMA and Placer Dome (Section 13.3) and the level of independent auditing of sampling and assaying by MRDI in 1994 through 1997.

20.6 | Density

Bulk density values for ore and waste units are based on 877 measurements made on core samples in 1995 and 1996 by E.C. Rowe and Associates, in 1996 and 1997 by CMA personnel, and in 1998 by Placer Dome. Bulk densities are assigned by a combination of lithology, stockwork intensity, and degree of oxidation. Methods conform to industry standard practices and are suitable for estimates of tonnage.

20.7 | Data Verification

Geological, geotechnical, and analytical information were developed over a period of multiple exploration programs between 1991 and 1999, involving Bema Gold, CMA, MRDI, and Placer Dome staff. Entry of information into databases utilized a variety of techniques and procedures to check the integrity of the data entered. With the exception of one period of drilling, assays were received electronically from the laboratories and imported directly into drill hole database spreadsheets.
MRDI (1997a) audited 5% of entries for geological attributes and assays against original logs and certificates for the 1991 to early 1996 drilling campaigns and found an error rate of 0.2%. MRDI (1997b) again audited the database for 1996 and 1997 drilling and found an error rate of 0.294%. AMEC audited all of 1998 and 1999 drilling data from Placer Dome and found no errors for assays and lithology for 1558 entries (4.5%).
The assay and geological databases are suitable to support resource estimates.

20.8 | Geological Interpretations

AMEC reviewed cross section and plan interpretations of lithology, stockwork intensity, oxidation, and potassic alteration and found these to conform reasonably to original logged information. Some smoothing was practiced to produce outlines suitable to use in resource estimates. Interpretations are reasonable and in concept are consistent with porphyry gold-copper deposits.

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20.9 | Metallurgical Processing

Metallurgical test work categorized ore types on the basis of metallurgical characteristics for comminution, optimal grind size, flotation response, cyanidation of tails (for gold) and trace element content. While the definition of the Cerro Casale Project continues to evolve, there is no new technical information that materially alters the validity of any previous work. Plant design concepts are reasonable and reflect typical, state-of-the-art, design concepts and application practices. The resultant sizing of major equipment items were examined reviewed by AMEC, with special attention to test work and design issues regarding grinding and flotation.

20.9.1 | Comminution

| · | Equipment selections (including both the two line and three line options) are based on generally-accepted application principles and practices.

| · | A two-line grinding plant is technically feasible but may carry a higher operating risk when encountering significant maintenance tasks such as mill relines.

| · | A two-line mill will also need a more carefully defined design basis including expected variations in grindability, mineral liberation size, and optimum grinds between and amongst rock types.

| · | It is important to continue to map the grinding characteristics of the Cerro Casale deposit. This may require the assessment of a wider variety of composites and combinations of rock types.

| · | The JK Drop Test Program may have overestimated the required SAG mill grinding power. The program should be repeated with larger core or bulk samples to fully understand the effects of rock size on impact grinding.

20.9.2 | Flotation

| · | Equipment selections are based on generally-accepted application principles and practices.

| · | Flotation test results are based on a 120 micron grind. A new set of bench scale tests is recommended to include results for the coarser grind now specified for the two-line grinding plant.

| · | Rougher flotation capacity is very conservatively applied (up to 40 minutes versus 20 to 25 minutes typical of comparable operations.

| · | Regrinding to 30 microns prior to cleaner flotation operations appears to be consistent with mineralogy notes of exceptionally fine-grained mineralization.

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20.9.3 | Dump Leaching of Oxide Ores

| · | Dump leaching of oxide ores will derive revenues from rock that previously represented a cost for removal and storage.

| · | Expected recoveries have been appropriately discounted to account for leaching of run-of-mine ores.

| · | Estimates of revenues and costs are reasonable.

20.9.4 | Other Plant and Process Issues

| · | The complete metallurgical complex also includes unit operations for thickening and filtration, cyanidation leaching of cleaner flotation tailings, and slurry transport of mineral concentrate.

| · | These operations, including flowsheets and equipment sizing, have been defined according to accepted laboratory test procedures and industry practices.

| · | AMEC considers this work to be appropriate for feasibility level evaluations.

20.10 | Mineral Resource and Mineral Reserve Estimates

20.10.1 | Resource Classification

The mineral resources of the Cerro Casale project were classified into Measured, Indicated, and Inferred Mineral Resources by PDTS. Parameters were chosen based on the gold variogram models. Measured Mineral Resources were set by a search ellipse defined by the first ranges of the variogram; Indicated Mineral Resources used a search ellipse defined by the second variogram ranges; and Inferred Mineral Resources were set using a search ellipse that was 1.5 times the second ranges of the respective variogram models. Only blocks that contained interpolated gold values were used in the Inferred category.
Inspection of the model and drill hole data on plans and sections combined with spatial statistical work and validation results done by PDTS and reviewed by AMEC support this classification scheme.

20.10.2 | Mineral Resources

Mineral resource estimates were done in 1999 from 3-dimensional block models utilizing Placer Dome in-house mine planning software (OP). PDTS concluded that the Cerro Casale gold model would be best represented by a combined lithologic-stockwork intensity model, whereas the copper model should be a combination of lithology-oxidation level-stockwork intensity parameters. AMEC concurs with this philosophy for development of geologic models or domains for use in grade interpolation at Cerro Casale.

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Domains were treated as soft boundaries with respect to gold and copper. PDTS chose a “semi-soft” philosophy to reflect the transitional nature commonly found between stockwork intensity domains of the same lithology. The Catalina Breccia, due to its distinctly higher grades, was treated as it own interpolation domain with hard boundaries to adjacent domains with respect to gold and copper. Also the oxide and mixed unit (C01) contact was treated as a hard boundary with respect to copper. AMEC concurs with this philosophy.
Capping thresholds for extreme grades of copper and gold were determined using histograms, CDF plots, and decile analysis. Generally, the distributions do not indicate a problem with extreme grades for copper nor gold (for most domains). Selected capping levels remove about 0.5% of metal. Notable exceptions are G03 for gold, which lost 4% metal, and the high-grade Catalina Breccia domain in which 3% Au and 2% Cu metal were cut. The capped grades were applied to composited assays.
Modelling for gold and copper grades consisted of grade interpolation by ordinary kriging (OK). Only capped grades were interpolated. Nearest-neighbor (NN) grades were also interpolated for validation purposes. The radii of the search ellipsoids were oriented to correspond to the variogram directions and second range distances. Block discretization was 3 x 3 x 3. A two pass approach was instituted each for gold and copper grade interpolation. Blocks needed a minimum of 6 composites in order for a block to receive an estimated grade. Maximum composite limits were set to 20. A second pass, mimicking all parameters of the first, was run strictly for Inferred mineral resources and used 1.5 times the first pass search ellipse size.
Bulk density values were assigned into the resource model by means of the copper domains. This is appropriate.
AMEC validated PDTS resource estimates using inspection of estimation run files, inspection of block grade sections and plans cross validation using change of support, and inspection for local biases using nearest-neighbor estimates on spatial swaths through the deposit. These checks showed no biases or local artifacts due to the estimation procedures.
AMEC reviewed premises used to derive the economic value of the contained metals, based on expected recoveries and smelting terms applied. AMEC also reviewed processing costs and their application to the net value function of ore blocks. These were properly developed.
MQes developed estimates of Cerro Casale mineral resources based on material within an optimistic Lerchs-Grossman pit shell. The pit shell was developed using metal prices of $550/oz gold, $1.75/lb copper, and estimated operating costs of $0.80/t mined, $0.29/t re-handled, $3.31/t milled, $1.85/t leached, and a G&A cost of $0.47/t milled. The pit shell fulfills the expectation of reasonable extraction test in declaring mineral resources. AMEC agrees with this logic and its implementation.

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Mineral Resources estimates comply with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000), CIM Definition Standards for Mineral Resources and Mineral Reserves (December 11, 2005) and Canadian National Instrument 43-101 of the Canadian Securities Administrators (December 30, 2005).

20.10.3 | Mineral Reserves

Mineral reserves are estimated using an elevated cutoff grade strategy for the ten mining phases, along with stockpiling low and high-grade ores during pre-production and normal production phases. A net revenue block model, referred to as the Profit Model, classified each block as ore or waste. AMEC agrees with this approach.
MQes prepared Cerro Casale mineral reserves estimates based on material within a designed pit (with roads), which was based on a Lerchs-Grossman pit shell developed using metal prices of $450/oz gold, $1.50/lb copper, and estimated operating costs of $0.80/t mined, $0.29/t re-handled, $3.31/t milled, $1.85/t leached, and a G&A cost of $0.47/t milled.
The life-of-mine waste-to-ore strip ratio is 2.9:1. Heap leach mining rates (oxides) start at 4.6 Mt/a in Year -2, peak at 38.3 Mt/a in Year 1, and decline through Year 8 as the oxide mineral reserves are exhausted. Sulphide and mixed ores are stockpiled in Year -1, produced at 45 Mt/a in Year 1, and at 54 Mt/a thereafter, with the exception of Years 3, 6, 9, 12, 15, 16, and 17 when part of the pit production is offset by treating stockpiled ores. The mine life is 17 years, excluding the pre-production period. The designed pit fulfills the expectation of reasonable extraction test in declaring mineral resources. AMEC agrees with this logic and its implementation.
Mineral reserve estimates comply with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000), CIM Definition Standards for Mineral Resources and Mineral Reserves (December 11, 2005) and Canadian National Instrument 43-101 of the Canadian Securities Administrators (December 30, 2005).

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20.11 | Mine Designs and Production Plans

Conventional open pit methods are planned for Cerro Casale. The mine plan features a ten-stage open pit, which is scheduled to deliver a nominal 150,000 t/d over a 17 year mine life. The final pit phase measures over 2,600 m in diameter and the highest sector of highwall will have a vertical extent of 1,500 m, ranking the proposed final highwall among the worlds tallest.
MQes used an economic model for pit designs, which incorporates metallurgical recoveries and processing costs by ore type, incremental mining costs, mine dewatering and geotechnical parameters. AMEC reviewed the economic modelling methodology and parameters applied. They are considered to be standard practice and appropriate for this deposit. An independent check by AMEC confirmed the results.
Pit designs use a 15 bench height; however, the block model was created using 17 m high blocks, which were re-blocked to 15 m high for design purposes. AMEC believes this approach is reasonable, but recommends revising the block model to use standard 15 m high blocks.
MQes revised the mine design to remove unbroken inter-ramp slopes in excess of 350 m vertical height, as recommended in the 2005 Technical Report.
Equipment selections are generally appropriate for the mine design, production rate and production schedule, but projected equipment availabilities are at the high end of rated capacities. The opportunity exists to reduce the fleet size by using larger trucks and shovels.
Generally speaking, the work done by MQes and Metalica appears to be of high quality and appropriate for the design stage and nature of the deposit. No fatal flaws were detected. Most areas where AMEC had concerns with the 2005 mine plan have been addressed. Compared to the 2005 mine plan, AMEC believes this plan contains less risk and appears to be achievable.

20.12 | Operating Cost Estimates

MQes developed operating cost estimates from information provided by Metalica, Bema, current operating costs from similar Chilean mining operations, new quotations, and in-house data. AMEC approves of the methodology used and the results obtained, with the following exceptions that AMEC would use to test project economic sensitivity:

| · | AMEC believes the fuel costs in the mine operating cost estimate are low. AMEC recommends using a long term fuel price of $0.60/l, which is the mean of two months diesel fuel costs at a nearby mine.

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| · | In AMEC’s opinion, the maintenance spares operating cost of $0.25/t milled is low. We recommend increasing this cost to $0.50/t, which is approximately half way between a standard factored estimate of 5% of the total direct field costs and the assumed value in the current economic model. AMEC also recommends evaluating this item in detail in further engineering studies.

| · | In AMEC’s opinion, process operating costs for consumables such as grinding liners and balls is overestimated. AMEC recommends reducing operating costs by $0.25/t.

20.13 | Capital Cost Estimates

MQes estimated total project capital costs to be $2,223 million, of which $1,961 million are pre-production costs and $263 million are sustaining capital costs. AMEC reviewed capital estimates and found them to be appropriate, with the following exceptions, which AMEC evaluated in a sensitivity analysis.

| · | In AMEC’s opinion, direct capital costs are underestimated by a total of $96.3 million. This value is comprised of increases in excavation, concrete, steel, equipment, concentrate pipeline, and electrical equipment costs.

| · | AMEC believes the indirect capital costs are underestimated by a total of $81.5 million. This value is comprised of an increase in contingency costs with a minor offset from an equipment rental cost reduction.

| · | The MQes model excludes working capital costs. AMEC suggests incorporating working capital into the cashflow model. Using 25% of the change in annual operating costs, working capital fluctuates over the project life, peaking at $177 million in Year 12. The remaining amount of $106 million is recovered in Year 17.

| · | AMEC believes the MQes model should include $25.8 million in G&A costs in Year -3.

| · | The MQes model excludes royalties. AMEC suggests incorporating $3 million production royalty to Minera Anglo American Chile Limitada in Year 1.

| · | The MQes model excludes an $80 million lump sum purchase payment to Barrick. AMEC suggests incorporating this cost in Year -3.

20.14 | Economic Analysis

The MQes economic analysis of the Cerro Casale project is based on a discounted cash flow analysis on a pre-tax basis, using Proven and Probable Mineral Reserves and annual production plans. Projections for annual revenues and costs are based on data developed for the mine, process plant, capital expenditures, and operating costs.
Discounted cash flow analysis indicates that the project offers a positive return. Payback period is 4.7 years. Life-of-mine is 17 years.

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As with many projects of this type, the Cerro Casale project is most sensitive to changes in metal price and rather less so to changes in operating cost and capital expenditures.
In AMEC’s opinion, the level of detail used in the economic analysis is appropriate for a feasibility study; although, we believe the capital and operating costs could be higher reducing the expected project value.

20.14.1 | Sensitivity Analysis

The MQes economic model does not appear to include an allocation for working capital; however, when standard estimates are used for working capital, there is an impact on payback but the internal rate of return remains positive.
All other inputs are appropriate and, apart from the first few years of development, all future annual cash flows are positive.
In addition to the recommended operating and capital cost revisions, AMEC suggests removing revenues generated from silver recovery, as there is insufficient data to support the in situ or recovered grades. After incorporating all of the potential cost increases the IRR and cumulative cash flows remain positive.
In addition to computing the IRR, the NPV at several discount rates was computed. An appropriate discount rate to use is the owner’s cost of capital. Excluding financing charges implies that the project is financed with all equity.

20.15 | Permitting and Environmental Studies

In accordance with legislative requirements of the Government of Chile described in Law N° 19.300 (Law on the General Basis on the Environment) and its regulations as outlined in Supreme Decree N° 30 (Regulation on the Impact Assessment System), environmental studies were conducted for the Cerro Casale Project and an Environmental Impact Study (EIS) was presented to the Regional Environmental Commission (COREMA) on 12 March 2001. Following a documented review process, approval was granted by COREMA on 1 February 2002 through “Resolución Exenta N° 014”. Through this document, the Cerro Casale Project has thus obtained the main environmental authorization required under Chilean legislative requirements.
The next stage of legislative compliance process is outstanding and will require the project to seek sectorial permits granted by the various agencies that have authority over environmental resources and construction, operation and closure of project infrastructure.
The future supplier of electrical power will need to obtain environmental permits for construction of power lines. It is reasonable to expect that administrative approval of power supply infrastructure will be granted.

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Compañía Minera Candelaria will need to obtain permits for CMA to build additional port facilities for concentrate shipping. It is reasonable to expect that CMC will negotiate terms for use of the port and that the necessary permits for construction of CMA facilities will be granted by the Chilean government.
Although there remains some exposure in that environmental permits remain to be secured for power lines and port facilities, and additional work is required regarding ARD potential of waste rock and potential downstream effects of tailings facilities, it is reasonable to expect that future permits will be granted and any potential environmental effects of waste rock and tailings, if determined to exist, can be addressed via design changes.
There are no existing impediments to obtaining easements for rights of way for access roads, water pipelines or concentrate pipelines.
There is still uncertainty regarding if mine wastes will produce ARD. The potential for elevated concentrations of base metals such as copper and zinc is yet to be determined. ARD assessment work to date has shown that most of the sulphur occurs as sulphate minerals which readily dissolve in water, and could potentially result in drainage waters that carry over 1,000 mg/L of sulphate. Preliminary models of waste rock water infiltration, however, show that there will be no net infiltration in periods with average annual precipitation and low (10 mm/a to 15 mm/a) infiltration in years with higher than average precipitation. ARD potential deserves additional study.
Impacts on surrounding water systems from water take operations conducted in the Piedra Pomez well field. Permits for use of ground water in the Piedra Pomez basin have been granted by the DGA. Groundwater exploration programs carried out by Placer Dome contractors have identified the Piedra Pomez basin as an endorreic system, or closed topographic and hydromorphic basin, based on geochemical studies. The geology of the basin is such that the basin may not be closed geohydrologically. Additional work may be warranted to confirm the lack of a hydrological connection with surrounding surface water systems.
Downstream impacts from operation of tailing impoundment and waste rock dump facilities. The tailings impoundment is based on conceptual designs and further study of the potential of seepage from the impoundment should be carried out in the future. The potential downstream impact of ARD should be revisited once more information regarding ARD potential is developed.

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| 21.0 | RECOMMENDATIONS

AMEC recommends preparing a comprehensive feasibility-level analysis, which includes a revised block model, updated mining plans, current processing concepts, and first principle cost estimates to increase confidence in the financial model and support a decision to further evaluate or develop the project.
Additional detailed recommendations for the Cerro Casale follow:

| · | Performing additional test work to confirm estimated copper and gold recovery rates.
| · | Performing additional regrinding tests, including jar tests, vendor tests, and simulations, to increase confidence in the design criteria and cost estimates.
| · | Performing additional test work to ensure the specified cleaner tails leach circuit will efficiently and economically handle the expected metal contents.
| · | Further testing of thickening and filtration concepts to further optimize these concepts and potentially reduce the associated costs.
| · | Preparing a revised block model with 15 m high blocks that correlate to the selected mining bench height.
| · | That multiple holes located within the respective search ellipses are used in estimating Measured and Indicated mineral resources rather than the presently used indirect method.
| · | That the ultimate pit and phase designs be reviewed by Piteau, with consideration given to the revised bench height.
| · | Constructing a geotechnical block model, which will allow the generation of pit designs that will more closely honor the geotechnical engineer’s inter-ramp angle, bench face angle, and berm width recommendations.
| · | Using the current pit phase designs as input for another iteration of smoothing to ensure all of the phases have access.
| · | Performing another pass at scheduling within the new pit phases to reduce the bench advance rates.
| · | Optimizing mine capital equipment expenditures by defining annual production requirements, haulage profiles, and mobile equipment requirements to suit the current mine plan.
| · | Further evaluation of the site layout with consideration give to the new pit design (and known mineral resources) and the tailings/pit buffer zone, plant location/layout, waste dump location, stockpile location, and heap leach pad location.
| · | Further evaluation of the selected truck/loader/shovel combinations.

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| · | Evaluating long term fuel prices and incorporating the results into the project planning and economic models.
| · | Further evaluating grinding media and liner costs, and incorporating the results into the project economic model.
| · | Negotiating energy supply contracts increase confidence in long term power availability and costs.
| · | That further evaluation of ARD potential be performed in order to reduce the level of uncertainty associated with the currently available ARD assessment and whether design changes in the waste rock facility are warranted
| · | Further evaluating maintenance supply parts costs.
| · | Preparing comprehensive capital cost estimates using current labor, materials, and equipment costs to increase confidence the project capital cost estimates.
| · | Developing a detailed silver model and complete test work to determine silver recoveries by grade and rock type.
| · | Performing additional test work such as reviewing the database, assaying sample rejects, and performing metallurgical tests to validate silver revenues in the MQes project economic model.
| · | Using a comprehensive project economic model with royalties, working capital, and taxes.
| · | The discount rate used to perform economic evaluations.

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22.0 | REFERENCES

AMBIMET LTDA., 1999, Mediciones de Calidad de Aire por Partículas PM10, Proyecto Aldebarán, Informe Final Campaña de Monitoreo Invierno 1999, Santiago, Chile, Diciembre 1999.
AMBIMET LTDA., 2000a, Mediciones de Calidad de Aire por Material Particulado Sedimentable, Proyecto Aldebarán, Informe Final Campaña de Monitoreo Período Julio 1999 a Marzo 2000, Santiago, Chile, Junio 2000.
AMBIMET LTDA., 2000b, Informe Meteorológico Anual 1999, Proyecto Aldebarán, Santiago, Chile, Mayo 2000.
AMBIMET LTDA., 2001, Informe Meteorológico Anual 2000, Proyecto Aldebarán, Santiago, Chile, Febrero 2001.
AMEC Americas Limited, March 2005, Cerro Casale Project, Chile, Technical Report and Qualified Persons Review, prepared for Placer Dome Inc.
Bechtel Mining and Metals, Capital and Operating Cost Review and Update, Cerro Casale Project, February, 2004.
Bema Gold Corporation, June 16, 2006, Purchase and Sales Agreement.
Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2000, CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines, CIM Standing Committee on Reserve Definitions, adopted by CIM Council, August 20, 2000.
Canadian Institute of Mining, Metallurgy and Petroleum (CIM), 2005, CIM Definition Standards for Mineral Resources and Mineral, CIM Standing Committee on Reserve Definitions, adopted by CIM Council, December 15, 2005.
CDN Water Management Consultant Inc., 2000, Proyecto Aldebarán, Modelo Hídrico y de Contaminantes, Vancouver, Canadá, Septiembre 2000.
E.C. Rowe, 2000, Depósito de Relave Cerro Casale, Memoria Descriptiva del Proyecto, Santiago, Chile, Octubre 2000.
EDRA, 1999, Hidrogeología Sector Quebrada Piedra Pómez, Santiago, Chile, Agosto 1999 (tres volúmenes).
Gobierno de Chile, Ley 19.300 Bases Generales sobre el Medio Ambiente.

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G&T Metallurgical Services, 1999, A Program of Flotation and Modal Studies - Project KM817, private report prepared for PDTS, April 1999.
G&T Metallurgical Services, 2000, An Assessment of Flotation Response - Project KM1011, private report prepared for PDTS, January 2000.
Gustavo Mieres y Juan Carlos Torres-Mura, 1999, Proyecto Aldebarán, Línea Base Vegetación, Flora y Fauna, Santiago, Chile, Septiembre 1999.
Kvaerner Metals, March 1997, Final Report - Basic Engineering, Cerro Casale Gold Project, report prepared for Compañia Minera Aldebaran
Metalica Consultores S.A., April 2005, Estudio do Planificacion, Minera Proyecto Aldebaran, private report prepared for Place Dome Latin America
Mineral Resources Development, Inc., 1997a, Oxide Feasibility Study, Cerro Casale Gold Project, Chile, private report prepared for Arizona Star Resource Corporation.
Miguel Cervellino, 1999, Proyecto Aldebarán, Línea Base del Patrimonio Cultural, Copiapó, Chile, Julio 1999.
Miguel Cervellino, 2000, Línea Base del Patrimonio Cultural para el Estudio de Impacto Ambiental del Proyecto Aldebarán. Emplazamiento de Sitios Patrimoniales en el Sector de Instalaciones Portuarias, Almacenamiento y Carguío en Punta Padrones, Costa de Caldera, Copiapó, Chile, Noviembre 2000.
Mine and Quarry Engineering Services, Inc., May 2006, Project Development Appraisal Studies, Cerro Casale Project, private report prepared for Bema Gold Corporation
Ministerio Secretaria General de la Presidencia de Chile, 2001, D.S. No 95 Reglamento del Sistema de Evaluación de Impacto Ambiental, 2001.
Mineral Resources Development, Inc., 1994, 1994 Exploration Program for the Aldebarán Property, private report prepared for Arizona Star Resource Corporation, October 1994
Mineral Resources Development, Inc., 1997a, Oxide Feasibility Study, Cerro Casale Gold Project, Chile, private report prepared for Arizona Star Resource Corporation.
Mineral Resources Development, Inc., 1997b, Preliminary Feasibility Study, Oxide and Sulphide, Cerro Casale Gold Project, Chile, private report prepared for Arizona Star Resource Corporation.
Mineral Resources Development, Inc., 1997c, Deep Sulphide Scoping Study, Cerro Casale Gold Project, Chile, private report prepared for Arizona Star Resource Corporation.

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Piteau Associates, 1999, Aldebarán Project, Cerro Casale Sulphide Deposit, Feasibility Geotechnical Assessments for the Open Pit, private report prepared for Compañía Minera Aldebarán.
Placer Dome Technical Services, 2000, Aldebaran Project, Chile: Feasibility Study, private report prepared for Compañia Minera Aldebaran
Placer Dome Technical Services, March 2004, Aldebaran Project, Chile: Feasibility Study Update, private report prepared for Compañia Minera Aldebaran
SENES Chile S.A., 1999a, Informe Final de Línea Base Vialidad e Infraestructura, Santiago, Chile, Septiembre 1999.
SENES Chile S.A., 1999b, Informe Final de Línea Base de Línea Base Geología, Geomorfología y Riesgo Geológico, Santiago, Chile, Septiembre 1999.
SENES Chile S.A., 1999c, Informe Final de Línea Base Socioeconómica, Santiago, Chile, Septiembre 1999.
SENES Chile S.A., 1999d, Informe Final de Línea Base de Suelos, Santiago, Chile, Septiembre 1999.
SENES Chile S.A., 1999e, Informe Final de Línea Base de Clima, Santiago, Chile, Agosto 1999.
SENES Chile S.A. 2000a, Informe Final Estudio de Impacto Vial, Proyecto Aldebarán, Santiago, Chile, Diciembre 2000.
SENES Chile S.A., 2000b, Informe Final Línea Base de Calidad de Aire, Santiago, Chile, Julio 2000.
SENES Chile S.A., 2000c, Informe Final Estudio de Línea Base Uso de Recursos, Santiago, Chile, Septiembre 2000.
SENES Chile S.A., 2001a, Línea de Base y Evaluación de Impacto Ambiental sobre el Valor Paisajístico, Noviembre 2001.
SENES Chile S.A., 2001b, Estudio de Impacto Ambiental Proyecto Aldebarán, Diciembre 2001.
Smee, B.W., May 1997. A Review of Quality Control Procedures and Results, Cerro Casale Project, Copiapó, Chile, private report prepared for Arizona Star Resource Corporation.

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Water Management Consultants Ltda., 1999, Aldebarán Preliminary (Phase I) Site Hydrology/Hydrogeology Scoping Study, Santiago, Chile, Diciembre 1999, con Resumen en Español.

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23.0 | DATE AND SIGNATURE PAGE

The undersigned prepared this Technical report, titled NI-43-101 Technical Report,Cerro Casale Project, Chile, dated 06 August 2006, in support of the public disclosure of Mineral Resources for the Cerro Casale property as of 24 July 2006. The format and content of the report are intended to conform to Form 43-101F1 of the National Instrument (NI 43-101) of the Canadian Securities Administrators.

| Signed and Sealed |  |
| /s/ Larry B. Smith |  |
| Larry B. Smith |  | 06 August 2006
| Signed and Sealed |  |
| /s/ William A. Tilley |  |
| William A. Tilley |  | 06 August 2006

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