<B>ARIZONA STAR CERRO CASALE TECHNICAL REPORT

EX-99 30 amectechnicalreport43101.htm TECHNICAL REPORT FOR THE CERRO CASALE PROJECT <B>ARIZONA STAR CERRO CASALE TECHNICAL REPORT

IMPORTANT NOTICE

Recognizing that Arizona Star Resource Corporation (ASR) has legal and regulatory obligations in a number of global jurisdictions, AMEC Americas Limited (AMEC) consents to the filing of this report with any stock exchange and other regulatory authority and any publication by ASR, including electronic publication on ASR 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 ASR by AMEC Americas Limited (AMEC). The quality of information, conclusions, and estimates contained herein is consistent with the level of effort involved in AMEC’s services, based on: i) information available at the time of preparation, ii) data supplied by outside sources, and iii) the assumptions, conditions, and qualifications set forth in this report. This report is intended to be used by ASR, subject to the terms and conditions of its contract with AMEC. That contract permits ASR 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.

	ARIZONA STAR RESOURCE CORP.
	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

CONTENTS

1.0	SUMMARY	1-1
	1.1 Introduction	1-1
	1.2 Project Description	1-2
	1.3 Project Location and Climate	1-2
	1.4 Project Ownership, Mineral Rights and Water Rights	1-4
	1.5 Permitting and Environmental Studies	1-5
	1.5.1 Environmental Studies and Environmental Impact Study Approval	1-5
	1.5.2 Additional Environmenta Permits and Approvals	1-6
	1.6 Geology	1-6
	1.7 Mineralization and Alteration	1-7
	1.8 Drilling Programs	1-7
	1.9 Sample Preparation and Assaying	1-8
	1.10 Assay Quality Assurance and Quality Control (QA/QC)	1-9
	1.11 Density	1-9
	1.12 Data Verification	1-10
	1.13 Geological Interpretations	1-10
	1.14 Metallurgical Processing	1-10
	1.14.1 Introduction	1-10
	1.14.2 Metallurgical Tests	1-11
	1.14.3 Throughput Capacity	1-11
	1.15 Mineral Resource and Mineral Reserve Estimates	1-12
	1.15.1 Mineral Resource and Mineral Reserve Statements	1-12
	1.15.2 Mineral Resource Estimation Procedures	1-13
	1.15.3 Resource Classification	1-14
	1.15.4 Mineral Reserves	1-15
	1.16 Mining Designs and Production Plans	1-16
	1.17 Operating Costs	1-17
	1.17.1 Total Unit Costs	1-17
	1.17.2 Mine Operating Costs	1-18
	1.17.3 Process Operating Costs	1-18
	1.17.4 Maintenance	1-19
	1.17.5 General and Administration	1-19
	1.18 Capital Costs	1-19
	1.19 Economic Analysis	1-22
2.0	INTRODUCTION AND TERMS OF REFERENCE	2-1
	2.1 Introduction	2-1
	2.2 Terms of Reference	2-2
	2.3 Units of Measure	2-2
	2.3.1 Common Units	2-2
	2.3.2 Common Chemical Symbols	2-5
3.0	DISCLAIMER	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 Introduction	4-6
	4.5.2 Environmental Approval of Power Supply Infrastructure	4-7
	4.5.3 Environmental Approval of Port Facilities	4-7
	4.5.4 Acid Rock Drainage (ARD) Potential	4-7
	4.5.5 Impacts on Surrounding Water Systems from Water Take 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 Location, Access, Local Resources and Infrastructure	5-1
	5.2 Climate, Physiography, Vegetation, and Fauna	5-3
	5.3 Environmental Impact Assessment	5-3
	5.3.1 Archaeological Assessment	5-3
	5.3.2 Fauna	5-5
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
	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	9-11
	9.4 Cerro Roman	9-11
	9.4.1 Geology	9-11
	9.4.2 Alteration and Mineralization	9-11
	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 Introduction	10-1
	10.2 Cerro Casale	10-1
	10.3 Eva	10-1
	10.4 Cerro Roman	10-2
	10.5 Estrella	10-2
	10.6 Anfiteatro	10-2
	10.7 Romancito	10-2
	10.8 Other Areas	10-2
11.0	DRILLING	11-1
	11.1 Introduction	11-1
	11.2 Drilling Methods	11-5
	11.2.1 Reverse Circulation Drilling	11-5
	11.2.2 Diamond Drilling Equipment	11-5
	11.3 Geological Logging Practices	11-6
	11.3.1 Introduction	11-6
	11.3.2 Reverse Circulation Chip Logging	11-6
	11.3.3 Core Logging	11-7
	11.3.4 Geotechnical Logging	11-7
	11.4 AMEC Review of Logging	11-8
	11.5 Core and RC Recovery	11-8
	11.6 Topography	11-8
	11.7 Drill Hole Collar Surveys	11-9
	11.8 Downhole Surveys	11-10
12.0	SAMPLING METHOD AND APPROACH	12-1
	12.1 Introduction	12-1
	12.2 Reverse-Circulation Drill Sampling	12-1
	12.3 Drill Core Sampling	12-2
	12.4 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 Introduction	14-1
	14.1.2 Data for 1991 to Early 1996 Drillng Campaigns	14-1
	14.1.3 Data for Late 1996 through 1997 Drilling Campaign	14-2
	14.1.4 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 Introduction	16-1
	16.2 Metallurgical Testwork	16-2
	16.2.1 Lithological Considerations	16-2
	16.2.2 Mineralogy	16-3

	16.2.3 Communution		16-4
	16.2.4 Selection of Optimum Grind Size		16-7
	16.2.5 Flotation		16-7
	16.2.6 Gravimetric Concentration		16-10
	16.2.7 Cyanidation		16-10
	16.2.8 Thickening		16-11
	16.2.9 Filtration and Transportable Moisture Limits		16-11
	16.2.10 Slurry Rheology		16-11
	16.2.11 Water Treatment		16-12
	16.2.12 Concentrate Minor Elements Composition		16-12
	16.3 Metallurgical Recoveries		16-12
	16.4 Net Smelter Return Calculations		16-14
	16.5 Plant Design Criteria		16-15
	16.5.1 Throughput Capacity		16-15
	16.5.2 Design Feed Grades		16-16
	16.5.3 Crushing and Grinding		16-16
	16.5.4 Flotation		16-17
	16.5.5 First Cleaner Tails Leaching		16-18
	16.5.6 Thickening		16-19
	16.5.7 Filtration		16-19
	16.5.8 Water Treatment		16-19
	16.6 Discussion		16-20
17.0	MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES		17-1
	17.1 Mineral Resource Estimates		17-1
	17.1.1 Introduction		17-1
	17.1.2 Geologic Models and Data Analysis		17-1
	17.1.3 Histograms, Cumulative Frequency Plots, and Boxplots		17-2
	17.1.4 Grade Scatterplots		17-6
	17.1.5 Contact Profile Analysis		17-7
	17.1.6 Estimation Domains		17-8
	17.2 Evaluation of Extreme Grades		17-8
	17.3 Variography		17-9
	17.4 Estimation		17-12
	17.4.1 Validation		17-12
	17.5 Mineral Resource Classification and Summary		17-13
	17.6 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 Mine Plan		19-1
	19.1.1 Economic Modelling		19-1
	19.1.2 Pit Shell Optimization		19-5
	19.1.3 Pit Stage Design		19-7
	19.1.4 Stockpile and Dump Design		19-13
	19.1.5 Production Schedule		19-15
	19.1.6 Equipment		19-17
	19.1.7 Mine Operations Manpower Requirements		19-17
	19.1.8 Major Equipment Productivity Estimates		19-18
	19.2 Metal Recovery		19-19
	19.3 Environmental Bonds, Remediation, and Reclamation		19-20
	19.4 Operating Costs		19-20
	19.4.1 Summary		19-20
	19.4.2 Mine Operating Costs	19-20
	19.4.3 Processing Plant Operating Costs	19-21
	19.4.4 Maintenance	19-23
	19.4.5 General and Administration	19-23
	19.5 Capital Costs	19-24
	19.5.1 Total Project Capital Costs	19-24
	19.5.2 Capital Cost Review	19-25
	19.5.3 Sustaining Capital Cost Review	19-29
	19.6 Economic Analysis	19-29
20.0	INTERPRETATIONS AND CONCLUSIONS	20-1
	20.1 Technical Basis for Resources and Reserves	20-1
	20.2 Permitting and Environmental Studies	20-1
	20.3 Mineralization and Alteration	20-1
	20.4 Drilling Programs	20-2
	20.5 Sampling, Sample Preparation, and Assaying	20-2
	20.6 Assay QA/QC	20.3
	20.7 Density	20-3
	20.8 Data Verification	20-4
	20.9 Geological Interpretations	20-4
	20.10 Metallurgical Processing	20-4
	20.11 Mineral Resource and Mineral Reserve Estimates	20-5
	20.11.1 Resource Estimates	20-5
	20.11.2 Resource Classification	20-6
	20.11.3 Mineral Reserves	20-6
	20.12 Mining Designs and Production Plans	20-7
	20.13 Operating Costs	20-8
	20.14 Capital Costs	20-8
	20.15 Economic Analysis	20-8
21.0	RECOMMENDATIONS	21-1
22.0	REFERENCES	22-1

TABLES

Table 1-1:	Cerro Casale Project Mineral Resource Summary – January 2000 (From PDTS, 2000)	1-15
Table 1-2:	Proven and Probable Mineral Reserves as of January 2000 using Elevated Profit Cutoff Strategy (from PDTS, 2000 Feasibility Study)	1-16
Table 1-3:	Unit Operating Costs	1-18
Table 1-4:	Distribution of Processing Operating Costs	1-18
Table 1-5:	Total Estimated Capital Costs	1-20
Table 1-6:	Post-Construction Sustaining Capital Costs (from PDTS, 2004)	1-21
Table 4-1:	Area of Interest	4-2
Table 4-2:	Mineral Concessions within Aldebaran 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 (from MRDI, 1997b)	13-9
Table 13-2:	Acme and Chemex Analyses of Standard, 1996-1997 (from MRDI, from 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
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:	Mineable Reserves with Smoothed Pit (@0.25$/t profit cutoff)	16-2
Table 16-2:	Overview of Measured Comminution Parameters	16-5
Table 16-3:	Placer Dome Model of Predictive Metallurgical Equations (from PDTS, 2000, Table 2.2)	16-13
Table 16-4:	Design Retention Time Derivation for Flotation Stages	16-17
Table 17.1:	Gold and Copper Geologic Models or Domains, Cerro Casale Project	17-2
Table 17-2:	Cutting Thresholds or Cap Grades for Gold and Copper Composite Data, Cerro Casale Project	17-9
Table 17-3	Gold and Copper Variogram Parameters for Cerro Casale Project Estimation Domains	17-11
Table 17-4	Global Model Mean Grade Values by Domain, Cerro Casale Project	17-14
Table 17-5:	Cerro Casale Mineral Resources – January 2000	17-15
Table 17-6:	Elevated Profit Cutoff Strategy (from PDTS, 2000)	17-16
Table 17-7:	Proven and Probable Mineral Reserves as of January 2000 using Elevated Profit Cutoff Strategy (from PDTS, 2000 Feasibility Study)	17-16
Table 19-1:	Process Costs	19-3
Table 19-2:	Base and Incremental Mining Costs	19-4
Table 19-3:	Recommended Slope Criteria (from Piteau Associates, 1999)	19-5
Table 19-4:	Simplified Inter-ramp Slope Angle Criteria (from PDTS, 2000)	19-6
Table 19-5:	Unit Operating Costs	19-20
Table 19-6:	Distribution of Processing Operating Costs	19-22
Table 19-7:	Total Estimated Capital Costs by Facility	19-24
Table 19-8:	Post-Construction Sustaining Capital Costs (from PDTS, 2004)	19-25

FIGURES

Figure 1-1:	Location of the Cerro Casale Gold-Copper Deposit, Northern Chile	1-3
Figure 4-1:	Mineral Claims and Area of Interest, Aldebaran (from 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 (from PDTS, 2000)	7-2
Figure 7-2:	Surface Geological Map of Cerro Casale (from PDTS, 2000)	7-5
Figure 7-3:	Cross Section 850E, Looking Northwest, Cerro Casale Deposit (from PDTS, 2000)	7-6
Figure 7-4:	Redox Units, Section 850E (from PDTS, 2000)	7-9
Figure 9-1:	Major Gold-Copper Occurrences in the Aldebaran Property (from PDTS (2000)	9-2
Figure 9-2	Measured + Indicated Gold Resources, Section 472200E (from PDTS, 2000)	9-4
Figure 9-3:	Measured + Indicated Copper Resources, Section 472200E (from PDTS, 2000)	9-5
Figure 9-4:	Measured + Indicated Gold Resources, 3832 Elevation (from PDTS, 2000)	9-6
Figure 9-5:	Measured + Indicated Copper Resources, 3832 Elevation (from PDTS, 2000)	9-7
Figure 9-6:	Intensity of Stockwork Veining, Section 850E (from PDTS, 2000)	9-8
Figure 9-7:	Potassium Feldspar Alteration, Section 850E (from PDTS, 2000)	9-9
Figure 9-8:	Geological Map of the Eva Deposit (from PDTS, 2000)	9-12
Figure 9-9:	Cross Section of the Eva Deposit (from PDTS, 2000)	9-13
Figure 9-10:	Geological Map of the Cerro Roman Deposit (from PDTS, 2000)	9-14
Figure 9-11:	North-South Cross Section of the Cerro Roman Deposit (from PDTS, 2000)	9-15
Figure 11-1:	Drill Collar Locations (from PDTS, 2000)	11-2
Figure 11-2:	Average and Median Drill Spacing by Elevation (from PDTS, 2000)	11-4
Figure 11-3:	Drill-Hole Collar Movements	11-9
Figure 13-1:	Relative Differences for Rig Duplicates (from MRDI, 1997b)	13-5
Figure 13-2:	Checks of Acme Gold Assays by Chemex (from MRDI, 1997b)	13-7
Figure 13-3:	Precision from Chemex Check Assays of Acme Gold Assays (from MRDI, 1997b)	13-7
Figure 13-4:	Chemex Check Assays of Acme Copper Assays (from 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 STD 12 – 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 STD 18 – 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
Figure 13-36:	Boxplot of All Density Measurements by Oxidation Categories	13-35
Figure 16-1:	Flotation Circuit Configuration for Laboratory Testwork	16-9
Figure 17-1:	Boxplot Summary of Gold Composite Data (Un-cut), Cerro Casale Project (from PDTS, 2000)	17-3
Figure 17-2:	Boxplot Summary of Gold Composite Data (Cut Grades), Cerro Casale Project (from PDTS, 2000)	17-4
Figure 17-3:	Boxplot Summary of Copper Composite Data (Un-cut), Cerro Casale Project (from PDTS, 2000)	17-5
Figure 17-4:	Boxplot Summary of Copper Composite Data (Cut Grades), Cerro Casale Project (from PDTS, 2000)	17-6
Figure 17-5:	Gold vs. Copper Scatterplot, Cerro Casale Project (from PDTS, 2000)	17-7
Figure 19-1:	Site Plan Map (from PDTS, 2000)	19-2
Figure 19-2:	Stage 1 Pit Design (from PDTS, 2000)	19-9
Figure 19-3:	Stage 2 Pit Design (from PDTS, 2000)	19-10
Figure 19-4:	Stage 3 Pit Design (from PDTS, 2000)	19-11
Figure 19-5:	Stage 4 (Ultimate) Pit Design (from PDTS, 2000)	19-12
Figure 19-6:	North Looking Section through Pit Stages (Generated by AMEC from surfaces provided by Placer Dome)	19-13

APPENDICES

Appendix A: List of Significant Assays

Project No.: 148562
June 2005

	ARIZONA STAR RESOURCE CORP.
	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

1.0

SUMMARY

1.1

Introduction

Arizona Star Resources Corporation (ASR) commissioned AMEC Americas Limited (AMEC) to review mineral resource and mineral reserve estimation work done to date for the Cerro Casale Project gold-copper project, northern Chile. AMEC was to determine if this estimation work was carried out in accordance with industry standard practices and if the mineral resource and mineral reserve estimates in the 2000 Feasibility Study and March 2004 Feasibility Study Update (described below) are compliant with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000) and Canadian National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators. The format and content of the report are intended to conform to Form 43-101F1.

Larry B. Smith, P.Geo., Dr. Stephen Juras, P.Geo., Jay Melnyk, P.Eng., and Pierre Lacombe, P.Eng., employees of AMEC, served as Qualified Persons responsible for preparation of this report. Larry Smith and Jay Melnyk visited the property 12 and 13 January 2005 and reviewed pertinent aspects of geology, exploration data, geological models, land status, infrastructure and mine designs. Lydia Létourneau, Manager Environmental Services for AMEC (Peru) S.A. visited the property on 15 January 2005 and reviewed environmental base-line studies, project environmental provisions and permitting. Ms. Létourneau reviewed additional documentation of environmental studies, well field studies and general permits at the offices of Compañía Minera Cerro Casale S.A. in Copiapo, Chile on 14, 16, and 17 January 2005. Roberto Aburto Durán, Manager of Engineering, AMEC International (Chile) visited the site on 12 and 13 January and reviewed infrastructure requirements and site conditions. Larry Smith and Jay Melnyk 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. Manuel Romero, an employee of AMEC International (Chile) and Roberto Aburto Durán reviewed supporting documentation for capital cost estimates at Placer Dome offices in Santiago. Dr. Stephen Juras obtained electronic copies of resource estimates and reviewed estimation procedures with Placer Dome staff in Vancouver, B.C. Canada. Pierre Lacombe obtained copies of metallurgical studies and process designs from Place Dome staff and reviewed these in AMEC’s Montreal, Quebec office. Dr. Ted Eggleston, P.Geo., an independent consulting geologist, reviewed historical assay quality assurance and quality control information for drilling campaigns in 1998 and 1999. Larry Smith reviewed historical assay quality assurance and quality control information for drilling campaigns from 1991 to 1997. Graham Wood, an employee of AMEC, reviewed cash flow estimates and financial evaluations at AMEC’s Vancouver office using information supplied by Placer Dome. All information was reviewed in sufficient detail to ensure that mineral resource and reserve estimates contained in the 2000 Feasibility Study and the March 2004 Feasibility Study Update comply with NI 43-101.

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 Technical Services Limited (PDTS). Capital cost estimates were updated by PDTS in March 2004. AMEC provides its opinions of these capital costs as of that date. Increases in capital and operating costs occurred through the remainder of 2004. In addition, there are opportunities for improvements that may come from re-examination of the technical design of the project. These changes will impact the project, but the degree to which for each cannot be determined until the technical and cost aspects of the project are again updated in 2005.

Significant assistance in preparation of electronic files and historical data were provided by Placer Dome staff, including Murray Canfield, David Prins, Horst Altschwager, Gonzalo Saldias, Flavio Fuentes, Victor Manuel Gonzalez, Oscar Munoz, Juan Carlos Rivera, Marc Jutras, John Folinsbee, and Michael Gingles.

Unless stated otherwise, all quantities are in metric units and currencies are expressed in constant 2004 US dollars.

“Placer Dome” is used in this report to refer generally to one or more companies within the Placer Dome Group of companies, including Placer Dome Latin America, a division of Placer Dome (CLA) Limited (PDLA), PDTS, and/or Placer Dome Exploration Inc. (PDX).

1.2

Project Description

Compañía Minera Casale (CMC), owned by the Bema Shareholders Group and the Placer Dome Shareholders Group, envisions a conventional open pit and milling operation producing 150,000 t/d of gold and copper ore from a porphyry gold-copper deposit. Ore will be processed in a semi-autogenous grinding (SAG) mill and flotation concentrator. Concentrates will be conveyed to the port at Punta Padrones near Caldera via a 250 km pipeline. Water for mine, mill and camp facilities will be pumped from well fields 120 km northeast of the project. Total mine life will be 18 years.

1,3

Project Location and Climate

The Cerro Casale gold-copper deposit is located in Region Three of northern Chile. The city of Copiapo is 145 km northwest of the deposit (Figure 1-1). The approximate geographic coordinates of the project are 27°47' S and 69^o 17' W. The international border separating Chile and Argentina is located approximately 20 km to the east.

The project 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 located at an elevation of 4450 m.

Figure -:

Location of the Cerro Casale Gold-Copper Deposit, Northern Chile

[amectechnicalreport43101008.jpg]

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

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.

The terrain surrounding the Cerro Casale deposit is adequate for construction of administration, camp, and mine facilities, as well as mill, concentrator, tailings and waste rock disposal facilities.

The project is approximately 180 km by road from Copiapo. The initial 25 km is paved highway leading south from Copiapo. After this, a 155 km gravel road winds its way through the Andes Mountains to site. Total driving time from Copiapo to site is about 3 hours.

Copiapo is served by a national airport with daily flights from Santiago. The city has most major services and utilities and serves as a regional centre for this part of Chile. The population of Copiapo is approximately 120,000 in habitants.

1.4

Project Ownership, Mineral Rights and Water Rights

The Cerro Casale Project is owned by CMC, a contractual mining company formed under the laws of the Republic of Chile. The share capital of CMC is indirectly owned by PDI as to 51%, Arizona Star Resource Corporation (ASR) as to 24.99% and Bema Gold Corporation as to 24.01%. The relationship of the CMC shareholders is governed by an Amended and Restated Shareholders’ Agreement dated as of June 5, 2003. The General Manager of the Project is Placer Dome Latin America, a division of Placer Dome Limited.

CMC owns 30 claim groups containing 4,105 patented mining claims and totalling 19,955 ha. Some of these claims partially overlap each other, reducing the actual ground covered by all patented mining claims to an area of 19,520 ha. All mineral rights are protected according to Chilean law, by payment of a mining patent.

Water exploration concessions are held in three areas: Piedra Pomez, Pedernales and Cerro Casale. Piedra Pomez and Pedernales are located 121 km and 210 km, respectively, north of Cerro Casale.

CMC holds permits for 17 wells drilled at Piedra Pomez with a total yield of 1,237.62 L/s. This area is expected to be the principal source of water for the Cerro Casale project.

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

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 from 1.0% to 3.0% Net Smelter Return based on the gold price ($425 to $600/oz, respectively).

After commencement of production, Placer Dome Latin America will be paid a net smelter royalty in accordance with the Net Smelter Royalty Agreement between CMC and Placer Dome Latin America dated 27 September 2004.

1.5

Permitting and Environmental Studies

1.5.1

Environmental Studies and Environmental Impact Study Approval

On-going environmental studies for the Cerro Casale Project were initiated by CMC in 1998. The scope of these studies includes baseline assessments of the main environmental components comprised of physical (surface and groundwater quality, hydrology, hydrogeology, soil, air, meteorology, etc.), biological (vegetation and fauna), cultural (archaeological), and human resources. Engineering assessments, impact evaluations, and development of environmental management plans also form part of environmental studies developed for the project. The study area covered the location of all project components including the proposed water supply well field located in the Piedra Pomez sector, the water pipeline from Piedra Pomez to Cerro Casale, mine site components (open pit, waste rock dump, tailings impoundment, support infrastructure and camp) in the Cerro Casale sector, the concentrate pipeline from Cerro Casale to the proposed port site at Punta Padrones and the proposed port site itself.

These studies led to the preparation of the Environmental Impact Study (EIS) presented to the Government of Chile’s responsible authority, COREMA, on 12 March 2001. Following a documented review process, approval for this EIS was granted on 1 February 2002. Through this approval the project has secured an important environmental authorization.

Based on AMEC’s review of the project, five items have been identified as potential environmental exposures that will require more study as the project advances. These include:

Environmental Approval of Power Supply 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.

Environmental Approval of Port Facilities. 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.

Acid Rock Drainage (ARD) potential. 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.

1.5.2

Additional Environmental Permits and Approvals

The next step in relation to the environmental process will be to obtain sectorial permits from the various agencies (refer to Chapter 4 for detail) that have authority over environmental resources and construction, operation and closure of project infrastructure.

1.6

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 centres 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, as well as numerous other smaller mineral prospects. The volcanic rocks overly 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 (Figure 7-1). Composite volcanic centres 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 centres. Gold-copper mineralization is principally associated with intense quartz-sulphide stockworks, potassic, and phylliic alteration.

1.7

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. 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.

1.8

Drilling Programs

Reverse-circulation (RC) and core drilling was carried out 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 totalling 122,747 m support the resource estimate for Cerro Casale. 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 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 below 200 m. Core holes are from 30 m to 1,473 m deep. Drilling tools produced NC (61 mm), HQ (61 mm), NQ (45 mm) and HX (63 mm) cores. Core recovery is poorly documented but appears to have exceeded 95%.

Most RC and core holes were drilled from the south to north 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 in shallow oxide mineralization is approximately 45 m. Average drill spacing in the core of the deposit in the interval between 3,700 m and 4,000 m is about 75 m. Drill 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 procedures conform to industry standard practices and have produced information suitable to support resource estimates. Sample recovery, to the extent documented, was acceptable. Sampling of core and RC cuttings was done in accordance with standard industry practices. 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.

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 honour original logs. Geological data and interpretations are suitable to support resource estimates.

1.9

Sample Preparation and Assaying

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.

1.10

Assay Quality Assurance and Quality Control (QA/QC)

Assay QA/QC protocols were observed throughout all drilling campaigns, with blind standard reference materials (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.

MRDI (1994) reviewed QA/QC results in detail for 1991 to 1994 and again (1997a) for core and RC holes drilled in 1995 and 1996. Overall, results indicated that sampling, preparation, and analytical procedures were adequate for obtaining reproducible (±20%) results for Au and Cu.

Smee and Associates (1997) evaluated QA/QC data for RC and core assays in the 1996 and 1997 drilling programs. SRM performance and assays of blanks, duplicate, and checks show acceptable analytical accuracy and precision.

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 0.1 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. Au grades above the 0.4 g/t internal cutoff are present in holes 99GT003, 99GT006 and CCD111. Coarse reject material should be reassayed for these holes prior to the next resource estimate update.

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.

1.11

Density

Bulk density values for ore and waste units are based on 877 measurements made on core samples in 1995 and 1996 core drilling campaign by EC Rowe and Associates (MRDI 1977a), 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.

1.12

Data Verification

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.

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.

1.13

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.

1.14

Metallurgical Processing

1.14.1

Introduction

The metallurgical evaluation of the ore characteristics by either of the current property owners was conducted over a period of three years (1997 to 1999), culminating with a series of batch flotation and cyanidation tests in late 1999. The outcome of this program was reported by G&T Metallurgical Services Ltd. of Kamloops (G&T) in January 2000.

The 2000 Feasibility Study was updated in March 2004 in order to reflect current costs for the capital and operating costs determination.

Plant design criteria relied mostly on the outcome of the grinding testwork completed in 1997, on the pilot plant work of June 1999 for dewatering circuits, and on the G&T flotation and cyanidation testwork of 1999. These trials focused on the treatment of various rock type composites.

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.

1.14.2

Metallurgical Tests

Metallurgical testwork appropriately categorized ore types on the basis of their 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. There was good agreement between the PDTS recovery models and actual locked cycle test results.

1.14.3

Throughput Capacity

The plant design criteria has been prepared for an average throughput of 150,000 t/d of ore when treating the hardest ore type, MDBX (MacPherson’s correlated autogenous work index of 18.4 kWh/t and Bond ball mill work index of 18.3 kWh/t). Average operating throughputs for all ore types are therefore 163,000 t/d, using a 92% plant availability. The oxides would be capable of a higher throughput still, at 180,000 t/d. This maximum throughput would not be achieved in the plant, and thus does not have to be accommodated for in the downstream processes of the plant, since the proportion of oxide in the mill feed stream is to be kept below 30% in order to achieve saleable concentrate copper grades.

A higher average throughput of 165,000 t/d was adopted for the first years (2 to 8) of operation and can be easily accommodated with the grinding circuit power base.

AMEC made cursory verifications of the major equipment sizing for crushing circuits to confirm the capacity of the chosen equipment to process the range of ores expected. Equipment checked was found to be sufficient for the required service.

All the flotation and regrinding circuit equipment selections involve proven technologies and are adequate for the proposed operation. Cleaner tail leaching circuit designs are adequate but may need the addition of a mercury retort system because of the mercury found in the flotation concentrate.

The water treatment system design is based on the Inco/SO₂ system, with SO₂ being provided by an elemental sulphur burner. This is a proven technology.

Use of small core for drop tests may have biased grinding requirements. AMEC supports the recommendation to conduct drop tests on larger core because if further testing demonstrates that SAG mill capacity is not a constraint; a substantial reduction in the capital cost could be realized by replacing three parallel lines with a dual SAG mill (each with three ball mills).

The calculation of the optimum grind target was comprehensive and is appropriate relative to present metal prices.

Design criteria are conservative relative to the interpretation of the available metallurgical data. Throughputs of softer ore blends may be greater than nameplate output of some of the equipment. Equipment sizing should consider peak throughputs that would result from processing softer ore blends. In particular, sufficient rougher retention time should be allocated for peak throughputs, as well as concentrate dewatering and leaching capabilities.

1.15

Mineral Resource and Mineral Reserve Estimates

1.15.1

Mineral Resource and Mineral Reserve Statements

Mineral Resources and Mineral Reserves are supported by appropriate exploration data, metallurgical tests, mine designs, and production plans that have been developed with generally accepted methods. Capital costs are supported with mine, processing, infrastructure, pipeline, port, water supply, and ancillary facilities designs at a feasibility study level. Operating costs are generally reasonable and supported by sufficient detail. Capital and operating costs do not reflect changes in costs since the March 2004 Feasibility Study Update. Place Dome intends on updating capital and operating costs in 2005 and, in addition, will re-evaluate technical aspects of the study dating from 2000.

Mineral Resources and Mineral Reserve estimates comply with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000) and Canadian National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators.

1.15.2

Mineral Resource Estimation Procedures

The mineral resource estimates in the 2000 Feasibility Study were calculated under the direction of Marc Jutras, P.Eng. of Placer Dome. The estimates, done in 1999, were made from 3-dimensional block models utilizing Placer Dome’s in-house mine planning software (OP). Cell size was 15 m east x 15 m north x 17 m high. Assays were composited into 2 m down-hole composites.

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. AMEC concurs with this philosophy for development of geologic models or domains for use in grade interpolation at Cerro Casale.

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.

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.

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 sparsely drilled regions yet multiple holes would be used in areas of denser 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 Placer Dome 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.

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). These values are supported by appropriate density measurements.

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-neighbour estimates on spatial swaths through the deposit. These checks showed no biases or local artefacts due to the estimation procedures.

1.15.3

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. AMEC would recommend that in future work there be direct verification that multiple holes located within the respective search ellipse will be used in estimating Measured and Indicated mineral resources rather than the 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 NI 43-101.

The mineral resources of the Cerro Casale project were tabulated based on material that fell within an optimistic ultimate pit shell. That shell was made based on a gold price of US$500/oz, US$1.25/lb for copper, and estimated costs (from experience on similar mines) of US$3.39/t. This shell fulfills the expectation of reasonable extraction test in declaring mineral resources at Cerro Casale. AMEC agrees with this logic and its implementation.

The mineral resources for Cerro Casale are reported at a gold cutoff grade of 0.4 g/t and are shown in Table 1-1. Mineral resources are entirely inclusive of mineral reserves.

Table -:

Cerro Casale Project Mineral Resource Summary – January 2000
(from PDTS, 2000)

Mineral Resource Category	Tonnage (Kt)	Grades		Contained Metal
Mineral Resource Category	Tonnage (Kt)	Gold (g/t)	Copper (%)	Gold (M oz)	Copper (Kt)
Measured	202,800	0.75	0.25	4.89	500
Indicated	910,800	0.70	0.26	20.53	2,368
Measured+Indicated	1,113,600	0.71	0.26	25.42	2,868
Inferred	171,200	0.63	0.33	3.49	565

Notes: 1. Mineral resources were defined within a Lerchs Grossman pit designed using $500/oz Au price, $1.25/lb Cu price and operating costs of $3.39/t. 2. Mineral resources are entirely inclusive of mineral reserves.

Mineral resources were last estimated in January 2000. These have not changed and are compliant with Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000) and Canadian National Instrument 43-101.

1.15.4

Mineral Reserves

Mineral reserves were estimated in the 2000 Feasibility Study using an elevated cutoff grade strategy for the four phases of mining and use of 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 based on a minimum profit value of $0.25/t after the deduction of all costs associated with mining, processing, administration, freight, smelting and refining from the recovered metal value. Blocks with values less than $0.25/t were classified as waste and assigned a negative value equivalent to the cost of mining that block. Metal prices used were $350/oz for gold and $0.95/lb for copper. Proven and Probable Mineral Reserves are listed in Table 1-2.

Life-of-mine waste:ore strip ratio is 2.67:1. Mining rates start at 296 Mt/a in Year 1, peak at 298 Mt/a in Year 2 and decline to 32 Mt/a by Year 18. Mine life is 18 years with two years of pre-production stripping.

Equipment selections are generally appropriate for the mine design, production rate and production schedule, but projected equipment availabilities are at the maximum for their respective ratings.

Table -:

Proven and Probable Mineral Reserves as of January 2000 using Elevated Profit Cutoff Strategy (from PDTS, 2000 Feasibility Study).

	Tonnage (Kt)	Grade		Contained Metal
	Tonnage (Kt)	Gold (g/t)	Gold (M oz)	Copper (Kt)	Copper (%)
Proven	200,266	0.72	4.64	501	0.25
Probable	834,799	0.68	18.25	2,170	0.26
Total	1,035,065	0.69	22.89	2,671	0.26

Notes: US$350/oz Au and US$0.95/lb Cu prices used. Metallurgical recovery equations are noted in Table 16-3 of this report.

Mine designs and production planning are suitable to support reserve estimates and are compliant with Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000) and Canadian National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators.

1.16

Mining Designs and Production Plans

Under the 2000 Feasibility Study all planned mining of the Cerro Casale deposit will be conducted by conventional open pit methods. The mine plan features a four stage open pit which is scheduled to deliver a nominal 150,000 t/d of mill feed at a maximum mining rate of 816,000 t/d over an 18 year mine life. The final pit stage will measure over 2,100 m in diameter and the highest sector of highwall will have a vertical extent of 1,224 m, ranking the proposed final pit amongst the world’s largest.

The primary crusher will be located 500 m south of the ultimate pit limit, although a 2 km haul road is required out of Stage 1 around the east side of Cerro Casale. Waste dumps and low-grade stockpiles will be located within 500 m of the pit entrance. The Río Nevado valley will be used to store waste rock. The northern edge of the waste rock dump will form the buttress for the tailing dam. The dumps and stockpiles will be built from the 4,087 m pit entrance elevation from the onset of mining. The truck fleet will be augmented by a trolley assist system in Year 5 as the majority of mining progresses below the pit entrance elevation of 4,087 m.

Mine design and planning steps were carried out using Placer Dome’s internally developed and proprietary ‘Pit Optimization Package’ (POP) software package, with exception of the stage limits which were defined using the NPV Scheduler package, and the smoothed pit stage design which was performed using the Vulcan mine planning package. AMEC considers these mine planning software packages to be robust, well accepted by the mining industry and appropriate for assessing the mining potential of the Cerro Casale deposit.

PDTS 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. An average gold price of $350/oz, a copper price of $0.95/lb and an exchange rate of US$1 per 510 Chilean pesos were used for the pit value model. Updates of capital and operating costs in 2004 used an exchange rate of US$1 per 600 Chilean pesos. AMEC reviewed the economic modelling methodology and parameters applied. They are considered to be standard practice 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 checked closely with the profit model values.

The ultimate pit was optimized in four stages with the net value maximized for each stage. Pit designs use a 17 bench height. AMEC recommends a standard 15 m bench height to match equipment selections.

The simplification of the inter-ramp slope recommendations used is reasonable for pit optimization and acceptable for feasibility level pit design. For the next feasibility update, AMEC recommends the construction of a geotechnical block model, which will allow the generation of pit designs that will more closely honour the geotechnical engineer’s inter-ramp angle, bench face angle and berm width recommendations.

Unbroken inter-ramp slopes in excess of 350 m vertical height should be avoided in preliminary mine plans.

1.17

Operating Costs

1.17.1

Total Unit Costs

Total unit operating costs, updated in March 2004 total $4.90/t of ore milled. Costs by area are shown in Table 1-3.

Table -:

Unit Operating Costs

Area	Cost $/t
Mine	1.64
Processing	2.63
Maintenance	0.12
Administration	0.51
Total	4.90

1.17.2

Mine Operating Costs

The open pit operating costs have been estimated on a yearly basis by determining major and support equipment requirements, including supplies, consumables, and manpower requirements. Cost information was derived from manufacturer’s information and extrapolated from existing Placer Dome operations.

1.17.3

Process Operating Costs

All the consumable costs reported in the final operating cost (OPEX) were revised in the March 2004 Feasibility Study Update (PDTS, 2004). The overall processing costs were revised from the 2000 Feasibility Study of $2.62/t to $2.63/t. Incremental costs for electricity and manpower were largely compensated by an expected reduction in the cost of consumables.

The resulting calculated processing costs distribution, as expected per the variations to major cost posts revealed in the 2004 feasibility update executive summary, are indicated in Table 1-4.

Table -:

Distribution of Processing Operating Costs

Area	Cost $/t
Operation labour	0.08
Maintenance labour	0.07
Reagents	0.31
Power	0.80
Grinding media	1.11
Operating supplies	0.02
Maintenance supplies	0.24
Total	2.63

1.17.4

Maintenance

Costs cover maintenance of mine and mill equipment, water supply, power transmission, substations, roads and yards, plant site buildings and accommodations camp. Labour costs are distributed into direct mine and mill operating costs.

1.17.5

General and Administration

G&A costs include personnel, accounting, warehousing, transport of employees, human resources, insurance, and head-office allocations.

AMEC reviewed costs in each area and found them to be reasonable for the time of estimate in the March 2004 Feasibility Study Update.

1.18

Capital Costs

Total capital costs are listed by facility in Table 1-5, as provided in Section 5 of the March 2004 Feasibility Study Update.

Table 1-6 shows post-construction sustaining capital costs total $287.9 million.

Direct costs include civil works in the pre-production stage, mine equipment, pre-production stripping, and construction of mine, process, camp, administration facilities, and general infrastructure. Methods used were adequate.

Quantities for civil works were estimated based on the general arrangement drawings developed for the project using historical unit prices available in Placer Dome’s database. The method used for the engineering development level is adequate.

Equipment for electrical works was estimated using Placer Dome’s database and local operating experience. The method used is adequate.

Selections and quantities of mining equipment are appropriate for the design of the project and considering the availability of equipment at the time of the 2000 Feasibility Study. PDTS intends to revise equipment selections in an update of the Feasibility Study in 2005.

Table -:

Total Estimated Capital Costs

	March 2004 US$ (millions)
Direct Costs
Plantsite & roads	38.1
Primary crusher	24.2
Coarse ore stockpile	17.4
Conveying	34.1
Grinding facilities	195.4
Flotation facilities	112.8
Leaching facilities	24.4
Refinery	5.8
Water supply	98.0
Shops & warehouses	20.2
First aid building	0.4
General office	6.4
Assay laboratory	3.5
Portside filter plant	15.6
Open pit – pre-production stripping & mining equipment	363.5
Power supply	27.1
Tailings disposal	31.7
Concentrate handling	9.3
Cyanide destruction	9.0
Concentrate pipeline	51.9
Accommodations	21.8
Vendors	7.0
Total Direct Costs	1,117.6
Indirect Costs
Construction overheads	83.3
Operations overheads	27.0
Project management	88.4
Design & engineering	75.5
Warehouse inventory	31.3
Freight	40.7
Taxes & duties	12.4
Total Indirect Costs	358.6
Total Construction Cost	1,476.3
Contingency	147.6
Commissioning	25.5
Total Capital Cost	1,649.4

Table -:

Post-Construction Sustaining Capital Costs
(from PDTS, 2004)

Area	Cost $ (millions)
Mining equipment	160.5
Pit dewatering & trolley assist	17.1
Tailings	22.9
Process plant & water system	71.4
Mine closure	16.0
Total	287.9

AMEC believes the following costs may be underestimated and should be revisited in the next feasibility study update:

•

a price of $1.50/m³ should be used for mass backfill, which may add $1 million to backfill in several sectors

•

an average price of concrete no lower than $650/m³ should be used, which may add $11 million to facility costs

•

mechanical platework should use a unit cost $3,500/t, for an increase of $3 million.

Indirect Costs

Indirect costs are those incurred in the construction and start-up phase and include all vendor, construction facilities and operation, freight, management fees, engineering procurement and construction management fees, spare parts, initial fills and consumables, and commissioning and owner costs.

In general, the bases for estimating indirect costs are in accordance with customary industry practices. AMEC recommends an additional $8 million for camp catering and maintenance based on other operations in Chile.

Sustaining Capital

Sustaining capital is required for mining and process equipment replacements, pit dewatering, installation of trolley assist for haul trucks and mine closure. Equipment selections and qualities appear reasonable relative to AMEC’s experience with similar scale projects.

1.19

Economic Analysis

Economic analysis of the Cerro Casale project is based upon a discounted cash flow analysis on a post-tax basis, using Proven and Probable Mineral Reserves and annual production plans as described in the 2000 Feasibility Study. 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 9.4 years. Life–of-mine is 18 years.

The model does not include an allocation for working capital, however, when standard estimates are used for working capital, there is only a small impact on return and the internal rate of return is still positive.

All other inputs are appropriate and, apart from the first few years of development, all future annual cash flows are 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 AMEC’s opinion, the level of detail used in the economic analysis is appropriate for a feasibility study and is a reasonable estimate of project value.

1.0

Project No.: 148562	Page #
June 2005

	ARIZONA STAR RESOURCE CORP.
	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

2.0

INTRODUCTION AND TERMS OF REFERENCE

2.1

Introduction

PDI commissioned AMEC to review resource and reserve estimation work done to date, for the Cerro Casale gold-copper deposit in northern Chile and prepare a Technical Report, as defined in National Instrument 43-101, Standards of Disclosure for Mineral Projects of the Canadian Securities Administrators. The format and content of this report are intended to conform to Form 43-101F1, Technical Report.

Larry B. Smith, P.Geo., an employee of AMEC, directed the review of resource and reserve estimates for the property. Mr. Smith reviewed the geology, exploration data, geological models, resource estimate procedures, and land position. Dr. Stephen D. Juras, P.Geo., an AMEC employee, reviewed resource estimation procedures and geostatistical support for estimates. Jay Melnyk, P.Eng., an employee of AMEC, reviewed mine designs, production plans, mineral reserve estimates and mine capital cost estimates. Pierre Lacombe, P.Eng., an employee of AMEC, reviewed metallurgical test data, process facility designs and process operating costs. Roberto Aburto and Manuel Romero, employees of AMEC International (Chile), reviewed infrastructure requirements, and general capital cost estimates. Graham Wood, an employee of AMEC, reviewed cash flow estimates and economic analyses. Lydia Letourneau, an employee of AMEC, reviewed environmental baseline studies, environmental management provisions, closure plans, and environmental permits. Dr. Ted Eggleston, P.Geo., an independent consulting geologist, reviewed historical assay quality assurance and quality control information for drilling campaigns in 1998 and 1999. Larry Smith reviewed historical assay quality assurance and quality control information for drilling campaigns from 1991 to 1997.

Information and data for the review and preparation of the report were obtained from Placer Dome offices in Vancouver, British Columbia; Santiago, Chile and Copiapo, Chile between 1 December 2004 and 20 January 2005. Some aspects of this report summarizing the geology of the deposit and historical protocols for drilling, sampling, sample preparation, assaying, drill hole surveying and density measurements were derived from a January 2000 Feasibility Study completed by PDTS Limited, from a Feasibility Study on oxide resources, a prefeasibility study on oxide and sulphide resources, and a scoping study on deep sulphide resources in 1997 by Mineral Resources Development, Inc. Data presented in these reports were verified by AMEC. Larry Smith, Jay Melnyk and Roberto Aburto visited the project site on 12 and 13 January 2005 and inspected drill core, drilling sites, outcrops of mineralization, general site plans, pipeline corridors and the proposed powerline route. Lydia Letourneau visited the site on 15 January 2005 and inspected environmental conditions, proposed tailings and waste rock disposal sites and monitoring wells.

AMEC’s opinions on resources, mine and process designs and reserves are based on the project as designed in the 2000 Feasibility Study. AMEC reviewed capital and operation costs updated in early 2004 by PDTS and provides its opinions of these capital costs as of that date. Increases in capital and operating costs occurred through the remainder of 2004. These increases will impact the project, but the degree to which cannot be determined until the technical and cost aspects of the project are again updated in 2005.

2.2

Terms of Reference

Unless stated otherwise, all quantities are in metric units and currencies are expressed in constant 2004 US dollars. This report is written for the entire project; the interests of any particular shareholder must therefore be deduced from the figures presented.

2.3

Units of Measure

2.3.1

Common Units

Above mean sea level

amsl

Ampere

Annum (year)

Billion years ago

British thermal unit

Btu

Candela

Carat

Carats per hundred tonnes

cpht

Carats per tonne

cpt

Centimetre

Cubic centimetre

cm³

Cubic feet per second

ft³/s or cfs

Cubic foot

ft³

Cubic inch

in³

Cubic metre

m³

Cubic yard

yd³

Day

Days per week

d/wk

Days per year (annum)

d/a

Dead weight tonnes

DWT

Decibel adjusted

dBa

Decibel

Degree

Degrees Celsius

°C

Degrees Fahrenheit

°F

Diameter

Dry metric ton

dmt

Foot

Gallon

gal

Gallons per minute (US)

gpm

Gigajoule

Gram

Grams per litre

g/L

Grams per tonne

g/t

Greater than

Hectare (10,000 m²)

Hertz

Horsepower

Hour

h (not hr)

Hours per day

h/d

Hours per week

h/wk

Hours per year

h/a

Inch

" (symbol, not ” )

Joule

Joules per kilowatt-hour

J/kWh

Kelvin

Kilo (thousand)

Kilocalorie

kcal

Kilogram

Kilograms per cubic metre

kg/m³

Kilograms per hour

kg/h

Kilograms per square metre

kg/m²

Kilojoule

Kilometre

Kilometres per hour

km/h

Kilonewton

Kilopascal

kPa

Kilovolt

Kilovolt-ampere

kVA

Kilovolts

Kilowatt

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

Litres per minute

L/m

Megabytes per second

Mb/s

Megapascal

MPa

Megavolt-ampere

MVA

Megawatt

Metre

Metres above sea level

masl

Metres per minute

m/min

Metres per second

m/s

Metric ton (tonne)

Micrometre (micron)

µm

Microsiemens (electrical)

µs

Miles per hour

mph

Milliamperes

Milligram

Milligrams per litre

mg/L

Millilitre

Millimetre

Million

Million tonnes

Minute (plane angle)

Minute (time)

min

Month

Newton

Newtons per metre

N/m

Ohm (electrical)

Ounce

Parts per billion

ppb

Parts per million

ppm

Pascal (newtons per square metre)

Pascals per second

Pa/s

Percent

Percent moisture (relative humidity)

% RH

Phase (electrical)

Pound(s)

Pounds per square inch

psi

Power factor

Quart

Revolutions per minute

rpm

Second (plane angle)

Second (time)

Short ton (2,000 lb)

Short ton (US)

Short tons per day (US)

tpd

Short tons per hour (US)

tph

Short tons per year (US)

tpy

Specific gravity

Square centimetre

cm²

Square foot

ft²

Square inch

in²

Square kilometre

km²

Square metre

m²

Thousand tonnes

Tonne (1,000 kg)

Tonnes per day

t/d

Tonnes per hour

t/h

Tonnes per year

t/a

Total dissolved solids

TDS

Total suspended solids

TSS

Volt

Week

Weight/weight

w/w

Wet metric ton

wmt

Yard

Year (annum)

2.3.2

Common Chemical Symbols

Aluminum	Al
Ammonia	NH₃
Antimony	Sb
Arsenic	As
Bismuth	Bi
Cadmium	Cd
Calcium	Ca
Calcium carbonate	CaCO₃
Calcium oxide	CaO
Calcium sulphide dehydrate	CaSO₄•2H₂O
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	MnO₂
Manganous hydroxide	Mn (OH)₂
Molybdenum	Mo
Nickel	Ni
Nitrogen	N
Nitrogen oxide compounds	Nox
Oxygen	O₂
Palladium	Pd
Platinum	Pt
Potassium	K
Silver	Ag
Sodium	Na
Sulphur	S
Tin	Sn
Titanium	Ti
Tungsten	W
Uranium	U
Zinc	Zn

	ARIZONA STAR RESOURCE CORP.
	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

3.0

DISCLAIMER

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. These reports indicate that all exploration and exploitation concessions, environmental permits and well field permits are secure and not under legal challenge.

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.

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.

4.0

PROPERTY DESCRIPTION AND LOCATION

4.1

Project Ownership and Agreements

Compañía Minera Casale (CMC) is a mining company incorporated under the laws of the Republic of Chile. Established under the Shareholders’ Agreement dated 5 June 2003, CMC is controlled 51% by Placer Aldebarán (Cayman) Limited, a member of the Placer Dome Shareholder Group and 49% by the Bema Shareholder Group. The Bema Shareholder Group is comprised of Arizona Star Resource (Bermuda) Ltd. (24.99% of shares of CMC) and BGO (Bermuda) Ltd. (24.01% of shares of CMC). 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.

The Shareholder Groups are defined under the agreement to mean:

•

Bema Gold Corporation (Bema), any of its Affiliates that may own interests from time to time, including Arizona Star Resource Corporation (Arizona Star) and its Affiliates, and any other Assignee of the Bema Shareholder Group (“the Bema Shareholder Group”); or

•

Placer Dome Inc. (PDI), any of its Affiliates that may own interests from time to time and any Assignee of the Placer Shareholder Group (“the Placer Shareholder Group”).

The shareholder groups are shareholders of CMC for the following purposes:

•

to explore and develop the Properties, construct facilities, and equip all or part thereof for commercial production

•

to profitably operate the Properties as mine over the long term

•

to market the doré, copper cathodes, copper concentrates (if any), and other products produced from the Properties

•

to engage in such other activities considered by the Participants to be necessary or desirable in connection with the foregoing.

In June 2003, an amendment to the 1998 Shareholders’ Agreement was signed by the parties. The amendment included terms for the payment of the holding costs for the period from the completion of the 2000 Feasibility Study to the commencement of construction, and terms for the repayment of the holding costs following commercial production.

In September 2004, a further amendment was signed by the parties to facilitate financing arrangements for the project and to cause the cash flows to be distributed consistent with the original intent of the Shareholders’ Agreement. The amendment included certain revisions to the financing structure whereby Placer would receive (1) an increased interest return on its shareholder loan funding to the project, and (2) a 3% NSR on revenues from the project. The amendments also provide for the option, at Placer Dome’s election, to fund the senior loans through Placer’s facilities. In return for compensation and as part of the amendments, Placer agreed to provide support, through the period of the pre-completion guarantee, for metal price hedge contracts required to secure the senior loans.

Except as provided in the Shareholders’ Agreement, all transactions, purchases, employment, operations, and negotiations with third parties will performed by the General Manager of CMC. Placer Dome Latin America is the current General Manager of CMC.

The Shareholders’ Agreement also requires PDI to accomplish the following or transfer the Placer Interest and Water Concessions to the Bema Shareholder Group for a nominal amount:

•

to cause PDLA to fund Maintenance and Permitting up to the US$900,000/a limit (plus applicable IVA charges)

•

deliver to the Bema Shareholder Group a quarterly certificate contemplated by Section 8.2 (a), (b) or (c) of the Shareholders’ Agreement

•

while a quarterly certificate contemplated by Section 8.2 (c) of the Shareholders’ Agreement is in effect, to use reasonable commercial efforts to arrange senior loans of an amount, which is at least 50% of the initial project capital requirements.

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 described in the Amended and Restated Shareholders’ Agreement (Table 4-1). Deposits with less exploration to date include Eva, Cerro Roman, Anfiteatro, Estrella, and Romancito Sur (Figure 4-1). CMC has carried out 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 -:

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.

Figure -:

Mineral Claims and Area of Interest, Aldebarán (from PDTS, 2000)

[amectechnicalreport43101014.jpg]

CMC-owned mining claims within the Area of Interest include 4,105 patented claims in 30 groups (Table 4-2), totalling 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 -:

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.

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 Chapter 21 as well as the opinion of Horst Altschwager and Flavio Fuentes of CMC. It has not been verified by AMEC and support documents (actual copies of water licenses or applications for licences) were not available for review.

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 7 applications for groundwater rights from 7 production wells. Groundwater rights for a total amount of 543 L/s have reportedly been granted to CMC but have not yet been permitted. 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^o193 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 have not yet been granted by the DGA.

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 depending 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

4.4

Other Costs

4.5

Environmental Exposures

4.5.1

Introduction

Based on AMEC’s review of the project, 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 are:

Environmental Approval of Power Supply Infrastructure

Environmental Approval of Port Facilities

Acid Rock Drainage (ARD) potential

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

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

4.5.2

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.

4.5.3

Environmental Approval of Port Facilities

The Cerro Casale Project proposes to use existing port facilities currently operated by Compañía Minera 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 Compañía Minera 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 Compañía Minera Candelaria. It is reasonable to assume that a contract will be negotiated and that approval for port modifications will be obtained.

4.5.4

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 Acid Rock Drainage (ARD)/leach metal or even produce any drainage”. The memorandum further stipulates that “any drainage from the Aldebarán (previous name of the current to 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.

AMEC recommends that further evaluation of ARD potential be carried out 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.

4.5.5

Impacts on Surrounding Water Systems from Water Take 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 neighbouring 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.

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. Amongst 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.

•

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.

5.0

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

5.1

Location, Access, Local Resources and Infrastructure

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 25 km is paved highway south from Copiapo, connecting with a 155 km gravel road southeast to the project site. Currently, total driving time from Copiapo to site is approximately 3½ hours.

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, and mine facilities, as well as mill, concentrator, tailings and waste rock disposal. Figures 5-2 and 5-3 show the mill and tailings/waste rock disposal sites, respectively.

Surface rights are held by the national government, which normally provides surface use permits for mining operations as a priority use.

Access to the property is via 180 km of paved and dirt regional roads from Copiapo. 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 skilled labour force is available in the Copiapo region and surrounding mining areas of northern Chile. A source of electric power must be negotiated. Suitable water supply is available from the presently permitted Piedra Pomez well field, located 121 km north of the project.

Figure -:

Location of Cerro Casale Project, Northern Chile

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Figure -:

Mill Site

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Figure -:

Tailings and Waste Rock Site

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5.2

Climate, Physiography, Vegetation, and Fauna

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.

The climate is typical for the northern Chilean Andes. Precipitation is generally limited to snowfall in April through September and rain is rare. Daytime temperatures in summer months get up to 23°C, with nighttime lows of 5°C. Daytime temperature in winter is around freezing, with nighttime temperatures dropping to -15°C.

5.3

Environmental Impact Assessment

Environmental baseline studies of the Cerro Casale Project were initiated in 1997 and continued monitoring of selected components (meteorology, hydrology, hydrogeology and water quality) is on going since then. Baseline work was conducted by a number of specialists identified as authors of the various baseline reports referenced in Chapter 21. SENES Chile S.A. was contracted to prepare the project’s EIS document presented to COREMA on March 12, 2001. Completion of the EIS was the result of joint efforts by all parties involved with support studies including but not limited to geology, engineering, design, evaluation, and baseline reports. The content of the EIS included baseline information on hydrology, hydrogeology, archaeology, biodiversity, fauna and flora, socio-economics, air quality, meteorology, climatology, geology, geomorphology, land use, and acid rock drainage. Also included is information aimed at fulfilling requirements outlined in applicable legislation and regulations as specified in Law 19.300 and D.S. 95 from 2001 respectively.

Information from relevant EIS sections is included below.

5.3.1

Archaeological Resources

Archaeological investigations were conducted as part of baseline environmental studies for the Cerro Casale Project. These investigations aimed at documenting the presence of archaeological sites and developing the necessary management plans to protect these resources from proposed project activities and infrastructure. Miguel Cervellino Giannoni lead studies developed for both the Cerro Casale area and the port site at Punta Padrones. Results are discussed in reports presented to CMC in July 1999 and November 2000 respectively.

A total of 69 sites were registered along the pipeline routes and in the Cerro Casale area. None were found in the camp area.

Field records include a description of resources found in the study area. These indicate that the study area has been subject to a sequence of human settlements starting at least in the Arcaic Period of northern Chile’s pre-history (about 6,000 to 4,000 years BC), and extend into the present.

Sites along the Water Pipeline Corridor

A total of four sites were registered along the water pipeline corridor that extends from Piedra Pomez to Cerro Casale. Two sites were classified as “Less Important” and two as “Important” showing a Fairly Good conservation pattern. Three sites correspond to archaeological sites and one site corresponds to an archaeological – anthropological site (re-occupation). The sites are villages or archaeological camps.

Cerro Casale Area

A total of nine sites were registered in the Cerro Casale area. Eight were classified as “Important” and one as “Very Important” as it shows a Fairly Good conservation pattern. All the sites correspond to archaeological sites of villages or camps, simple houses or tambos.

Sites along the Concentrate Pipeline Corridor

A total of 46 sites were registered along the concentrate pipeline corridor that extends from Cerro Casale to Punta Padrones. Eighteen sites were classified as “Less Important,” 19 as “Important” and nine as “Very Important”. The conservation pattern varies from Fairly Good, Good, Seriously Bad, to Fairly Bad. All the sites correspond to archaeological, historical, and anthropological sites as well as tambos, camps, barnyards, kilns, apacheta, and villages.

Area of Port Facilities

During this investigation, sites previously found in a survey conducted in 1983 were reclassified. Using the current scenario, 10 sites were registered; two sites were classified as “Less Important” and eight sites as “Important.” These sites were classified as archaeological and mainly consisted of shell accumulation areas with houses/graves. An evaluation of their state of conservation revealed that most have a Seriously Bad (3 sites) to Fairly Good (7 sites) conservation pattern.

5.3.2

Fauna

Fauna of the Cerro Casale Project area was also investigated as part of baseline environmental studies carried out by biologists Gustavo Mieres Urquieta and Juan Carlos Torres Mura. The study included three field campaigns designed to identify and characterize fauna present in the project area which was, as in the case of vegetation, sub-divided into three segments for study purpose, namely: the water pipeline corridor extending from Piedra Pomez to Cerro Casale, the Cerro Casale area and the concentrate pipeline corridor that extends from Cerro Casale to the port site in Punta Padrone. Results of these investigations are documented in a study report presented to CMC in September 1999.

Overall Results

The Cerro Casale Project area is only occupied by a low number of species characterized by a large habitat selection area. Species present are also widely distributed throughout many regions of the Andes.

Vertebrates are considered scarce in all three study areas. Their population density is low; however, fauna tend to concentrate in certain areas, specifically, close to water, thereby converting wetlands into the most sensitive areas.

All recorded species have a large distribution throughout the country with the exception of the reptiles Liolaemus juanortizi and Liolaemus rosenmanni, restricted to the Third region and the rodent Ctenomys fulvus, found in the Second and Third Region.

Water Pipeline Corridor from Piedra Pomez to Cerro Casale

Vertebrates recorded along the water pipeline corridor are composed of 39 species that include 3 reptiles, 28 bird and 8 mammals, all of which are native except for 1 mammal.

Nine species are cited in conservation lists. The reptile Liolaemus juanortizi, the two camelidae, Lama Guanicoe and Vicugna vicugna, and the Lagidium viscacia are listed as endangered. The reptile L. nigriceps and the mammal Ctenomys fulvus are considered vulnerable. The reptile L. roenmanni and the bird Attagis gayi are designated as rare; and the mammal Pseudalopex culpaeus is inadequately known.

Both the camelidae and the fox (Pseudalopex culpaeus) use large areas to secure food supplies; however, they tend to concentrate in more vegetated and better protected areas. The bird (Attagis gayi or perdicita) normally lives in wetlands and is not present in high density. The rest of the species are highly dispersed throughout the area.

Cerro Casale Area

The fauna of the Cerro Casale area is comprised of 45 species including 3 reptiles, 34 birds and 8 mammals. All species are native except for one mammal.

Ten species are cited in conservation lists. The reptile Liolaemus juanortizi, the two camelidae, Lama Guanicoe and Vicugna vicugna, and the Lagidium viscacia are all classified as endangered. The reptile L. nigriceps, the bird Chloephaga melanoptera and the mammal Ctenomys fulvus are considered as vulnerable. The reptile L. roenmanni and the bird Attagis gayi are listed as rare; and the mammal Pseudalopex culpaeus is reported as inadequately known.

Concentrate Pipeline Corridor from Cerro Casale to Punta Padrone

Ninety species of vertebrates were observed along the concentrate pipeline corridor. These included 2 amphibians, 11 reptiles, 61 birds, and 16 mammals. All observed species are native except for two bird and one mammal species.

Twenty-eight species are included in conservation lists. The two amphibians and the reptile Liolaemus juanortizi are classified as endangered, five reptiles are designated as vulnerable, four reptiles are considered rare and only one is not endangered. One bird is listed as endangered, three are considered as vulnerable and one classified as rare. Five mammals are endangered, one is vulnerable, and four are inadequately known.

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 Gold Corporation (Bema) and Arizona Star Resource Corporation (Arizona Star), both being members of 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. Present project equity is split between 51% to the Placer Shareholder Group and 49% to the Bema Shareholder Group.

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.

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 centres 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 centres 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).

Figure -:

Geology of the Maricunga Volcanic Belt (from PDTS, 2000)

[amectechnicalreport43101024.jpg]

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 down dip steeply to the southwest. The highest-grade mineralization is coincident with well developed quartz-sulphide stockworks in strongly potassic-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 totalling 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.

Table -:

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 -:

Surface Geological Map of Cerro Casale (from PDTS, 2000)

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Figure -:

Cross Section 850E, Looking Northwest, Cerro Casale Deposit (from 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.

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

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

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

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 -:

Redox Units, Section 850E (from 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 focussed 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.

1.0


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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 may be 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 -:

Major Gold-Copper Occurrences in the Aldebarán Property (from 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 centre 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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	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

Figure -:

Measured + Indicated Gold Resources, Section 472200E (from PDTS, 2000)

[amectechnicalreport43101064.jpg]

Figure -:

Measured + Indicated Copper Resources, Section 472200E (from PDTS, 2000)

[amectechnicalreport43101066.jpg]

Figure -:

Measured + Indicated Gold Resources, 3832 Elevation (from PDTS, 2000)

[amectechnicalreport43101068.jpg]

Figure -:

Measured + Indicated Copper Resources, 3832 Elevation (from PDTS, 2000)

[amectechnicalreport43101070.jpg]

Figure -:

Intensity of Stockwork Veining, Section 850E (from PDTS, 2000)

[amectechnicalreport43101072.jpg]

Figure -:

Potassium Feldspar Alteration, Section 850E (from PDTS, 2000)

[amectechnicalreport43101074.jpg]


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	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

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 (Cu₃S) 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 sporatic 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 (Figure 9-8).

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° (Figure 9-8) and dip approximately 70° south (Figure 9-9).

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 centred 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.

Figure -:

Geological Map of the Eva Deposit (from PDTS, 2000)

[amectechnicalreport43101080.jpg]


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	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

Figure -:

Cross Section of Eva Deposit (from PDTS, 2000)

[amectechnicalreport43101086.jpg]


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	CERRO CASALE PROJECT, CHILE
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Figure -:

Geological Map of the Cerro Roman Deposit (from PDTS, 2000)

[amectechnicalreport43101092.jpg]

Figure -:

North-South Cross Section of Cerro Roman Deposit (from PDTS, 2000)

[amectechnicalreport43101094.jpg]

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%.

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 suggest that 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°.

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.

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.


Project No.: 148562	Page #
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10.0

EXPLORATION

10.1

Introduction

Between the late 1980s and 1999, the Aldebarán area containing Cerro Casale was explored by Anglo American, Bema Gold Corporation, Arizona Star Resource Corporation, 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.

10.2

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 totalling 43,317 m and 88 core holes totalling 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 totalling 23,924 m.

10.3

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 totalling 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.

10.4

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, carried out surface magnetic and Induced Polarization surveys and drilled 41 RC holes totalling 7,250 m.

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

10.5

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 totalling 3,378 m. Placer Dome remapped the area in 1998 and trenched obvious areas of alteration. In 1999 Placer Dome drilled four core holes totalling 1,225 m.

10.6

Anfiteatro

Anglo American carried out 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 totalling 536 m. Placer Dome drilled three core holes totalling 998 m in 1990.

10.7

Romancito

Limited exploration work has been completed at Romancito. Regional mapping carried out 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 carried out. Two core holes totalling 794 m were drilled.

10.8

Other Areas

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

11.0

DRILLING

11.1

Introduction

Reverse-circulation (RC) and core drilling was carried out 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 -:

Cerro Casale Drilling

Year	Company	Type	Purpose	Holes	Metres
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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	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

Figure -:

Drill Collar Locations (from PDTS, 2000)

[amectechnicalreport43101102.jpg]


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	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

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.

1.1


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	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

Figure -:

Average and Median Drill Spacing by Elevation (from PDTS, 2000)

[amectechnicalreport43101112.jpg]


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	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

11.2

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.

11.2.1

Reverse Circulation Drilling

RC drilling in 1991 was carried out 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 carried out 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 precollar intervals for deeper core holes. The RC portions of these holes were sampled and assayed where mineralized.

11.2.2

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 totalling 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.

Three holes totalling 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.

11.3

Geological Logging Practices

11.3.1

Introduction

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.

11.3.2

Reverse Circulation Chip Logging

CMA geologists logged cuttings from each 2 m interval at the drill site using a hand lens. Colour, 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.

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 to 3%

•

3% to 7%

•

7% to 10%

•

>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.

11.3.3

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.

11.3.4

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.

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.

11.4

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 about 50 boxes of older core in a separate storage facility to inspect the general condition of the core and core recovery. Rock quality was found to be generally high with few intervals of strongly fractured rock and poor core recovery.

11.5

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.

11.6

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.

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 km² area around the pit area. Quoted vertical and horizontal accuracy is 2 m (MRDI, 1997a).

11.7

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 metre 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 -:

Drill Hole Collar Monuments

[amectechnicalreport43101118.jpg]

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.

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.

11.8

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.

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 locations 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.


Project No.: 148562	Page #
June 2005

	ARIZONA STAR
	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

12.0

SAMPLING METHOD AND APPROACH

12.1

Introduction

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.

12.2

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 metre 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% of samples were collected wet. Weight recovery was not measured for wet samples.

All collection, splitting and bagging of samples was carried out by CMA personnel.

12.3

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 labelling.

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.

12.4

List of Significant Assays

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

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.

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.

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.

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’s 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.

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 -:

Relative Differences for Rig Duplicates (from MRDI, 1997b)

[amectechnicalreport43101122.jpg]

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.

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.

Figure -:

Checks of Acme Gold Assays by Chemex (from MRDI, 1997b)

[amectechnicalreport43101124.jpg]

Note: X axis is g/t Au for Acme and Y axis is g/t Au for Chemex

Figure -:

Precision from Chemex Check Assays of Acme Gold Assays
(from MRDI, 1997b)

[amectechnicalreport43101126.jpg]

Note: X axis is percentile and Y axis is relative difference between analyses

Figure -:

Chemex Check Assays of Acme Copper Assays
(from MRDI, 1997b)

[amectechnicalreport43101128.jpg]

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.

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 -:

Check Assays by Chemex, 1996 and 1997 (from MRDI, 1997b)

	DDH Duplicates		RC Duplicates		DDH Duplicates		RC Duplicates
	Gold (g/t) Acme	Gold (g/t) Chemex	Gold (g/t) Acme	Gold (g/t) Chemex	Copper (%) Acme	Copper (%) Chemex	Copper (%) Acme	Copper (%) Chemex
	0.519	0.509	0.531	0.492	0.202	0.212	0.080	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 -:

Acme and Chemex Analyses of Standard, 1996-1997 (from 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

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.

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 mislabelled standard STD18.

Table -:

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 -:

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 -:

1998 Cerro Casale Standard (Blank) STD05 – Gold

[amectechnicalreport43101130.jpg]

Figure -:

1998 Cerro Casale Standard (Blank) STD05 – Copper

[amectechnicalreport43101132.jpg]

Figure -:

1998 Cerro Casale Standard STD12 – Gold

[amectechnicalreport43101134.jpg]

Figure -:

1998 Cerro Casale Standard STD12 – Copper

[amectechnicalreport43101136.jpg]

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.

Figure -:

1998 Cerro Casale Standard STD13 – Gold

[amectechnicalreport43101138.jpg]

Figure -:

1998 Cerro Casale Standard STD13 – Copper

[amectechnicalreport43101140.jpg]

Figure -:

1998 Cerro Casale Standard STD14 - Gold

[amectechnicalreport43101142.jpg]

Figure -:

1998 Cerro Casale Standard STD14 – Copper

[amectechnicalreport43101144.jpg]

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 -:

1998 Cerro Casale Standard STD18 - Gold

[amectechnicalreport43101146.jpg]

Figure -:

1998 Cerro Casale Standard STD18 – Copper

[amectechnicalreport43101148.jpg]

Figure -:

1998 Cerro Casale Standard (Blank) STD19 – Gold

[amectechnicalreport43101150.jpg]

Figure -:

1998 Cerro Casale Standard (Blank) STD19 – Copper

[amectechnicalreport43101152.jpg]

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.

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 -:

1998 Cerro Casale Gold Duplicate Data

[amectechnicalreport43101154.jpg]

Figure -:

1998 Cerro Casale Gold Duplicate Data

[amectechnicalreport43101156.jpg]

Figure -:

1998 Cerro Casale Copper Duplicate Data

[amectechnicalreport43101158.jpg]

Figure -:

1998 Cerro Casale Copper Duplicate Data

[amectechnicalreport43101160.jpg]

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 II.l.l1 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 -:

1998 Check Assay Statistics

Sample (8=471)	Mean	St. Dev	Max	75^th Percentile	Median	25^th 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

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 mislabelled standards and one is unexplained, but which was probably a mislabelled 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.

Figure -:

1999 Cerro Casale Standard STD12 – Gold

[amectechnicalreport43101162.jpg]

Figure -:

1999 Cerro Casale Standard STD12 – Copper

[amectechnicalreport43101164.jpg]

Figure -:

1999 Cerro Casale Standard STD13 – Gold

[amectechnicalreport43101166.jpg]

Figure -:

1999 Cerro Casale Standard STD13 – Copper

[amectechnicalreport43101168.jpg]

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 reassayed. 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 behaviour 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.

Figure -:

1999 Cerro Casale Standard STD14 – Gold

[amectechnicalreport43101170.jpg]

Figure -:

1999 Cerro Casale Standard STD14 – Copper

[amectechnicalreport43101172.jpg]

Figure -:

1999 Cerro Casale Standard STD18 – Gold

[amectechnicalreport43101174.jpg]

Figure -:

1999 Cerro Casale Standard STD18 – Copper

[amectechnicalreport43101176.jpg]

Figure -:

1999 Cerro Casale Standard (Blank) STD19 – Gold

[amectechnicalreport43101178.jpg]

Figure -:

1999 Cerro Casale Standard (Blank) STD19 – Copper

[amectechnicalreport43101180.jpg]

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 -:

1999 Cerro Casale Gold Duplicate Data

[amectechnicalreport43101182.jpg]

Figure -:

1999 Cerro Casale Gold Precision Estimate

[amectechnicalreport43101184.jpg]

Figure -:

1999 Cerro Casale Precision Estimate by Data Date

[amectechnicalreport43101186.jpg]

Figure -:

1999 Cerro Casale Duplicate Copper Data

[amectechnicalreport43101188.jpg]

Figure -:

1999 Cerro Casale Copper Precision Estimate

[amectechnicalreport43101190.jpg]

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 -:

1999 Check Assay Statistics

Sample (8=359)	Mean	Standard Deviation	Max	75^th Percentile	Median	25^th 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 carried out 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.

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 -:

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 -:

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.

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 -:

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 -:

Boxplot of All Density Measurements by Oxidation Categories

[amectechnicalreport43101192.jpg]

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 carried out with protocols conforming to industry standard practices. AMEC agrees with the assignment of densities by oxidation domain.

14.0

DATA VERIFICATION

14.1

Database Development and Integrity Checks

14.1.1

Introduction

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 directly 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).

14.1.2

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.

14.1.3

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.

14.1.4

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 this 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).

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 artefacts 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

15.0

ADJACENT PROPERTIES

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

16.0

MINERAL PROCESSING AND METALLURGICAL TESTING

16.1

Introduction

Metallurgical characteristics of Cerro Casale mineralization is based on a variety of information summarized in Volume 3 of the 2000 Feasibility Report (PDTS, 2000) and its accompanying appendices, which also outlines the process plant design and its planned operation. Additional information consists of:

•

Volume 1 of the Feasibility Report, dealing with the geology of the deposit.

•

Documents and data provided by PDTS (PDTS), covering the results from the most recent program of locked cycle flotation and cyanidation trials, as performed at G&T Metallurgical Services (G&T) at the end of 1999.

•

Histograms of the grade distributions for copper and gold in the block model produced by Placer Dome Exploration (PDX).

•

Flowsheets and mass balances produced with the plant design criteria.

AMEC acquired additional information from PDTS metallurgist, John Folinsbee, who contributed to the preparation of the feasibility study and supervised the latest portions of the metallurgical testwork programs.

Relevant data originating from individual testwork campaigns are found within the body of the Feasibility Report, or provided separately by Placer Dome. These were from the following documents:

•

A report produced by Hazen Research (Cerro Casale Metallurgical Study), commissioned by Bema Gold Corp., in November 1997.

•

Reports from G&T, A.R. MacPherson Consultants (MacPherson) and McClelland Laboratories, commissioned by Bema Gold Corp.

•

Three reports issued by the Placer Dome Research Centre (PDRC), documenting respectively the mapping, bench scale and pilot plant testwork programs realized by them.

•

A G&T report issued in early 1999 (A program of Flotation and Modal Studies – Cerro Casale, Maricunga District, Region III, Chile).

•

MacPherson grinding testwork report dated 4 December 1998 (Grindability Studies on Ore Samples from the Alderabán/Cerro Casale Property).

•

A report by Contract Support Services (CSS), on grinding simulations realized with the JKSimMet software, produced in January 1999 (Report to Placer Dome – Results of Phase 1 Simulation Study – Expected Performance of Proposed Alderabán Comminution Circuit).

•

From the same source as above, the report on a second phase of simulation work issued in May 1999.

•

A December 1999 updated analysis of the proposed grinding circuit from CSS, produced in light of the modifications brought during the feasibility study finalization.

The metallurgical evaluation of the ore characteristics was conducted over a period of three years (1997 to 1999), culminating with a series of batch flotation and cyanidation tests in late 1999. The outcome of this program was reported by G&T in January 2000.

Capital and operating costs were updated in March 2004 (PDTS, 2004).

At a late stage in the metallurgical testwork phase, the decision was taken to target a flotation feed size distribution typified by a P₈₀ of 120 µm. The plant design criteria relied mostly on the outcome of the grinding testwork completed in 1997, of the pilot plant work of June 1999 for dewatering circuits, and on the G&T flotation and cyanidation testwork of 1999. These trials focused on the treatment of various rock type composites.

AMEC reviewed metal recoveries, processing costs and smelter terms in regards to the calculation of net values for each resource block.

16.2

Metallurgical Testwork

16.2.1

Lithological Considerations

Placer Dome Exploration (PDX) geologists divided the Cerro Casale resources into ten lithological units, or rock types, for modelling purposes. Table 16-1 presents the rock types identified, reported mineable tonnage contained in the limit of the smoothed pit (as presented in the Feasibility Study Report, for a $0.25/t profit cutoff criteria), % of this value represented by each rock type and the average head grades for the blocks contained within the smoothed pit outline.

Table -:

Mineable Reserves within Smoothed Pit (@0.25$/t profit cutoff)

Fine-Grained Volcanic

Microdiorite Breccia

Rock Type	Designator	Tonnage		Average Grades
Rock Type	Designator	(Mt)	(%)	(%CuT)	(g/t Au)
Diorite Sulphide Upper	DSU	142.9	13.8	0.23	0.56
Diorite Sulphide Lower	DSL	73.5	7.1	0.30	0.73
Granodiorite Sulphide	GS	197.5	19.1	0.31	0.78
Volcanic Breccia	VB	383.8	37.1	0.25	0.64
Mafic Flow	MVF
	VPF
	MDBX	105.9	10.2	0.34	0.81
All Oxide	AO	102.6	9.9	0.12	0.71
Catalina Breccia	CBX	3.2	0.3	1.15	4.11
Undefined	UD	25.7	2.5	0.22	0.48
Total / Avg.	-	1,035.1	100.0	0.26	0.69

The undefined category may contain various mineralized dykes and breccias of individual volumes deemed too minor to be tracked individually. The VB, MVF, and VPF rock types evolved as separate entities from an earlier unit called Volcanic Sulphide (VS) that was encompassing them all. VB is reportedly making up 68% of the overall VS content, with a further 18% coming from VPF and 14% from MVF.

Per Table 16-1 data, five ore types are encompassing the bulk of the overall reserves: DSU, GS, VS, MDBX, and AO are inclusive of 90.1% of the total. Characterization of the metallurgical behaviour of these five ore types, with respect to establishing the plant design criteria is sufficient. This characterization assumes that the mine plan does not anticipate feeding the mill high proportions of one of the less-abundant ore types appearing in Table 16-1.

The material was further classified into three categories, according to its degree of weathering. Material exhibiting a loss of less than 10% of its original sulphide content is classified as a sulphide (DSL, DSU, VS, GS). 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).

Breccias are characterized by higher grades than the other rock types and are usually found as trending parallel the granodiorite, above and below it. As an exception, the Catalina Breccia is instead spatially associated with the microdiorite breccia. In addition to the higher concentration of copper and gold found in it, higher contents of zinc and lead can also be expected. It is reported that 2 m core intervals grading up to 13% Cu and 200 g/t Au were encountered in the Catalina Breccia.

16.2.2

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 were, in decreasing order: quartz, feldspar, mica, chlorite, gypsum, pyrite, chalcopyrite, and bornite.

Copper and gold are strongly correlated. Copper is found mainly in chalcopyrite but bornite and, to a lesser extent, chalcocite, digenite, covellite, chrysocolla and malachite are also occurring occasionally. Bornite to chalcopyrite ratio is found to increase at depth, with very little bornite above the 4000 m elevation. The average copper grade increases approximately 25% below the 4000 m elevation.

Chalcocite, covellite, chrysocolla, and malachite are found at the oxide/sulphide boundary.

Pyrite:chalcopyrite ratios are likely to play a role with respect to the achievable final concentrate copper grade produced (as perceived from some Placer Dome tests). The upper portions of the orebody, where the higher copper-grading bornite mineral is less abundant, should represent the most challenging ore at high pyrite to chalcopyrite ratios.

Gold mineralogy was not defined by any specific mineralogical work. 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 present at an average 3:1 ratio to gold, suggesting that the gold is present as electrum. Electrum may be present in fine grains (as seen by modal analysis of flotation products), within the fluid matrix that deposited the copper minerals, and/or as solid inclusions within these minerals. Modal analysis demonstrated (except for GS) that up to 85% of the gold content is associated specifically with chalcopyrite and less than 1% is found with pyrite. However, petrographic examination of high-grade CBX material showed that 85% of the gold was present as exposed grains along pyrite grain boundaries. An average gold grain size of 39 µm, within a range of 7 µm to 145 µm, was obtained from the analysis of a sample of CBX with 31 grains observed.

16.2.3

Comminution

Two phases of comminution parameter determinations were completed. Each may have included a number of the following: impact, SAG, rod and ball mill grindability, or abrasion index measurements. The first testwork phase was completed for Bema Gold Corp., through MacPherson, McClelland, and Hazen, while the second one was commissioned by Placer Dome in 1998 and realized by MacPherson and Hazen. Results from the drop weight tests conducted by Hazen were then interpreted by CSS for simulation work realized with the JKSimMet software in 1999.

An overview of the results obtained is presented in Table 16-2, where the following abbreviations are used:

•

BGC:

denotes samples prepared by Bema Gold Corp.

•

PDI:

denotes samples prepared by Placer Dome.

•

Ai:

abrasion index (paddle test, Allis Chalmers methodology)

•

CWi:

crushing (or impact) work index (Bond methodology)

•

RWi:

rod mill work index (Bond methodology) @ target product P80 of 1,190 µm (PDI determinations, not mentioned for BGCs)

•

CAWi:

correlated autogenous work index, brought at indicated P₈₀ (MacPherson methodology)

•

A, b, ta:

JKTech’s drop weight test parameters, where A*b is indicative of the ore resistance to impact and t_a is a measure of its resistance to grinding by attrition (higher values indicating incremental resistance)

•

BMWi:

ball mill work index (Bond methodology) @ target product P₈₀ of 147 µm (100 Mesh)

Table -:

Overview of Measured Comminution Parameters

Rock Type	Ai	CWi (kWh/t)	RWi (kWh/t)	CAWi@P₈₀(kWh/t@µm)	*Ab**	t_a	BMWi (kWh/t)	SG
MDBX (BGC-Comp. A)	0.422	6.0	17.6	18.4 @ 202			18.8	2.66
MDBX (PDI)	0.357	9.2	18.8	18.5 @ 144	27.3	0.38	16.5	2.70
DP Sulph (=DSL+DSU) (BGC-Comp. B)	0.263	6.9	18.9	20.9 @ 171			18.0	2.71
DP Sulph (=DSL+DSU) (PDI)	0.285	10.4	19.3	16.5 @ 198	33.2	0.38	16.9	2.75
DP Ox (=AO) (BGC-Comp. C)	0.109	11.8	15.1	14.4 @ 329			15.2	2.77
DP Mixed (=DSL+DSU+AO) (BGC-Comp. D)	0.230	13.1	17.6				17.1
DP Enrich (=supergene?) (BGV-Comp. E)	0.064	4.4	15.5				16.9
GS (PDI)	0.433	14.6	18.8	18.1 @ 158	36.4	0.47	16.5	2.75
VB (= subset of VS) (PDI)	0.335	10.5	19.3	17.0 @ 200	28.0	0.43	16.7	2.76
Weighted Avg.¹	0.323	10.5	19.1	17.0	31.0	0.41	16.7	2.75

Note: Calculations using only Placer Dome determinations (except for AO, from BGC) and weighing per percent of pit envelope tonnage indicated in Table 16-1, with VB values ascribed to all of VS.

The location of BGC samples is apparently unknown, lest for the fact that drill core material was used. The percent BGC samples were therefore not included in the calculation of averages since their representativity, and whether they would be within the pit envelope developed later, is unknown. The results for oxides were included in order to achieve a higher resources coverage, adding the 9.9% of the resources represented by this rock type to the total above.

Placer Dome samples used for the composite preparation were obtained from drill core material originating from 14 holes, with five of them contributing to the preparation of the MDBX composite; five and eight of DSL and DSU material, respectively, to create the DS sulphide composite; six for VB (assimilated into VS); and two holes for GS.

The types of tests conducted are appropriate for sizing the relevant comminution equipment. The product sizes selected for the BMWi determinations are adequate for the flotation feed size selected.

Ore types with parameters indicated in Table 16-2 represent 97.2% of the metallurgical ore types in the mineral reserve. JKTech parameters (A*b and ta) are covered by a lower percentage of these resources, at 87.3%.

It is not entirely appropriate to assume that the results obtained from a VB sample are characteristic of all of the VS material, while the two other components of this rock type classification (MVF and VPF) were excluded and represent approximately 32% of all material classified as VS. It is likely, for instance, that the VPF subset is a harder material than the VB, with its fine-grained nature.

AMEC is of the opinion that the distinction between VB and VS designations should be maintained when reporting results obtained for VB material only. Since VB comprises over a quarter of the pit envelope resources (68% of the 37.1% covered by VS), an appropriate distinction of the whole and its sole tested constituent is warranted. Future VS composites prepared for grinding parameter determinations should either maintain these rock types separately or properly respect their proportions in the final composite make-up. Considering the above, the actual coverage provided by the grinding determinations completed is reduced somewhat, but still sufficient for design purpose.

Future work should include more samples of BGC.

MacPherson concluded that critical size build-up was likely to occur in the SAG mill, from the RW_i frequently higher than the BW_i, indicative of a higher resistance to impact than abrasion. This characteristic dictated the inclusion of a pebble circuit, with crushing, within the SAG mill circuit (SABC configuration). AMEC concurs with this conclusion.

The latest round of grinding simulations, realized by Contract Support Services (CSS) in December 1999 with the JKSimMet software, accounted for a revised target grind finer than the one implied by the earlier CSS work. The result of this study was to increase the recommended diameter of the ball mills selected previously and validate the breakage characteristics expected from the largest operating SAG mill diameter retained (12.19 m diameter), of which only one to date is in operation (Cadia, Australia).

A recommendation was issued by CSS to Placer Dome at that time, requesting the performance of a new set of JKTech grinding parameter determinations to validate the simulations up to 1999. Previous tests were completed on half-core sections, which is limiting the top size of particles made available for the testwork procedures. This, in turn, is reducing the simulator’s capabilities at predicting the grinding kinetics of the coarsest size fractions fed to the mill. AMEC is in agreement with this recommendation, which will require full core sections for testing.

16.2.4

Selection of Optimum Grind Size

A series of batch rougher tests were conducted at different primary grind sizes with composites of DS (DSL+DSU), GS, VB, MDBX, and AO rock types. An economic evaluation model was developed from the resulting metallurgy, with relative revenue levels and associated operating cost estimates calculated to compare the different primary grind target scenarios. The trade-off analysis also considered the impact of lower mining cutoff grades.

The optimum grind size is influenced by the metal prices scenario assumed, with incremental recovery gains at a finer grind warranting the incremental costs as revenue differentials increase with higher metal pricing scenarios. For all rock types tested, (except AO) reducing the grind size, from 175 µm to 200 µm to 135 µm to 165 µm resulted in net revenue gain using a gold price range of $275/oz to $375/oz and a copper price of $0.75/lb. For AO, the metallurgical results were much less influenced by the fineness of grind within the range of grinds tested. The optimum grind size for AO decreased from 324 µm to 319 µm for the range of gold price indicated above.

A similar set of calculations using a higher price for copper of $0.95/lb reached similar conclusions but for finer optimum grinds. The lower gold price of $275/oz would call for a grind of 150 µm 180 µm as optimum while the high price of $375/oz is reducing these values to the 110 µm 145 µm range. Again, AO stands out with a consistent optimum grind throughout, in the much coarser 310 µm 315 µm range.

In all cases, the high end of the optimum P₈₀ ranges shown above is linked to GS, with DS establishing the low end and VB-MDBX yielding similar optimum P₈₀ just below the respective mid-range values.

The current metal pricing environment validates the selection of a fine grind target for flotation feed, with a P₈₀ of 120 µm being retained.

16.2.5

Flotation

Four main phases of flotation testwork were undertaken after Placer Dome became involved in the Cerro Casale Project.

Phase One

An initial phase consisted of 94 mapping tests on assay coarse rejects taken from 19 drill holes. Specific intercepts within these holes, representing single rock types and/or alteration patterns, were tested individually as 88 composites in order to establish the variability of the flotation response and decide how many discrete rock types should be defined for further flowsheet development. The results for these samples, as regrouped under four different rock type headings, were also used to establish by regression analysis the expected ore response versus feed grade equations later used for the economic model.

This work was performed by Placer Dome Research Centre (PDRC), using test procedures optimized during previous testwork on six rock type composites prepared from a single drill hole (CCD070, duplicating CCD019, in the northeastern portion of the orebody). The reported average flotation feed P₈₀, once grouping of the results per rock type was made, is between 109 µm and 130 µm.

Phase Two

Based on the response of individual samples previously tested, the mapping composites were recombined into seven new composites representing DSL, DSU, GS, MDBX, VS, VB and AO rock types. The optimization of flotation procedures, including determination of optimum primary grind size, was sought. Emphasis was placed on tests involving the DSL and DSU materials. These were then deemed as representing 49% of the orebody. Both PDRC and Met Engineers Ltd contributed to this testwork phase.

The tests were performed with flotation feed P₈₀ varying from 157 µm as the finest grind used, for trials with MDBX, up to 236 µm for AO. The average P₈₀ of all trials was 203 µm.

Phase Three

A pilot plant campaign was completed at the PDRC facilities. It involved six rock type composites (DSU, DSL, GS, MDBX, VB+VS, AO), which were tested at the PDRC as-is but mostly as mixtures thereof, using ratios to create a pilot plant feed stream representative of expected mine output over time (DSU, DSU/MDBX/VB+VS, DSU/MDBX/VB+VS/AO, DSL/GS/AO, DSL/GS). The P₈₀ for these trials was between 192 µm and 245 µm.

Confirmation bench-scale tests were completed in parallel by G&T, per their Project KM819, on five composites (DSL, DSU, GS, MDBX, VB), at similar coarse grinds as those used at the PDRC.

Important conclusions reached from this program, also applicable to the finer grind size adopted for plant design, included the realization that overdosing potassium amyl xanthate (PAX) in the scavengers was responsible for higher weight recoveries that did not yield improved metal recoveries. As well, it was found that the addition of cyanide in the flotation cleaning circuit (@ 8 g/t of plant feed equivalent) was beneficial in depressing pyrite more effectively than relying solely on the elevated pH levels used earlier. Improved final concentrate grades can thus be achieved. Finally, blending oxide material with the sulphides as mill feed would have to be done within limited ratios to prevent an excessive degradation of the final concentrate grade.

Phase Four

A series of batch tests, in open and locked cycle, were performed by G&T just before the 2000 Feasibility Study Report was completed. Composites of the rock types DSL, DSU, GS, MDBX, and on individual components later grouped under the VS designation (MVF, VB and VPF) were tested.

The first testwork program was mostly exploratory in nature, while the second and third ones listed above used much coarser grind targets than the one retained for preparation of the Feasibility Study Report (P₈₀ = 120 µm). The derivation of expected metallurgy, flotation and leach kinetics, as well as reagent consumptions is thus mainly based on the last testwork program, realized by G&T. Earlier testwork on 88 composites was also considered since it remained the most comprehensive source of information for assessing the metallurgical behaviour under variable feed grade conditions.

The flotation flowsheet adopted to carry out these programs is as presented in Figure 16-1. It consisted of a rougher/scavenger section, out of which the bulk of the final tailings are produced, as scavenger tails. The combined rougher/scavenger concentrate is sent to a regrinding stage where it is ground to a P₈₀ of 30 µm. This ground concentrate then proceeds to three stages of cleaning, with the first stage operated in open circuit with the last two recirculating their tailings to the freed to the regrind circuit.

The concentrate stream produced from the third cleaning stage is the final concentrate. The first cleaner tailings are routed to the gold cyanidation circuit, before being discarded (following cyanide destruction) along with the scavenger tailings in the tailings pond.

Figure -:

Flotation Circuit Configuration for Laboratory Testwork

[amectechnicalreport43101194.jpg]

16.2.6

Gravimetric Concentration

Gravimetric concentration was attempted with the seven composites (DSL, DSU, GS, MDBX, VB, VS, AO) created from blending the original mapping samples. In all cases, the gravity concentrates produced failed to reach acceptable grades for considering this approach as a way of recovering gold.

This outcome is consistent with G&T‘s gold behavioural model prepared for five of the composites (DSL, DSU, MDBX, VB, GS), which indicated that about 85% of the gold will behave like chalcopyrite in the flotation circuit, confirming a close mineralogical relationship between the main copper-bearing sulphide mineral and the gold.

16.2.7

Cyanidation

Cyanidation of tailings from the flotation circuit was investigated in parallel with the flotation testwork programs to recover additional gold.

A head sample of DSU material, ground to a P₈₀ of 200 µm, achieved a gold recovery of 81.3% (bottle roll test performed by PDRC during second phase of flotation testwork). Repeating similar tests with the same composite, but with the tailings from the first flotation cleaning stage (likely at P₈₀ of 25 µm), achieved an average gold recovery of 87.9% despite variable feed grades (at a constant residue grade of 0.05 g/t Au).

Standard bottle-roll cyanidation trials were conducted on both the scavenger and first cleaner tailings of various pilot plant runs completed at the PDRC in 1999. The average gold extraction from the scavenger tails (with the primary grind P₈₀ then used of about 200 µm and a cyanide concentration in the leach solution maintained at 250 ppm) was only 59%, with individual results ranging between 57% and 65%. In contrast, the recovery from the first cleaner tails surpassed 80% when a concentration of 1,000 ppm cyanide in the leach solution was maintained, but falling drastically if this concentration was reduced to 500 ppm (62.8% vs. 82.7% for tests with DSL+GS; 42.5% vs. 81.3% for tests with DSU). Copper dissolution was variable between ore composites, yielding higher values with those containing more bornite: DSU, MDBX and VS had 40% to 50% of their first cleaner tails copper content leached while DSL and GS leached 55% to 65%.

During the last phase of flotation trials (G&T, January 2000 report), first cleaner tails of multiple locked cycle trials were composited proportionally to each rock type in the mineralization. The resulting composite, having an average grind size P₈₀ of 20 µm, was submitted to cyanidation by standard bottle roll and cyanidation/carbon-in-leach (CIL) tests. For the bottle roll tests, a minimum concentration of 500 ppm of cyanide was required to optimize gold extraction. This equates to a consumption of 1.5 kg/t, coupled with 1 kg/t of lime (at 80% available CaO). In this fashion, a gold recovery of 88% could be achieved within 20 hours. Under similar conditions but at a reduced cyanide concentration of 250 ppm (1 kg/t consumed), the gold recovery decreased to 79%. At the 20 hour mark, about 25% of the copper in the feed was dissolved, regardless of the cyanide concentration maintained (with faster initial leaching kinetics differentiating the higher concentrations).

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 solubilized 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. This is likely caused by the additional demand brought by the improved copper leached.

16.2.8

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 that ultimately selected. Further settling testwork 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 an average conventional thickener unit area of 0.30 m²/t/d (0.20 m²/t/d for GS and VPF, with MVF, MDBX, and VB settling faster; and 0.36 m²/t/d for DSU, with DSL settling faster) could be used for design purpose. The resulting underflow density was 55% solids.

16.2.9

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 could not achieve a residual cake moisture below the transportable moisture limits (TML). Residual moisture above 17.5% were indicated, relative to TML (as determined by SGS) of 11.8% for DSU concentrate from pilot plant runs #16-17, and 11.3% for DSL+20%AO of pilot plant run #25. The equivalent cake moistures achieved by pressure filtration were below 12%.

Characterization of concentrate self-heating and pyrophoricity was not completed. With no pyrrhotite reported in the mineralogical composition of any rock types and secondary copper minerals found in the rock types limited to bornite (typically between 15% to 25% of total copper content), it is likely that the self-heating potential of the copper concentrate produced with Cerro Casale ores will be low. Nevertheless, the worst-case scenario (e.g., concentrate with VB or DSL ores) may warrant confirmatory testing.

16.2.10

Slurry Rheology

Viscosity measurements were reportedly made by Pocock on some thickened concentrate and tailings products obtained from the pilot plant trials realized by PDRC. The outcomes of such trials are not discussed in Pocock documentation; therefore, it is not known if high-torque specifications are required for thickening equipment.

16.2.11

Water Treatment

The introduction of a leach circuit on the first cleaner tailings requires a cyanide destruction circuit. The circuit design has to be capable of dealing with the elevated amounts of dissolved copper encountered as well as cyanide.

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.

16.2.12

Concentrate Minor Elements Composition

Assays of minor elements for some of the concentrates produced during the pilot plant trials (#16-17 on DSU, #25 on DSL) were reported by SGS. The DSU concentrate assayed 30 ppm mercury (usual penalty levels for Hg start at 10 ppm, with a typical levy of US$2/ppm above 10 ppm) but DSL contained only 6 ppm. Regular blending in the mill feed stream of the potentially offensive ore types will be required to reduce mercury content below penalty levels.

Both concentrates were below the normal 2.5% penalty level for combined lead and zinc contents, with about 1.8%. Some 2 m intercepts reported in the drill hole database are indeed quite high in zinc but this is accompanied with high copper grades as well, which will tend to keep in check the final zinc content in the concentrates.

16.3

Metallurgical Recoveries

Results obtained from the 88 composites tested in the first phase of flotation testwork were retained to establish the variability of the rougher flotation response relative to feed grades. Results were grouped by four different rock types for which rougher recovery versus feed grade was available. Regression equations were then developed to predict recovery for each element and adjusted by changing the intercepts of the linear equations while maintaining their slopes to intersect the actual results obtained from the locked cycle tests of the last phase of flotation testwork (G&T, January 2000 report)

The locked cycle test results were also used to derive the expected copper grade of the final concentrate and cleaner circuit recoveries for copper and gold. The mathematical expressions of the predictive equations are presented in Table 16-3.

Table -:

Placer Dome Model of Predictive Metallurgical Equations (from PDTS, 2000, Table 2.2)

Lithologies	Gold Recovery	Copper Recovery¹
Diorite Sulphide Upper	ST = 0.17 x Au + 0.090 RR = 100. – (ST x 137550) ¤ (Au x 150000) x 100. CLR = .985 Rec_Au% = RR x CLR	ST = 0.026 x Cu + 0.018 RR = 100. – (ST x 137550) ¤ (Cu x 150000) x 100. CLR = .965 Rec_Cu% = RR x CLR CG = 22.7%
Diorite Sulphide Lower	ST = 0.17 x Au + 0.095 RR = 100. – (ST x 137850) ¤ (Au x 150000) x 100. CLR = .985 Rec_Au% = RR x CLR	ST = 0.026 x Cu + 0.023 RR = 100. – (ST x 137850) ¤ (Cu x 150000) x 100. CLR = .964 Rec_Cu% = RR x CLR CG = 24.2%
Granodiorite Sulphide	ST = 0.116 x Au + 0.0574 RR = 100. – (ST x 137550 ¤ (Au x 150000) x 100. CLR = .985 Rec_Au% = RR x CLR	ST = 0.029 x Cu + 0.0112 RR = 100. – (ST x 137550) ¤ (Cu x 150000) x 100. CLR = .966 Rec_Cu% = RR x CLR CG = 25.7%
Volcanic Sulphide	ST = 0.17 x Au + 0.040 RR = 100. – (ST x 140200) ¤ (Au x 150000) x 100. CLR = .985 Rec_Au% = RR x CLR	ST = 0.064 x Cu + 0.0 RR = 100. – (ST x 140200) ¤ (Cu x 150000) x 100. CLR = .978 Rec_Cu% = RR x CLR CG = 27.6%
MicroDiorite Breccia	ST = 0.17 x Au + 0.063 RR = 100. – (ST x 139500) ¤ (Au x 150000) x 100. CLR = .985 Rec_Au% = RR x CLR	ST = 0.026 x Cu + 0.021 RR = 100. – (ST x 139500) ¤ (Cu x 150000) x 100. CLR = .976 Rec_Cu% = RR x CLR CG = 28.6%
Oxide	ST = 0.17 x Au + 0.095 RR = 100. – (ST x 135000) ¤ (Au x 150000) x 100. CLR = .985 Rec_Au% = RR x CLR	ST = 0.078 x Cu - 0.0026 RR = 100. – (ST x 135000 ¤ (Cu x 150000) x 100. CLR = .970 Rec_Cu% = RR x CLR CG = 12.2%
Catalina Breccia	ST = 0.17 x Au + 0.063 RR = 100. – (ST x 139500) ¤ (Au x 150000) x 100. CLR = .985 Rec_Au% = RR x CLR	ST = 0.026 x Cu + 0.021²RR = 100. – (ST x 139500) ¤ (Cu x 150000) x 100. CLR = .976 Rec_Cu% = RR x CLR CG = 28.6%
Other	ST = 0.116 x Au + 0.0574 RR = 100. – (ST x 137550) ¤ (Au x 150000) x 100. CLR = .985 Rec_Au% = RR x CLR	ST = 0.029 x Cu + 0.0112 RR = 100. – (ST x 137550) ¤ (Cu x 150000) x 100. CLR = .966 Rec_Cu% = RR x CLR CG = 25.7%

Notes: 1. Table 2.2 is mistakenly showing the gold head grade as being an input to the calculation of the rougher recovery (RR), while the copper head grade should have appeared in its place. 2. Corrected intercept from 0.018 to 0.021 to copy MDBX equations, as intended.

Recovery functions are multiple regression equations that incorporate the effect of the grind size in the expected rougher tailings grade for gold and copper. The setting of a primary grind target of 120 µm allowed simplification of the equations’ expression to what is displayed in Table 16-1.The Catalina Breccia was assigned the recovery function of unit MDBX. Category “Other” (minor materials) was assigned the same recovery function as unit GS.

The equations of Table 16-3 are estimates of scavenger tailings (ST) copper and gold grades, final concentrate copper grade (CG) and cleaner recoveries (CLR) for copper and gold. The additional equations make use of the scavenger tails grades and expected rougher weight recovery to derive a rougher recovery (RR) required to predict metallurgical results required to assign a value to the ore block. The block model evaluation routine (of “orevnu.f”) is moreover capping both copper and gold recovery to a maximum of 94%.

Table 16-3 underlies the fact that AO must be processed as a diluted blend (<30% of overall weight) with sulphide-bearing rock types since a more concentrated blend will result in concentrate grades too low for smelters.

The fixed rougher weight recovery per rock type, as implied by the equations in Table 16-3, provides an approximation of the actual rougher weight recovery since this amount would vary with the type of minerals found in the feed stream. The indicated rougher weight recoveries per rock type are not in line with the average values obtained from the locked cycle tests performed by G&T where MDBX, GS, DSU and DSL were showing an average weight recovery of 8% and Table 16-3 is calling for 8.3% for DSU and GS, 8.1% for DSL and 7.0% for MDBX. For the VS material (tested individually as VB, MVF and VPF by G&T), the actual proportional weight recovery of 7.4% obtained should be compared with the Placer Dome model setting said recovery at 6.5%. Higher weight recovery than the actual test results would overstate the rougher recovery calculated. The discrepancies noted above are within the level of accuracy of the whole revenue calculation steps required and are therefore not considered as significant.

In Table 16-3, the fixed gold recovery in the cleaning stage, of 98.5% for all rock types, represents the overall contribution to the recovery expected from both the flotation and the leaching of the first cleaner tails.

The agreement between the Placer Dome model and actual locked cycle test results is generally good, reflecting mostly that the former hold equations specifically adjusted to intercept said results. Only one major deviation from actual locked cycle test results (G&T KM1011 testwork) and projected values per the Placer Dome model was registered: for GS, tests 15 and 16 with average head grades of 0.31% Cu and 0.84g/t Au achieved gold rougher recovery of 75.6% to 76.1% while projection from Table 16-3 would yield 83.1%. This difference is not deemed to have an impact on the revenue calculations.

16.4

Net Smelter Return Calculations

The mathematical expressions for recovery and grade developed by Placer Dome are presented in the previous section.

These, along with indicated smelting contract conditions to be applied to the sales of the copper concentrate and doré gold, as found in the “orevnu.f” file, are used 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 $5.25/t of dry concentrate is indicated but the details (e.g., ore types, contents, scales of application and penalty rates) used to derive this number are not indicated. From Section 16.2.12 of this report, the most likely element to incur smelting penalties will be mercury.

Payment for silver in the copper concentrate is assumed. Silver revenues are not supported with recovery of silver in metallurgical testwork; therefore, revenue from silver should not be included in economic analyses until these data are available.

16.5

Plant Design Criteria

The sizing of individual equipment, from the application of the adopted design criteria, was not completely confirmed during AMEC’s review. Testwork parameters and procedures applied to achieve the scale-up were assessed by AMEC and found to be following industry standard practices.

16.5.1

Throughput Capacity

Plant design criteria have been prepared for an average throughput of 150,000 t/d of ore. This in fact is reflecting the minimum capability of the plant since the relevant grinding circuit parameters for the hardest ore type, MDBX, were applied (MacPherson’s correlated autogenous work index of 18.4 kWh/t and Bond ball mill work index of 18.3 kWh/t). Operated throughputs for all ore types are 163,000 t/d, with a 92% plant availability. The oxides would be capable of a higher throughput still, at 180,000 t/d. This maximum throughput would not be achieved in the plant, and thus does not have to be accommodated for in the downstream processes of the plant, since the proportion of oxide in the mill feed stream is to be kept below 30% in order to achieve saleable concentrate copper grades.

The resulting grinding circuit design is capable of an average throughput above the nameplate capacity of 150,000 t/d, with the exception of days with solely MDBX entering the mill. A higher average throughput of 165,000 t/d was therefore adopted for the first years (2 to 8) of operation, which possibly could be accommodated with the grinding circuit power base.

16.5.2

Design Feed Grades

The design feed grade selection influences the sizing of the regrinding and cleaning circuit by dictating mass pull from the roughers. Peak processed tonnages are used to calculate the required flotation cell volumes and regrinding circuit power base.

A design copper feed grade is not given in the design criteria. A maximum rougher weight recovery may have been assumed but it is not stated.

16.5.3

Crushing and Grinding

AMEC made cursory verifications of the major equipment sizing to confirm the capacity of the chosen equipment to process the range of ores expected. Equipment checked was found to be sufficient for the required use.

Two independent lines of crushing and stockpiling, and three parallel lines of ore reclaiming and SABC (SAG milling, pebble crushing and ball milling) are required to achieve the design throughput. The variable speed capability implemented for the SAG and ball mill motors will allow handling a large range of ore blends. This feature will ensure a flexible operation, providing flotation feed of target size and throughput tuning capability.

The reduction of the stockpile live capacity, from an earlier 135,000 t to 100,000 t is not advisable: on an availability adjusted operating basis, 100,000 t equates to just over 13 hours of operation, at the nominal 165,000 t/d sought in the first years of operation. If forced too regularly to push dead material over the feeders, additional rehandling costs are incurred and a more erratic feed size distribution feeding the SAG mills will lead to difficulties in maintaining a steady mill operation. The presence of two primary crushers may alleviate somewhat the usual design criteria that would call for 16 to 24 hours of live capacity for a SAG mill-feeding stockpile.

Provision for a covered stockpile may be justified, either from environmental obligations or for a requirement to control dust from spreading over large areas of the property and potentially damaging to some surrounding equipment (electrical switchgear), to the safety of personnel traveling on surrounding access or mine production roads, or to the general work environment in the yard or buildings. If a cover is to be excluded from the final design, dominant wind directions are to be carefully evaluated and no infrastructure should be planned in the downwind direction.

Care will have to be taken to segregate properly the AO and other sulphide rock types on the stockpile, as well as providing AO to at least one feeder of each of the three grinding circuits. Failing this, the proportion of AO presented in the feed of each line will be difficult to achieve and maintain at the levels indicated per the short-term mine planning.

The MacPherson correlated autogenous work index determination is the preferred approach to sizing of a SAG mill while the standard Bond approach can be used for the ball mill, with some power credit for feed preparation with a SAG mill instead of a rod mill allowable. The JKSimMet software provides additional insight in the design and is mostly used for simulating the effect of proposed changes within existing plants. Its use requires an extensive sampling survey with data to first be inputted in the simulator to obtain a calibrated model for the plant. The JKSimMet simulations performed by Contract Support Services (CSS) and the benchmarking of the Cadia Hill operations, where SAG mill of the dimensions considered at Cerro Casale are in use, indicate the SAG mill will be the limiting factor in the grinding circuit. This eliminates the possibility of considering only two lines of SAG milling, followed by three ball mills per SAG mill, instead of the current three lines of SAG milling followed by two ball mills per SAG mill. Larger ball mills could then have been considered, as a way of reducing further the capital and operating costs for the grinding circuit.

One of the recommendations expressed by CSS was to obtain drop weight test parameters on core fragments from a larger drill core diameter to minimize the potential bias that the use of the smaller cores may have induced in the interpretation of the results. AMEC supports CSS’ recommendation because if further testing demonstrates that SAG mill capacity is not the constraint, a substantial reduction in the capital cost could be realized by replacing three parallel lines with a dual SAG mill (each with three ball mills).

16.5.4

Flotation

The flotation circuit adopted for the plant design follows the circuit configuration used in the latest pilot-plant trials, per Figure 16-1.

Retention times were specified on the basis of the locked cycle trials of G&T (2000). Usual scale-up factors were used to pass from laboratory to industrial plant requirements. The resulting factors are indicated in Table 16-4.

Table -:

Design Retention Time Derivation for Flotation Stages

Flotation Stage	Flotation time (min)		Scale-Up factor	Cell Volume Provided (m³)
Flotation Stage	Laboratory	Plant (retained)	Scale-Up factor	Cell Volume Provided (m³)
Roughers	15	30	2.0	9600 (6 lines x 10-160 m³ cells)
1^st Cleaners	7	12	1.7	840 (3 lines x 7–40 m³ cells
2^nd Cleaners	4	6	1.5	120 (3 lines x 4–10 m³ cells
3^rd Cleaners	2	5	2.5	90 (3 lines x 3–10 m³ cells

For an operated feed rate of 163,000 t/d, the cell volumes provided are sufficient to maintain the desired plant retention time, allowing as well for an aeration factor estimated at 12%. The indicated design rougher feed density of 35% solids (actually indicated as ball mill cyclone overflow density) seems high although: benchmarking indicates 27% to 30% as more the norm for porphyry. A lower value would drive the volume requirement upwards and bring the retention time provided below the selected 30 minutes.

Tank cells of 200 m³ have been installed at other mine sites for test purpose and a few are on order for full plant circuits. They may be considered for potential capital cost reduction, if indicated performance feedback is positive.

The flotation circuits use established tank and conventional (rectangular or U-shaped) flotation cells. Recirculating loads within the cleaning circuits are minimal, per the locked cycle tests indications, and thus have a marginal influence on the cell volume requirements calculated.

Second and third cleaning stages rely on small cells, with a limited number installed per bank (4 and 3, respectively). Three cells are considered normally as a minimum to prevent undue short-circuiting within the cells. Adoption of cells with a smaller individual volume (8 vs. 10 m³) would push the number of cells required upwards and mitigate this potential problem. Alternatively, one row of cells for each stage of flotation could be provided, with somewhat larger cells. The last units could then be operated without aeration upon shutdown of one SAG mill line for maintenance purpose.

Tower mills are used for regrinding duties, with two units of 900 kW each provided for each of the three parallel cleaning circuits. As per the above, the regrinding duty could be redesigned in a single grouping, with five 1,100 kW units needed to cover and equivalent duty when all SAG mill lines are operated and individual units shut down as required with lower total throughput.

All the flotation and regrinding circuit equipment selections are proven technologies and are appropriate for the test results.

16.5.5

First Cleaner Tails Leaching

The leaching time of 24 hours with a free cyanide concentration in the slurry maintained at 500 ppm is in agreement with the indicated requirements from testwork data.

The balance of the circuit, including the leaching and CIP tanks, carbon stripping and regeneration systems, is of conventional design. Viscosity tests were performed by Pocock but were not available for AMEC’s review. The design of the agitation mechanism for the leach and carbon tanks should have taken the outcome of these tests to ensure appropriate installed power base and feed density selection (set at 40% solids) to these process stages.

The refinery equipment may require the addition of a mercury retort furnace, considering the high mercury content found in some of the flotation concentrate.

16.5.6

Thickening

A high-rate thickener is specified for the final tailings duty while conventional units are sought for the concentrate and first cleaner tails, to minimize losses in the supernatant.

Thickeners (at 30 m diameter for concentrate, 2 m to 91 m diameter for tailings and 80 m diameter for the 1^st cleaner tails) have been sized from the raw testwork data.

It is customary to add a safety factor of 25% above the conventional thickening test results to account for wall effects produced in the laboratory testwork procedures. A unit area requirement of 0.34 m²/t/d was used for the concentrate. This is to be compared with the indicated requirement of 0.15 m²/t/d to 0.40 m²/t/d stemming from the testwork, with the high end of this range only applicable to dealing with 100% DSU concentrate. Since this type of feed stream is very unlikely in the plant, the selected unit area is adequate. For the first cleaner tails, the provision of 0.42 m²/t/d is in line with the range of 0.23 m²/t/d to 0.60 m²/t/d from testwork, with the high end due to pure AO ore type that will never be delivered to the mill by itself.

16.5.7

Filtration

A single pressure filter of a proven design (Larox) is specified for the filtration of the concentrate. CMC may want to consider two units of smaller individual capacity instead, with appropriately sized thickened concentrate stock tanks still to be provided ahead of the filters. This is a more flexible configuration, considering that the mechanical availability of such filters is about 85% while there is a likely minimum continuous requirement for filtering the concentrate produced from two out of the three grinding lines.

A partial decoupling between the processing plant and the filtration plant is provided in the current design by the concentrate thickener underflow tanks, the concentrate pipeline, and the receiving tanks at the filtration plant. Despite these provisions, if the single filter was to become incapacitated beyond the demands for regular preventive maintenance, it may become a factor limiting the process plant availability. AMEC believes having two concentrate filtration units rather than one will allow more flexibility and will reduce the possible reduction in plant capacity.

16.5.8

Water Treatment

The cyanide destruction treatment system design is based on the Inco/SO₂ system, with SO₂ being provided by an elemental sulphur burner. This is a proven technology, especially suited to deal with free and complexed cyanide species.

The circuit design was based on destroying up to 320 kg/h of cyanide complexes dissociable in weak acid solutions (CN_wad), containing as well up to 280 ppm of copper. The treated slurry would then contain less than 1 ppm of CN_wad.

16.6

Discussion

The recommendation made by CSS, requesting a new round of drop weight test parameter determinations on core samples of larger diameter than those available earlier, should be followed upon. These tests may lead to the realization that the SAG mill capacity is not as limited as now assumed, which may allow for a reconfiguration of the primary grinding circuit into two lines instead of three with each one comprising a SAG mill followed by three ball mills. The latter may be of a larger diameter than currently envisioned but the overall capital cost will be lower than currently estimated.

The calculation of the optimum grind target was comprehensive and is appropriate for the present metal prices.

Metallurgical recovery projections were found to generally reflect the results of the last locked cycle test campaign completed by G&T.

Design criteria that could be examined in detail were found to be sufficiently conservative relative to interpretation of the available metallurgical test data.

17.0

MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1

Mineral Resource Estimates

17.1.1

Introduction

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 (OP). Project limits are 4000 to 4405 East, 23900 to 27905 North, and 2506 m to 5005 m elevation. Projects limits are in truncated UTM coordinates with 470000 subtracted from easting coordinates and 6900000 subtracted from northing coordinates. Cell size was 15 m east x 15 m north x 17 m high.

17.1.2

Geologic 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.

Table -:

Gold and Copper Geologic Models or Domains, Cerro Casale Project

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

17.1.3

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. 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.

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Figure -:

Boxplot Summary of Gold Composite Data (Un-cut), Cerro Casale Project (from PDTS, 2000)

[amectechnicalreport43101200.jpg]


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Figure -:

Boxplot Summary of Gold Composite Data (Cut Grades), Cerro Casale Project (from PDTS, 2000)

[amectechnicalreport43101202.jpg]


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Figure -:

Boxplot Summary of Copper Composite Data (Un-Cut), Cerro Casale Project (from PDTS, 2000)

[amectechnicalreport43101204.jpg]


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Figure -:

Boxplot Summary of Copper Composite Data (Cut Grades),
Cerro Casale Project (from PDTS, 2000)

[amectechnicalreport43101210.jpg]

17.1.4

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 centre of the deposit. Higher copper grades are generally found at depth.

Figure -:

Gold vs. Copper Scatterplot, Cerro Casale Project (from PDTS, 2000)

[amectechnicalreport43101212.jpg]

17.1.5

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.

17.1.6

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.

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 -:

Cutting Thresholds or Cap Grades for Gold and Copper Composite Data, Cerro Casale Project

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.

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.

For both metals, the undefined domains (G15 and C15) were estimated using the respective variogram parameters of the volcanic, no stockwork intensity domain.


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Table -:

Gold and Copper Variogram Parameters for Cerro Casale Project Estimation Domains

	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


Project No.: 148562	Page #
June 2005

	ARIZONA STAR
	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

17.4

Estimation

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.

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.

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.

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-neighbour estimates. (The nearest-neighbour 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.

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-neighbour 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-neighbour estimate or declustered composite data, thus the nearest-neighbour 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 and Summary

Table -:

Global Model Mean Grade Values by Domain, Cerro Casale Project

	Nearest-Neighbour 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.

The mineral resources of the Cerro Casale project were tabulated based on material that fell within an optimistic ultimate pit shell. That shell was made based on a gold price of US$500/oz, US$1.25/lb for copper, and estimated costs (from experience on similar mines) of US$3.39/t. This shell fulfilled the expectation of reasonable extraction test in declaring mineral resources at Cerro Casale. AMEC agrees with this logic and its implementation.

The mineral resources for Cerro Casale are reported at a gold cutoff grade of 0.4 g/t and are shown in Table 17-5. Mineral resources are entirely inclusive of mineral reserves. Mineral resources were last estimated in January 2000. These have not changed and are compliant with Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000) and Canadian National Instrument 43-101.

Table -:

Cerro Casale Mineral Resources – January 2000

Mineral Resource Category	Tonnage (Kt)	Grades		Contained Metal
Mineral Resource Category	Tonnage (Kt)	Au (g/t)	Cu (%)	Au (M oz)	Cu (Kt)
Measured	202,800	0.75	0.25	4.89	500
Indicated	910,800	0.70	0.26	20.53	2,368
Measured+Indicated	1,113,600	0.71	0.26	25.42	2,868
Inferred	171,200	0.63	0.33	3.49	565

AMEC considers the resource estimate as prepared by Placer Dome in 2000 to be reliable and relevant.

17.6

Mineral Reserves

Strategies for different cutoff values by mining stage are listed in Table 17-6. Ore in Stage 1 is defined as >$4.00/t profit cutoff for oxide ore and >$2.00/t profit cutoff for sulphide ore. The $4.00/t value was used only for Stage 1 to improve mill head grades and reduce the proportion of oxide ore from Stage 1 to the mill. Ore in Stages 2 and 3 is that material at profit cutoffs of $2.00/t for both oxide and sulphide ore. The ultimate pit reserves are stated at the minimum profit value cutoff of $0.25/t. Mineralization with a grade less than the cutoff values and greater than the minimum cutoff value of $0.25/t will be stockpiled and delivered to the mill as required periodically throughout the mine life. Oxide and sulphide ores will be stockpiled separately.

During the pre-production period high-grade oxide and sulphide ore will be stockpiled separately from the low-grade oxide and sulphide ore. Stockpiled high-grade will be reclaimed within the first two years of the production period. Stockpiled low-grade material will be reclaimed throughout the life of the operation as required.

Total Proven and Probable Mineral Reserves are listed in Table 17-7. A comparison of the total smoothed reserves and the total un-smoothed reserves shows that smoothing and the provision of access resulted in a 5.5% decrease in overall ore tonnage, and a 2.1% decrease in waste stripping. Gold and copper grades were virtually the same in both cases.

Table -:

Elevated Profit Cutoff Strategy (from PDTS, 2000)

Pit Stage	Oxide Profit ($/t)	Sulphide Profit ($/t)
Stage 1	4.00	2.00
Stage 2	2.00	2.00
Stage 3	2.00	2.00
Ultimate Pit	0.25	0.25

Table -:

Proven and Probable Mineral Reserves as of January 2000 using Elevated Profit Cutoff Strategy (from PDTS, 2000 Feasibility Study)

	Ore Kt	Au Grade (g/t)	Contained (oz Au)	Cu Grade (%)	Contained (Kt Cu)
Proven	200,266	0.72	4,636,000	0.25	501
Probable	834,799	0.68	18,251,000	0.26	2,170
Total Proven & Probable	1,035,065	0.69	22,887,000	0.26	2,671

Notes: 1. US$350/oz Au and US$0.95/lb Cu prices used. 2. Metallurgical recovery equations are noted in Table 16-3 of this report.

This reserve has a life-of-mine waste:ore strip ratio of 2.67:1.

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) and Canadian National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators.


Project No.: 148562	Page #
June 2005

18.0

OTHER RELEVANT DATA AND INFORMATION

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

	ARIZONA STAR
	CERRO CASALE PROJECT, CHILE
	TECHNICAL REPORT AND QUALIFIED PERSONS REVIEW

19.0

REQUIREMENTS FOR TECHNICAL REPORTS ON PRODUCTION AND DEVELOPMENT PROPERTIES

19.1

Mine Plan

The mine plan incorporated in the 2000 Feasibility Study envisions that the Cerro Casale deposit will be mined using conventional open pit methods. The mine plan features a four stage open pit which is scheduled to deliver a nominal 150,000 t/d of mill feed at a maximum mining rate of 816,000 t/d over an 18 year mine life. The final pit stage will measure over 2,100 m in diameter and the highest sector of highwall will have a vertical extent of 1,224 m, ranking the proposed final pit wall amongst the world’s highest.

The following sections provide summary descriptions of the key mine planning steps. Mine plans were developed using Placer Dome’s internally developed and proprietary ‘Pit Optimization Package’ (POP) software package, with exception of the stage limits which were defined using the NPV Scheduler package, and the smoothed pit stage design which was performed using the Vulcan mine planning package. AMEC considers these mine planning software packages to be robust, well accepted by the mining industry and appropriate for assessing the mining potential of the Cerro Casale deposit.

19.1.1

Economic Modelling

A net revenue block model, referred to as the Profit Model, classifies each block as ore or waste. If a block generates a positive profit value greater than $0.25/t after the deduction of all costs associated with mining, processing, administration, freight, smelting and refining from the recovered metal value, the block is assigned that profit value. If a value of at least $0.25 is not generated, the block is classified as waste and assigned a negative value equivalent to the cost of mining that block. The $0.25/t profit criterion represents the cost of rehandling stockpiled material to the primary ore crusher. All low-grade ore, which may be stockpiled for future processing, must be able to bear the rehandle cost.

Figure -:

Site Plan Map (from PDTS, 2000)

[amectechnicalreport43101226.jpg]

Only measured and indicated resources are treated as ore. Inferred materials are treated as waste in the profit model.

AMEC has reviewed the economic modelling methodology and parameters applied. They are considered to be standard practice 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, and are discussed in more detail in Section 16.3 and Table 16-3 of this report. 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 cost of processing the different ore types used in the open pit optimization is variable, due in part to the differences in cyanide consumption characteristics and crushing/grinding work indices of the ore types. The costs applied by process group and downstream product costs are listed in Table 19-1. The cost of processing, administration, and plant services has been estimated on a dry tonne of ore basis.

Table -:

Process Costs

Process Group	Diorite Sulphide Upper & Lower	Granodiorite Sulphide & Undefined	Volcanic Sulphide	Breccias	Oxide
Process Costs $/t	2.63	2.82	2.79	2.93	1.83
Cost Item	Units	$/Unit
Freight	$/t concentrate	31.00	Land + Ocean
Smelting	$/t concentrate	87.00	Includes $2.00/t penalty.
Refining	$/lb of Cu $/oz Au in conc.	0.09 6.50	-
Deduction	%	-	1 percentage point deduction from concentrate grade
Participation	%	-	10% of copper price variance from $0.90/lb
Marketing & Insurance	%	-	0.03% of product value
Moisture	%	-	10% moisture content in shipped concentrate
Losses	%	-	0.2% of concentrate lost in transit
Administration	$/t ore	0.66	-
Plant Services	$/t ore	-	Included in Administration

Mining Cost

The mining cost and incremental mining cost per bench below the 4087 m elevation used in the profit model are listed in Table 19-2. All mining will be carried out on 17 m benches, with pit stage wall berms at 34 m intervals. De-watering requirements, which are discussed below, have an average operating cost of $0.001/t over the life of the mine, which were included in the mine operating cost estimate.

Table -:

Base and Incremental Mining Costs

Item	Cost $/t
Ore and Waste	0.43
Pit Entrance Elevation (m)	4087
Incremental Cost per Bench Below Entrance Elevation	0.005

The incremental cost increase per bench below the pit entrance elevation of one-half cent is reasonable.

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.

Metal Price

An average gold price of $350/oz and copper price of $0.95/lb was used in the profit model calculations.

Exchange Rates

The Profit Model was constructed in units of 1999 US Dollars. An exchange rate of US$1.00 = 510 Chilean Pesos was used. This differs from the exchange rate of US$1.00 = 600 Chilean Pesos that was used for estimating operating and capital costs in the 2004 Feasibility Study Update.

19.1.2

Pit Shell Optimization

Wall Slopes

Geotechnical assessments of Cerro Casale were conducted by Piteau Associates in 1998 and 1999, in their report “Aldebarán Project, Cerro Casale Sulphide Deposit, Feasibility Geotechnical Assessments for the Open Pit.” Piteau Associates’ involvement with the Cerro Casale project dates back to 1995, when they performed geotechnical studies for the then targeted oxide deposit.

Geotechnical investigations were conducted using data collected from six geotechnical diamond drill holes totalling over 5,000 m, piezometers installed in five of the geotechnical drill holes, exploration drilling, and mapping of both limited naturally occurring surface outcrops, exploration road cuts, trenches and small pits. Recommendations by Piteau Associates are preliminary and should be revisited in future updates of the feasibility study.

Piteau Associates recommendations for wall configurations (with 34 vertical metres of bench face between berms) are listed in Table 19-3.

Table -:

Recommended Slope Criteria (from Piteau Associates, 1999)

Preliminary Structural Domain	Sector Azimuth Range	Recommended Preliminary Interramp Slope Designs
Preliminary Structural Domain	Sector Azimuth Range	Interramp Slope Angle	Bench Face Angle	Berm Width (m)
Central Intrusives	228 – 40	38	55	19.7
	40 – 82	44	65	19.4
	82 – 142	38	55	19.7
	142 – 228	44	65	19.4
Volcanics / Volcaniclastics	90 – 226	38	55	19.7
	226 – 288	47	65	16.9
	288 – 318	38	55	19.7
	318 – 26	44	60	15.6
	26 – 78	41	60	19.5
	78 – 190	43	60	16.8
Volcanic Flows	226 – 270	38	55	19.7
	270 – 316	41	60	19.5
	316 – 84	46	65	17.0
	84 – 120	38	55	19.7
	120 – 226	42	60	18.1

The complexity of the above criteria, consisting of primary structural domains, subdivided by radially defined sectors, was simplified into the slope sector table shown in Table 19-4 below for pit optimization.

Table -:

Simplified Inter-ramp Slope Angle Criteria (from PDTS, 2000)

Pit Wall Position; Azimuth
0°	90°	170°	190°	266°	288°	318°	348°
Inter-ramp Wall Slope
44°	44°	40°	38°	38°	40°	42°	44°

The simplification of the inter-ramp slope recommendations is reasonable for pit optimization and acceptable for feasibility level pit design. For the next feasibility update, AMEC recommends the construction of a geotechnical block model, which will allow the generation of pit designs that will more closely honour the geotechnical engineer’s inter-ramp angle, bench face angle, and berm width recommendations.

Piteau Associates recommended that unbroken inter-ramp slopes in excess of 350 m vertical height be avoided in preliminary mine plans. The west side of the ultimate pit design has an unbroken slope with a nominal vertical height 750m and a maximum height of 850 m. For the next feasibility update, this slope will require the addition of two ramp width breaks in the slope to meet this design criterion. These slope breaks could be in the form of large cleanable catch benches or a second ramp system that would enhance the flexibility of the mine plan.

Optimization Methodology

The ultimate pit limits were defined using Placer Dome’s ‘Pit Optimization Package,’ 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 honours 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 four relatively large pit stages, targeting high production, low cost mining geometries. The pit stage limits were defined using NPV-Scheduler software, with each stage representing a nominal production period of from one to six years. Each pit was optimized to maximize the operating profit for that increment of production. This progression from the most profitable to the next most profitable increment guided the overall development of the pit staging.

AMEC considers the optimization methodology to be appropriate for this deposit.

19.1.3

Pit Stage Design

Methodology

The final design of the four pit stages involved the following procedure:

•

smoothing of the bench profiles

•

designing road access for each bench of each pit stage while maintaining workable pushback widths between stages

•

designing the bottom bench of each stage with minimum acceptable dimensions, given the relatively large size of the open pit mining equipment selected.

A 17 m high bench configuration was selected to maximize the efficiency of the specified large scale mining equipment. Although AMEC recognises the potential benefits of mining a 17 m bench as opposed to a more common 15 m bench, two operational issues could negate those benefits:

•

The specified blasthole drill has a maximum hole depth capacity of 19.8 m, which is equal to the 17 m bench height, plus planned subgrade of 2.8 m. A second drill rod will need to be added to obtain the proper subgrade in areas where the pit floor is higher than planned.

•

The specified loading tools will be unable to scale the full bench face back to the designed highwall without the assistance of dozer or backhoe slot cutting the crest from the bench above.

AMEC recommends using a standard 15 m bench height as the basis for feasibility level planning and suggests that with operational experience, upside potential exists to mine thicker benches.

Pit Access

Access to the pit from the topographic entrance level is via a 10% decline ramp. The entrance bench level is at elevation 4087 m for all pit stages. Ramps were 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 insufficient for ramps in which trolley assist will be used. For trolley assist ramps, a 44 m wide ramp is recommended to allow for passing trucks that have been pulled out of the trolley assist.

Access to Stage 1 is unique. Situated on the north side of Cerro Casale, Stage 1 requires a single ramp system heading west from the entrance and spiralling downward in a clockwise direction to the bottom of the pit at elevation 3951 m. The waste haul will be directly east of the pit entrance, and the ore will be hauled to the south around the east side of Cerro Casale, a distance of about 2 km to the primary crusher.

Access to Stages 2, 3, and 4 is by a common, double heading system. One haul road will descend to the northeast from the south ore pit entrance location, and the other will descend in a southwest direction from the east side waste entrance location. The two haul roads will meet at a common junction at elevation 4021 m. A single road will then continue to spiral counter-clockwise downward to the pit bottom.

Figures 19-2 to 19-5 show plan view images of the four stage designs. Figure 19-6 shows a generalized north looking sectional view of the four stages.

Figure -:

Stage 1 Pit Design (from PDTS, 2000)

[amectechnicalreport43101228.jpg]

Figure -:

Stage 2 Pit Design (from PDTS, 2000)

[amectechnicalreport43101230.jpg]

Figure -:

Stage 3 Pit Design (from PDTS, 2000)

[amectechnicalreport43101232.jpg]

Figure -:

Stage 4 (Ultimate) Pit Design (from PDTS, 2000)

[amectechnicalreport43101234.jpg]

Figure -:

North Looking Section through Pit Stages
(Generated by AMEC from surfaces provided by Placer Dome)

19.1.4

Stockpile and Dump Design

Low-grade Ore Stockpiles

Low-grade oxide and sulphide materials will be stockpiled for reclamation and processed either later in the mine life or as necessary to maintain mill feed tonnage. There will be separate oxide and sulphide stockpiles located to the east of the plantsite and constructed from the 4087m elevation on top of a portion of the waste dump.

Parameters used in stockpile design were:

•

Density of stockpiled material, rock

1.95 t/m³

•

Constructed overall dump slope

37°

•

Constructed overall stockpile height

40 m

•

Approximate basal area

110 ha

Maximum capacity of either stockpile will be in the order of 50 Mt of material. Combined maximum capacity of both the oxide and sulphide stockpiles will approach 80 Mt.

Waste Dumps

Approximately 2,768 Mt of waste, exclusive of stockpiled ore, will be mined over the 18 year life of the mining operation. Of this total, some 80 Mt of waste rock has been scheduled for tailing dam construction over the mine life.

The Río Nevado valley east and south-east of the open pit is well situated and able to contain the open pit waste. Valley floor elevations range from 3820 m in the north to 3725 m in the south end of the dump area. The waste will be contained between the eastern and western sides of the valley. The north face of the waste dump will form the downstream backing for the tailing 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 area will contain the requisite waste tonnage with a finished top elevation of 4160 m. The south side of the dump body will be 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 are:

•

Density of rock fill in dump

1.95 t/m³

•

Angle of repose of material

37°

•

Constructed dump slope, south end

22°

•

Constructed dump slope, north end

22°

Acid Rock Drainage

ARD assessment work has shown that most of the sulphur in Cerro Casale waste rock 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. Some of the rock and tailing 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.

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 or special handling / segregating of waste types in this study.

Mine Access Roads

All roads will be crowned and ditched to enhance drainage. Road dimensions and characteristics will be:

•

Width

35 m

•

Gradient

10% maximum

•

Berm

2.0 m high berm on outside edge of ramps

Dust suppression will be provided by three 90 t water trucks.

Tailing Dam Construction

The dam will be constructed by the mine, using a combination of locally available materials and pit run waste. Dam construction will entail a downstream construction technique utilizing run-of-mine waste rock that will be hauled to location with the mine haul-trucks. The downstream slope of the tailing dam will initially abut the north toe of the waste dump and eventually will be covered by the waste. The tailing dam and the north end of the waste dump will be constructed concurrently as needed to provide adequate volume in the tailing pond. Appropriate tonnages and distances have been included in the haulage cycle calculations to account for tailing dam construction.

19.1.5

Production Schedule

Summary

Production scheduling and mining rates were established from the following:

•

Significant pre-stripping in the first two pit stages will be required during the construction period. The last three months of Year -1 is deemed the commissioning period during which mill throughput will work up to nominal capacity.

•

Waste stripping will be required for each of the four stages.

•

Waste production will be scheduled to avoid large fluctuations in equipment and personnel requirements. A minimum pushback width of 100 m will be required for the size of equipment being considered.

•

Nominal milling rates will be 54,750,000 t/a (150,000 t/d). Additional low-grade ore will be included in the mill feed in the first eight years, providing material in excess of the nominal milling rate, to utilize 95% of available mill horsepower while processing the softer ores. Achieved mill throughput will maximize at 62,350 Mt/a in Year 2.

Mining rates start at 296 Mt/a in Year 1, peak at 298 Mt/a in Year 2 and decline to 32 Mt/a by Year 18. An elevated cutoff strategy and a high proportion of oxide ore stockpiling were employed in this study in an attempt to improve mill feed head grade in the early production years.

Another scheduling objective was to limit the oxide ore content in the mill feed. A ratio of 30% oxide to 70% sulphide was considered the maximum to provide a reasonable copper recovery and concentrate grade. Stockpiling allows the achievement of the requisite blend limits. If processed as mined from the pit, oxide ore would represent over 65% of the ore produced in Year 1. To alleviate this problem, oxide ore from the Stage 1 pit with a value less than $4.00/t was stockpiled and introduced into the mill feed over the mine life.

During the pre-stripping period, high-grade sulphide ore grading >$2.00/t and high-grade oxide ore >$4.00/t will be stockpiled apart from the lower grade stockpiles. These materials will be included in the ore stream within the first two years of production. The intent of the production schedule was to provide +$2.00/t ore to the mill from the onset of the production period. However, the mining rates necessary to achieve this goal were found to be unrealistic (+400 Mt/a). Therefore, low-grade ore from the pit and stockpiles will be included in the mill feed in order to provide the requisite mill tonnage and to maintain mining rates in a workable range of around 290 Mt/a.

AMEC has reviewed the detailed production schedule and finds that it conforms to the following design.

Stage 1

Pre-production stripping will commence in Stage 1 in Year -2 with high-grade oxide and sulphide stockpiled at elevated cutoff values of $4.00/t and $2.00/t of ore respectively. During Year 1 and 2 of this mining phase, the high-grade oxide and sulphide ores stockpiled during the pre-production phase will be reclaimed. The low-grade oxide and sulphide material below a cutoff of $4.00/t and $2.00/t will be stockpiled or included in mill feed as required. Mining will continue in Stage 1 until Year 2.

Stage 2

Pre-production mining will start in Year –2, concurrent with the Stage 1 pit. The low-grade oxide below $2.00/t and low-grade sulphide below $2.00/t will continue to be stockpiled or included in mill feed as required. The permanent haulage ramp system will be incorporated in the Stage 2 design from the entrance bench elevation of 4087 m to the 4011 m elevation. Mining will continue in this stage until Year 7.

Stage 3

Waste stripping starting in Year 3 will be limited to access development and bench establishment. There will be considerable pre-stripping required before ore is encountered in substantial quantities. As much as 480 Mt of waste will be stripped from this stage prior to ore extraction in Year 6. As with Stage 2 the low-grade oxide and low-grade sulphide materials below a cutoff of $2.00/t, not included with high-grade mill feed, will be stockpiled. This stage will be completed in Year 13.

Stage 4: Ultimate Pit

The stripping in the Stage 4 pushback starts in Year 7 and lasts for 4 years. Ore production will begin in Year 11. The internal cutoff for this pit phase is reduced to the minimum stockpile cutoff of $0.25/t. Production from Stage 4 will cease in Year 18.

19.1.6

Equipment

The mining rate at Cerro Casale starts at 296 Mt/a in Year 1 (811,000 t/d), and remains above 250 Mt/a for eleven consecutive years. The mining rate decreases to 160 Mt/a in Year 12 and decreases annually to 32 Mt/a in Year 18.

Open pit material movement will be carried out with haul trucks and a combination of electric shovels, hydraulic excavators, and large front-end loaders. The initial production fleet will consist of eight 381 mm blasthole drills, fifty 308 t trucks, five 1,200 t class electric shovels, two 575 t class diesel driven hydraulic shovels and three 25 m³ rubber-tired loaders. The truck fleet will be augmented by a trolley assist system in Year 5 as the majority of mining progresses below the pit entrance elevation of 4087 m.

Mining in the open pit will take place on 17 m benches. Double benching will be employed in the final pit walls of each pit stage, leaving a 20 m wide berm every 34 m.

AMEC believes the equipment selected is appropriate for a large low-cost open pit mining operation and that the scheduled build up of the mining fleet is appropriate for the production schedule.

19.1.7

Mine Operations Manpower Requirements

Summary

Manpower requirements are based on the equipment required to achieve the production schedule at the performance levels discussed in the section on Operating Costs. Pit operations will be scheduled for 24 hours per day, 7 days per week, 365 days per year. Four rotating crews and shift pit supervisors will be required, working 12 hour shifts on a seven days on, seven days off schedule. Non-shift personnel will work a 4 days on, 3 days off roster.

AMEC reviewed the staffing plan through construction and operation and considers the staffing plan to be appropriate for the mine plan.

Mine Maintenance Manpower Requirements

Major equipment suppliers will provide initial warranty coverage. Optional repair hire services can be negotiated. This would allow the Maintenance Department time to recruit and train mechanics, welders, and electricians. As this department expands and warranty volume diminishes, equipment supplier personnel could be phased out.

Notwithstanding the above, the maintenance costs for this study assume owner maintenance from the onset of mining operations.

Mine maintenance manpower was estimated using the approximate ratio of 0.7 maintenance people for every equipment operator. In Year 4 for example, there are 328 equipment operators and 232 maintenance personnel for a ratio of 71%. Number of people in each maintenance classification within the maintenance crew, i.e., mechanic, welder, etc., are pro-rated. AMEC has reviewed manpower statistics from Placer Dome’s Chilean operations and considers the 0.7 to one ratio to be appropriate.

Variations in crew complement occur on an annual basis, depending on the amount and age of equipment in use.

19.1.8

Major Equipment Productivity Estimates

Use of Time Assumptions

For the major equipment, the shovels, drills, and trucks, certain assumptions are made in order to establish the required operating hours, equipment productivity and number of units required. The base assumptions are:

•

Operating days per year

360

•

Operating shifts per day

•

Hours operated per shift

•

Hours of delay per shift (shift change, service, lunch)

1.08

•

Operating hours per shift

10.92

•

Effective minutes per operating hour

•

Effective hours per shift

9.10

The hours of delay per shift estimate above consists of: 15 minutes for shift change, 45 minutes for lunch/breaks and 10 minutes for blasting delays per day. Staggered breaks for truck drivers and replacement operators on shovels and drills during lunch breaks will be incorporated to attain the planned equipment utilization.

Added to these assumptions in order to determine equipment fleet sizing are: equipment availability and utilization factors, down days due to weather, a reduction in the effective minutes per operating hour, e.g., less than 50 in the initial years, (48 minutes in Year -1 and 49 minutes in Year 1), to reflect operator training.

AMEC comments specific to equipment productivity for each area are given below.

Drilling

AMEC has concerns that adequate explosive power distribution may not be achieved at the toe of the bench in the more competent rock, given the wide blasthole spacing. Poorly fragmented or unlocked material at the toe of the bench will restrict the loader dig rates and lead to increased pit floor remediation work. Blast hole spacings should be confirmed in future updates of the feasibility study.

Loading

For year two of the mine plan, each of the five electric shovels must load 40,300 kt/a, or 110.4 kt/d. The life of mine average production rate is 107.5 kt/d. These production rates are at the very end of AMEC’s experience with the specified electric shovel at large efficient mines, and may require over-trucking to achieve.

Hauling

Operating hours per annum and the average ore and waste truck cycles are based on haul profiles and manufacturer indicated truck performance. Annual ore and waste profiles were developed using the average annual bench elevation by pit stage.

The owner truck fleet, initially at 50 units, increases to 60 in Year 4. The requirement is actually 62, 70, and 64 units in Years 4, 7, and 8 respectively. The owner fleet is kept constant at 60 units by leasing the additional trucks for these high requirement years.

19.2

Metal Recovery

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 $5.25/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 16.2.12 of this report, the most likely element to incur smelting penalties will be mercury.

19.3

Environmental Bonds, Remediation, and Reclamation

According to Flavio Fuentes of Placer Dome Latin America, 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.

19.4

Operating Costs

19.4.1

Summary

Unit operating costs, updated in March 2004, total $4.90/t of ore milled. Costs by area are also shown in Table 19-5.

Table -:

Unit Operating Costs

Area	Cost ($/t)
Mine	1.64
Processing	2.63
Maintenance	0.12
Administration	0.51
Total	4.90

19.4.2

Mine Operating Costs

The March 2004 update to the feasibility study addressed changes in operating and capital costs. The following sections reference the updated costs.

The open pit operating costs were estimated on a yearly basis by determining major and support equipment requirements, including supplies, consumables, and manpower requirements. Cost information was derived from manufacturer’s information or was extrapolated from existing Placer Dome operations.

The following major cost centres were included:

•

operating labour

•

maintenance labour

•

engineering and geology

•

mine operating costs

•

drilling

•

blasting

•

hauling

•

roads and dumps

•

general services

•

pit de-watering.

19.4.3

Processing Plant Operating Costs

Summary

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.

All the consumable costs reported in the final operating cost (OPEX) were revised in the March 2004 Feasibility Study update (PDTS, 2004). The overall processing costs were revised from the 2000 Feasibility Study of $2.62/t to $2.63/t. Incremental costs for electricity and manpower were largely compensated by an expected reduction in the cost of consumables. Recent rises in the cost of grinding media and liners may offset these savings however.

The resulting calculated processing costs distribution, as expected per the variations to major cost posts revealed in the 2004 Feasibility Study update executive summary, are indicated in Table 19-5.

Table -:

Distribution of Processing Operating Costs

Item	Cost ($/t)
Operation labour	0.08
Maintenance labour	0.07
Reagents	0.31
Power	0.80
Grinding Media	1.11
Operating Supplies	0.02
Maintenance Supplies	0.24
Total	2.63

Manpower

The staffing of the plant includes metallurgical staff, operations, and maintenance personnel for the mill, tailings and water systems, filtration plant and concentrate loading. AMEC reviewed staffing in terms of total cost rather than by numbers in each area.

Grinding Media and Liners

The costs for grinding balls and liner wear were derived from empirical equations.

A credit equivalent to 30% of the calculated ball mill ball consumption could have been taken to reflect the better wear rates achieved with the improved steel metallurgy of modern grinding media. Tower mill media wear is routinely fixed at 0.1 kg/kWh, based on experience with similar regrinding duties. Such an assumption is reasonable for a feasibility study.

Reagents

Reagent consumptions were assessed from the G&T locked cycle tests of 2000.

Cyanide destruction reagent costs were based on consumption allowances, based on the treatment of similar slurries elsewhere. They are the least reliable inputs in this category. Actual neutralization test work results should have been obtained at this study stage, providing confirmation as well that no interfering ion in solution would prevent obtaining the targeted cyanide level in the treated slurry.

Electricity

The plant electrical costs were calculated on the basis of a detailed equipment list and assignment of operating load, for each piece of equipment, including plant utilization effect. The resulting costs, based on the design grades and tonnages, were then adjusted for the actual yearly throughputs.

This assessment is sufficiently detailed for this type of study.

Maintenance Supplies

The calculation should have been based on percentages of the indicated capital expenses (CAPEX), per operating area and per type of CAPEX categories (structural, architectural, mechanical, piping, electrical, instrumentation, etc.). This is an acceptable method of evaluating the likely requirements for maintenance parts.

Processing Operating Costs - Comments

The derivation of the operating costs was thoroughly done. AMEC has the following comments:

•

No manpower schedule could be reviewed. Adequacy of staffing level could therefore not be commented

•

A credit for ball mill ball consumption could have been taken to account for the better ball metallurgy that evolved since the derivation of the empirical equations used to assess it.

•

More recent pricing environment for balls and metal liners may have to be taken into consideration just prior to the final evaluation of the project since the volatility of the materials used in their production has been quite high since the last review has been carried out.

19.4.4

Maintenance

19.4.5

General and Administration

G&A costs include personnel, accounting, warehousing, transport of employees, human resources, insurance, and head-office allocations. G&A cost estimates are reasonable.

19.5

Capital Costs

19.5.1

Total Project Capital Costs

Total capital costs by facility are provided in Table 19-7, as referenced in the Section 5 of the March 2004 Feasibility Study Update provided to AMEC.

Table -:

Total Estimated Capital Costs by Facility

	March 2004 US$ (millions)
Direct Costs
Plantsite & roads	38.1
Primary crusher	24.2
Coarse ore stockpile	17.4
Conveying	34.1
Grinding Facilities	195.4
Flotation Facilities	112.8
Leaching Facilities	24.4
Refinery	5.8
Water supply	98.0
Shops & warehouses	20.2
First aid building	0.4
General office	6.4
Assay laboratory	3.5
Portside filter plant	15.6
Open Pit – pre-production stripping & mining equipment	363.5
Power supply	27.1
Tailings disposal	31.7
Concentrate handling	9.3
Cyanide destruction	9.0
Concentrate pipeline	51.9
Accommodations	21.8
Vendors	7.0
Total Direct Costs	1,117.6
Indirect Costs
Construction overheads	83.3
Operations overheads	27.0
Project management	88.4
Design & engineering	75.5
Warehouse inventory	31.3
Freight	40.7
Taxes & duties	12.4
Total Indirect Costs	358.6
Total Construction Cost	1,476.3
Contingency	147.6
Commissioning	25.5
Total Capital Cost	1,649.4

Post-construction sustaining capital costs total $287.9 million and include the following as shown in Table 19-8.

Table -:

Post-Construction Sustaining Capital Costs
(from PDTS, 2004)

Area	Cost $ (millions)
Mining Equipment	160.5
Pit Dewatering & Trolley Assist	17.1
Tailings	22.9
Process Plant & Water System	71.4
Mine Closure	16.0
Total	287.9

19.5.2

Capital Cost Review

AMEC reviewed capital costs for mine facilities and infrastructure using layout drawings, process flowsheets and engineering drawings. Civil, concrete, steel, and piping drawings were not available. AMEC reviewed the estimating methods used by PDTS and compared the totals against projects of similar design and size. Emphasis was given to major capital items and unit prices for each. Cost tables provided by Placer Dome were in a format of unit items (civil works, equipment, steel, platework, etc.) rather than by each major facility as presented in Section 5 of the March 2004 Feasibility Study Update.

Direct Costs

Direct costs include civil works in the pre-production stage, mine equipment, pre-production stripping, sustaining equipment capital, and construction of mine, process, camp, administration facilities and general infrastructure.

Mine Capital items include:

•

pre-production mining and support equipment

•

two years of pre-production stripping.

Process facilities are designed to process a minimum of 150,000 t/d of gold-copper ore and include:

•

Primary Crushing, with two Primary Crushing stations, direct dump, including 60" to 110 " gyratory crushers, and all related facilities.

•

Coarse Ore Stockpile, without cover, with 3 discharge tunnels, including apron feeders and SAG feed conveyors.

•

Conveying System, includes stockpile feed conveyors, SAG feed conveyors, and pebble crushing plant conveyors.

•

Grinding Facilities, include 3 grinding lines that comprises one SAG mill (12.2 m diameter x 6.7 m long) wrap around drive and 2 ball mills (7.3 m diameter x 10.4 m long) wrap around drive with their respective cyclone cluster, and a pebble crushing plant.

•

Flotation Facilities, includes rougher flotation cells (160 m³), three stage cleaning, 6 vertical regrinding mills (6 unit with their cyclone clusters), two tailing thickeners (91 m diameter), one concentrate thickener (30 m diameter, steel tank) with the storage tanks, reagent facilities for lime, frother, collectors, flocculant, and copper sulphate.

•

Leaching facilities, for gold recovery, include leaching (5) and CIP tanks (6 tanks), one cleaner tails (80 m diameter), screens, and related equipment.

•

Refinery facility, to produce “doré” bullion, including carbon regeneration equipment, cyanide mixing and storage facilities, solution tanks, electrowinning cells, furnace, oven and vault.

•

Cyanide Destruction Facility includes sulphur dioxide plant, cyanide destruction reactors/tanks, and sulphate storage stockpile.

•

Port Site Facilities, using the existing Candelaria Port. Port facilities include concentrate reception, storage, filtering and conveying to existing system, and an evaporation pond.

•

Concentrate Pipeline, 246 km from minesite to Candelaria Port, including pump stations, valves stations, storage tanks, and control. Route mainly follows the access road and international road, and along Copiapo Valley.

Infrastructure capital items include:

•

Access road, from existing international road and 20 km to the process plant facilities. Improvement of the existing road is been considered and a new road form construction and operation camps will constructed.

•

Plant site at elevation of 3800 m.

•

Water supply at Piedra Pomez well field located 121 km from to the north of the plant.

•

Power supply facilities consider that the power supplier will provide energy to the plant. One 230 kV transmission line is anticipated to be built. Power supply will include all transformers and yard equipment. In the construction phase power will be supplied by four 1,600 kW generators. An additional four 1,600 kW generators are considered for the mine pre-stripping phase.

•

Shops and warehouse facility includes truck and mobile equipment shops, light vehicles shop, maintenance shop (6,440 m²) and wash facility, including all equipment.

•

Camp facilities – operation camp is designed for 1,232 people and will be located 20 km from the plant site. Access will include a new road.

•

First-aid facilities.

•

General office facilities in Copiapo.

•

Pit and administration building (1,000 m²) and guardhouse.

•

Assay laboratory, including lab equipment and services (770 m²).

Indirect Costs

Indirect costs are those incurred in the construction and start-up phase and include:

•

vendor labour

•

construction facilities

•

construction support

•

construction camp facilities

•

camp catering and operations

•

construction equipment

•

freight

•

insurance, taxes, and duties

•

field office operation

•

project management fees

•

engineering and procurement services

•

engineering fees

•

commissioning

•

spare parts

•

initial fills

•

consumables

•

contractor mobilization and demobilization

•

Owner’s costs

•

corporate office costs.

Capital Estimates

AMEC concludes the following for each capital area:

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. The method used for the engineering development level is adequate.

Equipment for electrical works was estimated using Placer Dome database and local operating experience. The method used is adequate.

Selections and quantities of mining equipment are appropriate for the design of the project and considering the availability of equipment at the time of the 2000 Feasibility Study. Placer Dome intends to revise equipment selections in an update of the feasibility study in 2005.

Indirect Costs

The construction camp is based on a manpower curve showing a peak of 6,600 men and a quotation from “Tecno Fast Atco,” the rest of the camp associated facilities were estimated according to Placer Dome database. This is adequate.

Construction activities were estimated from Placer Dome operations in Chile and are adequate.

Owner’s management staff was estimated with its associated expenses.

Engineering and Procurement Services were estimated by Placer Dome and a Consultant per design package and are appropriate.

Spare Parts, Reagents and Consumables are based on an estimate prepared by the Metallurgical Engineering team in accordance with recent consumptions and prices quoted in March 2004. AMEC is aware of significant increases in prices for consumables in the remainder of 2004.

Ocean freight and inland transportation costs were estimated by the Placer Dome procurement group at 7% of the value of equipment and materials, some allowances were also included for airfreights. This is appropriate.

An average of 2% to cover taxes on imported goods was considered, since imports may arrive from countries with which there is a free trade treaty (with no payment) or from countries without a treaty (these may be of the order of 6%). Taxes for foreign technical services were estimated.

Commissioning cost was estimated considering a 3-month start-up period; it includes the staff, manufacturers’ supervision, construction contractor’s support, equipment and materials.

Comment

In general, the bases for estimating indirect costs are in accordance with customary industry practices. Work quantities are sufficiently detailed to provide the basis for an adequate estimate. AMEC believes that additional costs may be incurred in the following areas and these should be reviewed in the next feasibility study update:

•

AMEC recommends a price of $1.50/m³ for mass backfill, which may add $1 million to backfill in several sectors

•

AMEC recommends an average price of concrete no lower than $650/m³, which may add $11 million to facility costs

•

Mechanical platework should use a unit cost $3,500/t, for an increase of $3 million.

19.5.3

Sustaining Capital Cost Review

AMEC reviewed estimates for sustaining capital required over the life of the mine. These consist of mining and process equipment replacements, pit dewatering, installation of trolley assist for haul trucks and mine closure. Equipment selections and qualities appear reasonable relative to AMEC’s experience with similar scale projects.

19.6

Economic Analysis

Discounted cash flow analysis indicates that the project offers a positive return. Payback period is 9.4 years. Life-of-mine is 18 years.

The model does not appear to include an allocation for working capital; however, when standard estimates are used for working capital, there is only a small impact on return and the internal rate of return is still positive.

All other inputs are appropriate and, apart from the first few years of development, all future annual cash flows are 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 AMEC’s opinion, the level of detail used in the economic analysis is appropriate for a feasibility study.


Project No.: 148562	Page #
June 2005

20.0

INTERPRETATIONS AND CONCLUSIONS

20.1

Technical Basis for Resources and Reserves

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. AMEC reviewed the March 2004 capital and operation costs and provides its 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.

Increases in capital and operating costs occurred through the remainder of 2004. These increases will impact the project, but the degree to which cannot be determined until the technical and cost aspects of the project are again updated in 2005.

20.2

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.

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.

20.3

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.

20.4

Drilling Programs

A total of 224 RC and 124 core holes totalling 122,747 m support resource estimates for Cerro Casale. 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 below 200 m. Core holes are from 30 to 1,473 m deep. Drilling tools produced NC (61 mm), HQ (61 mm), NQ (45 mm) and HX (63 mm) cores. Core recovery is poorly documented but appears to have exceeded 95%.

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.

The quality of logging of attributes related to mineralization is adequate to support resource estimates.

20.5

Sampling, Sample Preparation, and Assaying

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.

20.6

Assay QA/QC

Assay QA/QC protocols were observed throughout all drilling campaigns, with blind standard reference materials (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 Placer Dome Research Centre, 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.

20.7

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.

20.8

Data Verification

The assay and geological databases are suitable to support resource estimates.

20.9

Geological Interpretations

20.10

Metallurgical Processing

The metallurgical test work has been fairly extensive. The plant design has relied on the grinding testwork completed in 1997, the pilot plant work of June 1999 for dewatering circuits and on the G&T flotation and cyanidation testwork in 1999. The Feasibility Study Report was updated in March 2004 in order to reflect current costs for the capital and operating costs determination.

Metallurgical recovery equations for gold and copper were developed for eight ore types. There was good agreement between the Placer Dome recovery models and actual locked cycle test results.

Plant designs are reasonable. The resultant sizing of individual equipment, from the application of the adopted design criteria, was not completely confirmed during AMEC’s review, although AMEC verified the testwork parameters and procedures applied to achieve the scale-up were assessed and found these to be following standard practices. The design criteria that could be reviewed in more detail during the course of this review were found to be sufficiently conservative relative to the interpretation of the available metallurgical data upon which they were based. Some of the equipment sizing should be considering peak throughputs that would result from processing softer ore blends

Metallurgical testwork, interpretations of metal recoveries and ore characteristics and resulting plant designs were done in sufficient detail and are reliable enough to support reserve estimates.

20.11

Mineral Resource and Mineral Reserve Estimates

20.11.1

Resource Estimates

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.

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.

Modelling for gold and copper grades consisted of grade interpolation by ordinary kriging (OK). Only capped grades were interpolated. Nearest-neighbour (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-neighbour estimates on spatial swaths through the deposit. These checks showed no biases or local artefacts due to the estimation procedures.

20.11.2

Resource Classification

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.

The mineral resources of the Cerro Casale project were tabulated based on material that fell within an optimistic ultimate pit shell. That shell was made based on a gold price of US$500/oz, US$1.25/lb for copper, and estimated costs (from experience on similar mines) of US$3.39/t. This shell fulfilled the expectation of reasonable extraction test in declaring mineral resources at Cerro Casale. AMEC agrees with this logic and its implementation.

Mineral resource estimates and classification are compliant with Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Standards on Mineral Resources and Mineral Reserves (2000) and Canadian National Instrument 43-101.

20.11.3

Mineral Reserves

Mineral reserves were estimated in the 2000 Feasibility Study, using an elevated cutoff grade strategy for the four phases of mining and use of 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.

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.

Mineral reserve estimates are compliant with Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000) and Canadian National Instrument 43-101.

20.12

Mining Designs and Production Plans

Under the 2000 Feasibility Study, all planned mining of the Cerro Casale deposit will be conducted by conventional open pit methods. The mine plan features a four stage open pit which is scheduled to deliver a nominal 150,000 t/d of mill feed at a maximum mining rate of 816,000 t/d over an 18 year mine life. The final pit stage will measure over 2,100 m in diameter and the highest sector of highwall will have a vertical extent of 1,224 m, ranking the proposed final highwall amongst the world’s tallest.

PDTS 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. An average gold price of $350/oz, a copper price of $0.95/lb and an exchange rate of US$1 per 510 Chilean Pesos were used in the 1999 Profit Model for pit designs. An exchange rate of US$1 per 600 Chilean Pesos was used for operating and capital cost estimate updates in 2004. 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.

The simplification of the inter-ramp slope recommendations is reasonable for pit optimization and acceptable for feasibility level pit design.

Mine plans are suitable to support mineral reserve estimates and are compliant with Canadian Institute of Mining, Metallurgy, and Petroleum (CIM) Standards on Mineral Resources and Reserves (2000) and Canadian National Instrument 43-101.

20.13

Operating Costs

Total Operating Costs

Operating costs total $4.90/t of ore and were updated in March 2004. Major cost centres are mining, processing, maintenance, and administration. All appropriate costs are included.

Additional inflation of consumable costs occurred industry-wide in the remainder of 2004. These will be reconsidered in a future feasibility study update.

Mining

The open pit operating costs total $1.64/ore tonne and have been estimated on a yearly basis by determining major and support equipment requirements, including supplies, consumables and manpower requirements. Cost information was derived from manufacturer’s information and extrapolated from existing Placer Dome operations in Chile are therefore are suitable.

Processing

All the consumable costs reported in the final operating cost (OPEX) were revised in the March 2004 Feasibility Study update (PDTS, 2004).

20.14

Capital Costs

Total estimated capital cost is $1,649.4.0 million. Post-construction sustaining capital costs total $287.9 million and include mining equipment, pit dewatering and trolley assist, tailings construction, process plant and water system equipment and mine closure. AMEC reviewed capital estimates and found them to be appropriate.

20.15

Economic Analysis

Economic analysis of the Cerro Casale project is based upon a discounted cash flow analysis on a post-tax basis, using Proven and Probable Mineral Reserves and annual production plans in the 2000 Feasibility Study. Projections for annual revenues and costs are based on data developed for the mine, process plant, capital expenditures and operating costs. Capital and operating costs used were from the March 2004 Feasibility Study Update.

Discounted cash flow analysis indicates that the project offers a positive return. Payback period is 9.4 years. Life–of-mine is 18 years.

All other inputs are appropriate and, apart from the first few years of development, all future annual cash flows are 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 AMEC’s opinion, the level of detail used in the economic analysis is appropriate for a feasibility study and appropriately reflects the value of the project.

21.0

RECOMMENDATIONS

AMEC recommends that the feasibility study be updated for changes in capital and operating costs since March 2004. Technical aspects of the feasibility study, such as mining equipment selections, have not been updated since 2000 and may represent areas of opportunity to improve costs.

AMEC has the following additional recommendations:

•

Energy supply contracts should be renegotiated.

•

Further studies on the acid-rock drainage potential of waste rock should be initiated to better determine if design changes in the waste rock facility are warranted.

•

Coarse rejects for samples in batches 135 to 234 from the 1999 drilling campaign should be pulled from storage and reassayed for the next model update. Re-assays can be restricted to those samples in holes 99GT003-006 and CCD111.

•

Silver recoveries in the future should be based on actual metallurgical tests.

•

Drop weight tests should be performed on large drill core to determine if SAG mill capacity may not be as constrained as presently thought. A reduction in capital costs might be achieved if drop tests were influenced by small core.

•

The refinery equipment may require the addition of a mercury retort furnace, considering the high mercury content found in some of the flotation concentrate.

•

AMEC would recommend that in future work Placer Dome require that multiple holes located within the respective search ellipse will be used in estimating Measured and Indicated mineral resources rather than the presently used indirect method.

•

For the next feasibility update, AMEC recommends the construction of a geotechnical block model, which will allow the generation of pit designs that will more closely honour the geotechnical engineer’s inter-ramp angle, bench face angle and berm width recommendations.

•

Unbroken inter-ramp slopes in excess of 350 m vertical height should be avoided in preliminary mine plans.

•

AMEC recommends using a standard 15 m bench height as the basis for feasibility level planning and suggests that with operational experience, upside potential exists to mine thicker benches.

•

These loading rates assumed for electric shovels are at the very high end of AMEC’s experience with the selected electric shovels at large efficient mines. These rates should be reviewed in the next study update.

•

An updated economic model should include working capital.

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.

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.

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.

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.

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.

Piteau Associates, 1999, Aldebarán Project, Cerro Casale Sulphide Deposit, Feasibility Geotechnical Assessments for the Open Pit, private report prepared for Compañía Minera Aldebarán.

PDTS, 2000, Aldebarán Project, Chile: Feasibility Study, private report prepared for Compañía Minera Aldebarán.

PDTS, 2004, Aldebarán Project, Chile: Feasibility Study Update, private report prepared for Compañía Minera Aldebarán, March 2004.

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.

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

APPENDIX A

LIST OF SIGNIFICANT ASSAYS


Project No.: 148562	Appendix A
June 2005