Technical Report Dated October 26, 2005

Method

Number of Samples per Run

Quality Control Sample Allocation

Fire Assay

2 standards, 3 duplicates, 1 blank

Other Methods

2 standards, 1 duplicate, 1 blank

SECURITIES AND EXCHANGE COMMISSION

Washington, D.C. 20549

FORM 6 - K

Report of Foreign Private Issuer
Pursuant to Rule 13a-16 or 15d-16 of
the Securities Exchange Act of 1934

For the month of October, 2005

PLACER DOME INC.

Suite 1600, 1055 Dunsmuir Street
P.O. Box 49330, Bentall Postal Station
Vancouver, British Columbia
Canada V7X 1P1

Indicate by check mark whether the registrant files or will file annual reports under cover Form 20-F or Form 40-F.

Form 20-F o Form 40-F þ

Indicate by check mark whether the registrant by furnishing the information contained in this Form is also thereby furnishing the information to the Commission pursuant to Rule 12g3-2(b) under the Securities Exchange Act of 1934.

Yes o No þ

SIGNATURES

Pursuant to the requirements of the Securities Exchange Act of 1934, the registrant has duly caused this report to be signed on its behalf by the undersigned, thereunto duly authorized.


	PLACER DOME INC. (Registrant)
	By:	/s/ Geoffrey P. Gold
		Geoffrey P. Gold
		Vice-President, Assistant Secretary and Associate General Counsel

Date: October 28, 2005

Pueblo Viejo Project Province of Sanchez Ramirez, Dominican Republic 43-101 Technical Report and Quali.ed Person’s Review Prepared for: Placer Dome Inc. By: Susan Lomas P.Geo. Effective Date: 26 October 2005 Project Number: 147107

IMPORTANT NOTICE

Recognizing that Placer Dome Inc. (PDI) has legal and regulatory obligations in a number of global jurisdictions, AMEC Americas Limited (AMEC) consents (a) to the filing of this report with any stock exchange and other regulatory authority and any publication of this report by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, and (b) to the publication of this report and/or excerpts thereof and/o information therein by PDI on its company website or otherwise, and (c) to all other uses by PDI of this report and/or excerpts thereof and/or information therein in connection with its business.

This report was prepared as a National Instrument 43-101 Technical Report, in accordance with Form 43-101F1, for PDI by 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 PDI, subject to the terms and conditions of its contract with AMEC. That contract permits PDI 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.

Susan F. Lomas, P.Geo.
111 Dunsmuir Street, Suite 400
Vancouver, BC
Tel.: (604) 664-3413
Fax: (604) 664-3057
susan.lomas@amec.com

CERTIFICATE OF AUTHOR

I, Susan F. Lomas, P.Geo., am a Professional Geoscientist, employed as Principal Geologist of AMEC Americas Limited and residing at 2620 Lauralynn Drive in the City of North Vancouver in the Province of British Columbia.

I am a member of the Association of Professional Engineers and Geoscientists of British Columbia. I graduated from the Concordia University of Montreal with a Bachelor of Science degree in geology in 1987.

I have practiced my profession continuously since 1987 and have been involved in : mineral exploration for gold, nickel, copper, zinc, and silver, in Canada, United States, Mexico, Venezuela and Ghana and in underground mine geology, ore control and resource modelling for gold, nickel, copper, zinc, silver, potash and industrial mineral properties in Canada, United States, Venezuela, Peru, Ecuador, Brazil, Dominican Republic, Thailand, China, Mongolia, New Caledonia, Ghana, and Russia.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.

I am currently a Consulting Geologist and have been so since February 1998.

I am responsible for the preparation of the technical report titled Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic and dated 26 October 2005 (the “Technical Report”). I prepared Sections 1 through 11, 15, 20 and 21. of the Technical Report. I prepared Sections 19.3 and 19.4 will the assistance of Mr. Ron Pearce, Chief Estimator for AMEC and Mr. Simon Handelsman, Financial Consultant for AMEC. Between 10 and 14 January 2005 I visited the Pueblo Viejo property and reviewed exploration programs, exploration data, geological models and historical records for the purpose of preparing a technical report on all mineral resources and reserves within the project. I was assisted in the site review of resource estimations, mining, pit designs and reserve estimation by Mr. Tracy E. Barnes, P.Eng., a Qualified Person in the areas of resource estimation, reserve estimation and mining engineering. Mr. Barnes prepared Sections 17 and 19.1 of the Technical Report. Dr. Ted Eggleston, R.Geo, a Qualified Person in the area of exploration data reviews, sampling, assaying and assay quality assurance, reviewed sampling information and assay quality assurance and quality control information at Placer Dome Inc. offices in Vancouver, British Columbia. Dr. Eggleston prepared Sections 12, 13 and 14 of the Technical Report. Mr. Lynton Gormely, P.Eng, a Qualified Person in the area of metallurgy and process, assisted with an office review of metallurgical tests, process designs and process operating costs. He prepared Sections 16 and 19.2. I was further assisted in review of the capital costs by Mr. Ronald Pearce, AMEC Chief Estimator, review of financial analysis by Mr. Simon Handelsman, Finance Consultant, review of environmental conditions and permitting by Mr. Dennis Krochak and review of twin hole data analyses by Mr. Kevin Francis, R.Geo. The resulting Technical Report was prepared under my supervision.

I have had no other prior involvement with the property that is the subject of the Technical Report.


AMEC Americas Limited
111 Dunsmuir Street, Suite 400
Vancouver, B.C. V6B 5W3
Tel	(604) 664-3471
Fax	(604) 669-3057	www.amec.com

Dr. Ted Eggleston is a Registered Profession Geologist, employed as an independent consulting geologist and residing at 011 Flying Eagle Trail, Cotapaxi, Colorado 81223. He graduated from Western State College in Colorado in 1975 with Bachelor of Arts Degree in Geology. He graduated from the New Mexico Institute of Mining and Technology with a Master of Science Degree in Geology in 1982 and a Ph.D. in geology in 1986. Dr. Eggleston is a Registered Professional Geologist in the State of Wyoming, USA (PG1830). Since 1975 Mr. Eggleston has continually been involved in mineral exploration projects for precious and base metals and industrial minerals in the United States, Canada, Chile, Perú, Argentina, Ecuador, Colombia, Brazil, Costa Rica, Mexico, Myanmar, Russia, Zambia, Ghana, Norway, and Turkey.

Tracy E Barnes is a Registered Professional Engineer in the State of Colorado, Registration No. 33381, and is employed by Barnes Engineering Services, Inc as President and Principal Mining Engineer, and residing at 12945 W 15^th Drive in the City of Golden in the State of Colorado, USA. He is a member of the Society for Mining, Metallurgy, and Exploration, the American Statistical Association, and the International Association for Mathematical Geology. He graduated from the University of Washington, Seattle, Washington with a Bachelor of Science in Mining Engineering degree in 1975. Mr. Barnes has practiced his profession continuously since 1975 and has been involved in: gold operations in the United States, New Zealand and Chile and preparation of mine audits, scoping, pre-feasibility, and feasibility level studies for gold properties in United States, Canada, Chile, Peru, New Zealand, and Brazil.

Lynton Gormely is a Registered Professional Engineer in the Province of British Columbia, Registration No. 10005, and is employed by AMEC Americas Ltd as a Principal Process Engineer, and residing at 5441 Cliffridge Avenue, North Vancouver, B.C. He is a member of the Canadian Institute of Mining Metallurgy and Petroleum. He graduated from the University of British Columbia, Vancouver, B.C., with a Doctor of Philosophy in Chemical Engineering degree in 1973. Mr. Gormely has practiced his profession continuously since 1973 and have been involved in test work supervision, preparation of process audits, scoping, pre-feasibility, and feasibility level studies for gold properties in Canada, the United States, Australia, Papua New Guinea, Mongolia, and China.

I am not aware of any material fact or material change with respect to the subject matter of the Technical Report, which is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

I, Mr. Tracy Barnes, Dr. Ted Eggleston and Mr. Lynton Gormely are independent of Placer Dome Inc. in accordance with the application of Section 1.5 of National Instrument 43-101.

I have read National Instrument 43-101 and certify that the Technical Report has been prepared in compliance with that Instrument. I further certify that, as of the date of this certificate, the Technical Report contains all of the information required under Form 43-101F1 in respect of the property that is the subject of the report.

Dated at Vancouver, British Columbia, this 26^th day of October 2005.

Signature on File


Susan Lomas, P.Geo.

Page 2 of 2

Susan F. Lomas, P.Geo.
111 Dunsmuir Street, Suite 400
Vancouver, BC
Tel.: (604) 664-3413
Fax: (604) 664-3057
susan.lomas@amec.com

CONSENT of AUTHOR


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

AND TO: Placer Dome Inc.

I, Susan F. Lomas, P.Geo., do hereby consent to the filing of the technical report prepared for Placer Dome Inc. titled Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic and dated 26 October 2005 (the “Technical Report”) with the securities regulatory authorities referred to above.

I further consent (a) to the filing of the Technical Report with any stock exchange and other regulatory authority and any publication of the Technical Report by them for regulatory purposes, including electronic publication in the public company files on their websites accessible by the public, and (b) to the publication of the Technical Report and/or excerpts thereof and/or information therein by Placer Dome Inc. on its company website or otherwise, and (c) to all other uses by Placer Dome Inc. of the Technical Report and/or excerpts thereof and/or information therein in connection with its business, and (d) to the use of my name as a Qualified Person in connection with the Technical Report in any continuous disclosure documents or other regulatory documents or filings of Placer Dome Inc., in all applicable jurisdictions.

I certify that I am not aware of any material fact or material change with respect to the subject matter of the Technical Report, which is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

I confirm that I have read the Placer Dome Inc. news release dated September 27, 2005 entitled “Placer Dome Approves Pueblo Viejo Project and Provides Cerro Casale Update” and that its contents are consistent with the information in the Technical Report.

Dated this 26^th of October 2005.


Signature on File

Susan Lomas, P.Geo.


AMEC Americas Limited
111 Dunsmuir Street, Suite 400
Vancouver, B.C. V6B 5W3
Tel	(604) 664-4315
Fax	(604) 664-3057	www.amec.com

Lynton S. Gormely, Ph.D., P.Eng.
400 – 111 Dunsmuir Street
Vancouver, B.C., V6B 5W3
Tel: (604) 664 3312
Fax: (604) 664 3057
lynton.gormely@amec.com

CERTIFICATE OF QUALIFIED PERSON

I, Lynton S. Gormely am a Registered Professional Engineer in the Province of British Columbia, Registration No. 10005, I am employed by Amec Mining & Metals as a Principal Process Engineer, and I reside at 5441 Cliffridge Avenue, North Vancouver, B.C.

I am a member of the Canadian Institute of Mining Metallurgy and Petroleum. I graduated from the University of British Columbia, Vancouver, B.C., with a Doctor of Philosophy in Chemical Engineering degree in 1973.

I have practiced my profession continuously since 1973 and have been involved in: test work supervision, preparation of process audits, scoping, pre-feasibility, and feasibility level studies for gold properties in Canada, the United States, Australia, Papua New Guinea, Mongolia, and China.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.

I am currently a Consulting Engineer and have been so since 1974.

I served as the Qualified Person writing sections 16, 19.2, and 19.3 of the technical report titled Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic and dated 26 October 2005 (the “Technical Report”), relating to metallurgical test work, process design, and process operating costs. The work was completed in AMEC’s office and this work did not require me to visit the Pueblo Viejo site.

I have had no other prior involvement with the property that is the subject of the Technical Report.

I am not aware of any material fact or material change with respect to the subject matter of the Technical Report which is not reflected in the Technical Report, the omission to disclose which makes the Technical Report misleading.

I am independent of Placer Dome Inc. in accordance with the application of Section 1.5 of National Instrument 43-101.

I have read National Instrument 43-101 and certify that the portions of the Technical Report for which I served as a Qualified Person have been prepared in compliance with that Instrument. I further certify that, as of the date of this certificate, the portions of the Technical Report for which I served as a Qualified Person contain all of the information required under Form 43-101F1 in respect of the property that is the subject of the report.

Dated at Vancouver, British Columbia, Canada, this 26^th day of October 2005.


Signature on File

Lynton S. Gormely, Ph.D., P.Eng.


AMEC Americas Limited
111 Dunsmuir Street, Suite 400
Vancouver, B.C. V6B 5W3
Tel	(604) 664-5311
Fax	(604) 669-3057	www.amec.com

Lynton S. Gormely, Ph.D., P.Eng.
400 – 111 Dunsmuir Street
Vancouver, B.C., V6B 5W3
Tel: (604) 664 3312
Fax: (604) 664 3057
lynton.gormely@amec.com

CONSENT of AUTHOR


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

AND TO: Placer Dome Inc.

I, Lynton S. Gormely, do hereby consent to the filing of the technical report prepared for Placer Dome Inc. titled Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic and dated 26 October 2005 (the “Technical Report”) with the securities regulatory authorities referred to above.

Dated this 26^th day of October 2005.


Signature on File

Lynton S. Gormely, Ph.D., P.Eng.


AMEC Americas Limited
111 Dunsmuir Street, Suite 400
Vancouver, B.C. V6B 5W3
Tel	(604) 664-5311
Fax	(604) 669-3057	www.amec.com

Ted Eggleston, Ph.D., P.Geo
P.O. Box 67, Hillside, CO 81232
Tel (719) 942-4418; Fax (719) 942-3009
egglesto@ris.net

CERTIFICATE OF AUTHOR

I, Ted Eggleston Ph.D., P. Geo, am an Associate Geologist for AMEC Americas (formerly AMEC E&C Services) Limited, and residing at 011 Flying Eagle Trail, Cotapaxi, Colorado 81223.

I graduated from Western State College in Colorado in 1975 with Batchelor of Arts Degree in Geology. I graduated from the New Mexico Institute of Mining and Technology with a Master of Science Degree in Geology in 1982 and a Ph.D. in geology in 1986. I am a Registered Professional Geologist in the State of Wyoming, USA (PG1830).

Since 1975 I have continually been involved in mineral exploration projects for precious and base metals and industrial minerals in the United States, Canada, Chile, Perú, Argentina, Ecuador, Colombia, Brazil, Costa Rica, Mexico, Myanmar, Russia, Zambia, Ghana, Norway, and Turkey.

I have read the definition of “Qualified Person” set out in National Instrument 43-101 (“NI 43-101”) and certify that by reason of my education, affiliation with a professional association (as defined in NI 43-101) and past relevant work experience, I fulfill the requirements to be a “qualified person” for the purposes of NI 43-101.

I reviewed information on sampling, assaying and assay quality assurance and quality control information for drilling on the Pueblo Viejo Project and historical data on ore and waste densities. This work was undertaken at Placer Dome’s offices in Vancouver, B.C. on December 2 and 3, 2004. I am responsible for the preparation of Sections 12, 13 and 14 in the Technical Report titled “Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic ” and dated 26 October 2005 (the “Technical Report”). I have no prior involvement with the Property that is the subject of this Report.

I am not aware of any material fact or material change with respect to the subject matter of this technical report that is not reflected in this report, the omission to disclose which would make this report misleading.

I am independent of Placer Dome Inc. in accordance with the application of Section 1.5 of National Instrument 43-101.

Dated at Hillside, Colorado, USA, this 26^th day of October 2005.


Signature on File

Ted Eggleston, Ph.D., P. Geo.

Ted Eggleston, Ph.D., P.Geo
P.O. Box 67, Hillside, CO 81232
Tel (719) 942-4418; Fax (719) 942-3009
Email: egglesto@ris.net
CONSENT of AUTHOR


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

AND TO: Placer Dome Inc.

I, Ted Eggleston Ph.D., P. Geo, do hereby consent to the filing of the technical report prepared for Placer Dome Inc. titled Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic and dated 26 October 2005 (the “Technical Report”) with the securities regulatory authorities referred to above.

Dated this 26^th day of October 2005.

Signature on File

Ted Eggleston, Ph.D., P. Geo.

12945 West 15^th Drive, Golden, Colorado 80401-3572 USA

Tracy E Barnes, P.E.
12945 W 15^th Drive
Golden, Colorado, USA
Tel: (303) 202-3667
Fax: (303) 202-3696
tbarnes@barnesengr.com

CERTIFICATE OF QUALIFIED PERSON

I, Tracy E Barnes am a Registered Professional Engineer in the State of Colorado, Registration No. 33381, employed by Barnes Engineering Services, Inc as President and Principal Mining Engineer, and residing at 12945 W 15^th Drive, Golden, Colorado, USA.

I am a member of the Society for Mining, Metallurgy, and Exploration, Inc., the American Statistical Association, and the International Association for Mathematical Geology. I graduated from the University of Washington, Seattle, Washington with a Bachelor of Science in Mining Engineering degree in 1975.

I have practiced my profession continuously since 1975 and have been involved in: gold operations in the United States, New Zealand and Chile and preparation of mine audits, scoping, pre-feasibility, and feasibility level studies for gold properties in United States, Canada, Chile, Peru, New Zealand, and Brazil.

As a result of my experience and qualifications, I am a Qualified Person as defined in National Instrument 43-101.

I am currently a Consulting Engineer and have been so since May 1996.

I served as the Qualified Person writing sections 17 and 19.1 of the technical report titled Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic and dated 26 October 2005 (the “Technical Report”), relating to mineral resource and reserve estimates, mine design, mine planning, and mining operations. I visited the Pueblo Viejo site between 10 and 14 January 2005.

I have had no other prior involvement with the property that is the subject of the Technical Report.

I am independent of Placer Dome Inc. in accordance with the application of Section 1.5 of National Instrument 43-101.



Voice: (303) 202-3667	Facsimile: (303) 202-3696	e-mail: tbarnes@barnesengr.com

Dated at Golden, Colorado, United States of America, this 26^th day of October 2005.


Signature on File

Tracy E. Barnes, P.E.

12945 West 15^th Drive, Golden, Colorado 80401-3572 USA

Tracy E.Barnes,P.E.
12945 W 15^th Drive
Golden, Colorado, USA
Tel: (303) 202-3667
Fax: (303) 202-3696
tbarnes@barnesengr.com

CONSENT of AUTHOR


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

AND TO: Placer Dome Inc.

I, Tracy E. Barnes, do hereby consent to the filing of the technical report prepared for Placer Dome Inc. titled Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic and dated 26 October 2005 (the “Technical Report”) with the securities regulatory authorities referred to above.



Voice: (303) 202-3667	Facsimile: (303) 202-3696	e-mail: tbarnes@barnesengr.com

12945 West 15^th Drive, Golden, Colorado 80401-3572 USA

Dated this 26^th day of October 2005.


Signature on File

Tracy E. Barnes



Voice: (303) 202-3667	Facsimile: (303) 202-3696	e-mail: tbarnes@barnesengr.com

Contents


1.0 SUMMARY			1-1
1.1 Project Description			1-1
1.2 Special Lease Agreement			1-2
1.3 Environmental Review			1-2
1.3.1 Assessment of Existing Conditions			1-2
1.3.2 Future Liabilities and Risks			1-2
1.4 Property History			1-3
1.5 Geological Setting and Mineralization			1-4
1.6 Drilling and Verification of Data			1-5
1.6.1 Rosario Drilling			1-5
1.6.2 GENEL Drilling			1-6
1.6.3 MIM Drilling			1-6
1.6.4 PDDC Drilling			1-6
1.6.5 Twinned-Hole Comparison			1-6
1.7 Sampling Method and Approach			1-6
1.7.1 Pre-PDDC Drilling Programs			1-6
1.8 Sample Preparation, Analyses, and Security			1-7
1.9 Density			1-7
1.10 Mineral Processing and Metallurgical Testing			1-8
1.11 Mineral Resource Estimates			1-8
1.12 Mining Operations			1-9
1.13 Mineral Resource and Mineral Reserve Statements			1-10
1.14 Mill Operating Cost			1-10
1.15 Capital and Operating Cost Estimates			1-11
1.16 Economic Analysis			1-11

2.0 INTRODUCTION AND TERMS OF REFERENCE			2-1
2.1 Terms of Reference			2-1

3.0 DISCLAIMER			3-1

4.0 PROPERTY DESCRIPTION AND LOCATION			4-1
4.1 Project Location			4-1
4.2 Special Lease Agreement			4-1
4.3 Environmental Review			4-6
4.3.1 Information Sources			4-6
4.3.2 Assessment of Existing Conditions			4-6
4.4 Liabilities and Risks			4-7
4.4.1 Pre-operation			4-8
4.4.2 Operations			4-9
4.4.3 Closure and Post Closure			4-11
4.5 Conclusions			4-12
4.6 Permits			4-12

5.0 ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY			5-1
5.1 Climate and Physiography			5-1
5.2 Access			5-1

6.0 HISTORY			6-1



Project No.: 147107		TOC i


6.1 Pre-1969			6-1
6.2 Rosario/AMAX (1969-1992)			6-1
6.3 Privatization (1996)			6-3
6.3.1 GENEL JV			6-4
6.3.2 MIM			6-7
6.3.3 Newmont			6-8

7.0 GEOLOGICAL SETTING			7-1
7.1 Regional Geology			7-1
7.2 Local Geology – Moore and Monte Negro Deposits			7-4
7.2.1 Lithology and Stratigraphy			7-4
7.2.2 Laboratory Studies			7-7

8.0 DEPOSIT TYPES			8-1

9.0 MINERALIZATION			9-1
9.1 General Description			9-1
9.2 Metal Occurrence and Distribution			9-2
9.2.1 Gold			9-2
9.2.2 Silver			9-4
9.2.3 Zinc			9-4
9.2.4 Copper			9-5
9.2.5 Lead			9-5
9.3 Mineralization			9-5
9.3.1 Moore			9-5
9.3.2 Monte Negro			9-9
9.4 Hydrothermal Alteration			9-12
9.5 Weathering			9-13
9.6 Ore Controls Used in Resource Estimates			9-13

10.0 EXPLORATION			10-1
10.1 Structural Pit Mapping			10-1
10.2 Regional Mapping and Sampling			10-1
10.3 Tailings Evaluation			10-1

11.0 DRILLING			11-1
11.1 All Drilling Campaigns			11-1
11.2 Rosario Drilling			11-1
11.3 GENEL Drilling			11-2
11.4 MIM Drilling			11-3
11.4.1 Pre-Placer Dome Drill Hole Surveying			11-3
11.5 PDDC Drilling			11-3
11.5.1 Summary			11-3
11.5.2 Surveying			11-4
11.5.3 Geological Logging			11-7
11.6 Discussion of Drilling Programs			11-9
11.6.1 Rosario Pseudo-Twin Assay Pairing			11-9
11.6.2 Cross Section Review			11-10
11.6.3 Conclusions			11-10

12.0 SAMPLING METHOD AND APPROACH			12-1
12.1 Pre-PDDC Drilling Programs			12-1
12.2 PDDC Diamond Drilling			12-2



Project No.: 147107		TOC ii


13.0 SAMPLE PREPARATION, ANALYSES, AND SECURITY			13-1
13.1 Summary			13-1
13.2 Sample Preparation Procedures			13-2
13.2.1 Rosario			13-2
13.2.2 GENEL			13-2
13.2.3 MIM			13-3
13.2.4 PDDC			13-3
13.3 QA/QC Procedures			13-5
13.3.1 PDDC 2002			13-5
13.3.2 ALS Chemex Quality Control			13-6
13.3.3 Acme Check Assay Program			13-7
13.3.4 GENEL QA/QC – 1996 to 1997			13-7
13.3.5 1985 Rosario Check Assays			13-8
13.3.6 1978 Rosario Check Assays			13-8
13.4 Discussion of Sampling, Sample Preparation and QA/QC at Pueblo Viejo			13-9

14.0 DATA VERIFICATION			14-11
14.1 Summary			14-11
14.2 Verification of Pre-PDDC Data			14-12
14.2.1 Database Development			14-12
14.2.2 Comparison of PDDC, GENEL, and Rosario Drill Hole Assays			14-12
14.2.3 Cross Sectional Review of MIM, Rosario, and PDDC Drilling			14-13
14.3 Verification of PDDC Data			14-13
14.4 Historical Twinned-Hole Comparisons			14-13
14.5 Decay and Cyclicity (Down-hole Contamination) of RC Holes			14-16
14.5.1 Introduction			14-16
14.5.2 Conclusions			14-17
14.6 Density			14-18

15.0 ADJACENT PROPERTIES			15-1

16.0 MINERAL PROCESSING AND METALLURGICAL TESTING			16-1
16.1 Introduction			16-1
16.2 Metallurgical Studies Prior to Placer Dome			16-1
16.3 Placer Dome Metallurgical Testwork			16-3
16.3.1 Introduction			16-3
16.3.2 Whole Ore Pressure Oxidation			16-4
16.3.3 Limestone Calcining and Slaking Testwork			16-5

17.0 MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES			17-6
17.1 Summary of Conclusions			17-6
17.2 Methodology			17-6
17.2.1 Geological Modelling			17-8
17.2.2 Data Compositing			17-11
17.2.3 Data Analysis			17-11
17.3 Variogram Analysis			17-12
17.4 Interpolation Plans			17-12
17.4.1 Moore			17-12
17.4.2 Monte Negro			17-13
17.5 Model Validation			17-14
17.6 Resource Classification			17-15
17.7 Mineral Resource Statements			17-19
17.8 Mineral Reserves			17-19



Project No.: 147107		TOC iii


18.0 OTHER RELEVANT DATA AND INFORMATION			18-1

19.0 REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT AND PRODUCTION PROPERTIES			19-2
19.1 Mining Operations			19-2
19.1.1 Introduction			19-2
19.1.2 Mine Operations Summary			19-3
19.1.3 Mine Design and Mine Plan			19-6
19.1.4 Mining Operations			19-8
19.2 Process Metal Recoveries			19-10
19.2.1 Overview			19-10
19.2.2 Process Description			19-11
19.3 Markets			19-22
19.4 Contracts			19-22
19.5 Environmental			19-23
19.6 Taxes			19-24
19.7 Capital Cost Estimates			19-24
19.7.1 Mill Operating Cost			19-25
19.8 Economic Analysis			19-26

20.0 CONCLUSIONS			20-1
20.1 Introduction			20-1
20.2 Conclusions			20-1

21.0 REFERENCES			21-1

Tables


Table 1-1:	Proven and Probable Mineral Reserves, 27 September 2005 at $400/oz Au Price	1-10
Table 1-2:	Mineral Resources, 27 September 2005 at $450/oz Au Price	1-10
Table 6-1:	Historical Geological Reserve, Fluor 1985 (3.0 Au g/t cutoff)	6-2
Table 6-2:	Historical Geological Resource Results, SW/AMS 1992	6-3
Table 6-3:	Historical Resource Estimate, GENEL (Rosario) 1997 / PAH 2001 (2.0 g/t Au cutoff grade)	6-6
Table 6-4:	Historical Measured + Indicated Resource Estimate, El Dorado Pit, MRDI 1998	6-7
Table 9-1:	Mineralogically Determined Deportment of Gold	9-3
Table 11-1:	Drill Hole Dip Angles by Deposit	11-1
Table 11-2:	Rosario Drill Hole Summary	11-4
Table 11-3:	GENEL Drill Hole Summary	11-2
Table 11-4:	MIM Drill Hole Summary	11-3
Table 12-1:	Sample Interval Data for Rosario, GENEL and MIM Drill Holes	12-1
Table 12-2:	Sample Length and Grade Comparison Results	12-2
Table 13-1:	Summary of Pueblo Viejo Assaying Procedures	13-5
Table 13-2:	ALS Quality Control Sample Frequency	13-7
Table 14-1:	List of Twin Holes	14-15
Table 14-2:	Intervals with Possible Down-hole Contamination	14-18
Table 16-1:	Summary of Metallurgical Test Programs (from PAH)	16-2
Table 16-2:	Valuable Constituents of Ore (October 2005)	16-2
Table 17-1:	Number of Drill Holes and Metres by Drilling Type	17-7



Project No.: 147107		TOC iv


Table 17-2:	Zone Names in Monte Negro Deposit	17-9
Table 17-3	Monte Negro Zone Names	17-10
Table 17-4:	Moore Model Limits	17-13
Table 17-5	Monte Negro Model Limits	17-13
Table 17-6:	Resource Classification and Estimation Statistics	17-16
Table 17-7:	Resource Summary at 1.7 g/t Au Breakeven Cutoff	17-19
Table 17-8:	Mineral Resources, 27 September 2005, Exclusive of Mineral Reserves	17-19
Table 17-9:	Proven and Probable Mineral Reserves, 27 September 2005 at $400 Gold Price	17-20
Table 19-1:	Mineral Reserves in Moore and Monte Negro Pits at $375 Gold Price	19-5
Table 19-2:	Proven and Probable Mineral Reserves for Moore and Monte Negro Deposits, 27 September 2005 at a $400 Gold Price	19-5
Table 19-3:	Inferred Limestone Mineral Resource by Quarry Site	19-6
Table 19-4:	Capital Costs by Major Items	19-25

Figures


Figure 4-1:	Location of Pueblo Viejo Project, Dominican Republic	4-3
Figure 4-2:	Pueblo Viejo Property Location Map	4-5
Figure 7-1:	Regional Geology	7-2
Figure 7-2:	Property Geology	7-3
Figure 7-3:	Moore Section 94600N – Lithology (Looking North)	7-4
Figure 7-4:	Moore Section 94850N – Lithology (Looking North)	7-5
Figure 7-5:	Monte Negro Section 95800N – Lithology (Looking North)	7-7
Figure 9-1	Moore Deposit Structure and Mineralization	9-6
Figure 9-2:	Alteration and Mineralization, Section 94600 N, West Flank and Vein Zone, Moore Deposit	9-8
Figure 9-3	Monte Negro Deposit Structure and Mineralization	9-10
Figure 9-4:	Cross Section of Alteration and Mineralization, Monte Negro Central Zone at 95650N (Looking North)	9-11
Figure 11-1:	Location of All Historical Drill Holes – Moore Deposit	11-2
Figure 11-2:	Location of All Historical Drill Holes – Monte Negro Deposit	11-3
Figure 11-3:	Core Holes in Moore Deposit	11-5
Figure 11-4:	Monte Negro Drill Hole Location Map	11-6
Figure 14-1:	AMEC Comparison of PDDC and Rosario Drill Hole Assays with 10 m Search Distance	14-13
Figure 14-2:	Idealized Example of Cyclicity and Decay	14-17
Figure 17-1:	Moore Section 94600 N Lithology and Structure	17-9
Figure 17-2:	Monte Negro Section 95800N Lithology and Structural Domains	17-10
Figure 17-3:	Moore Section 94500N Showing Mineralization and Resource Classification	17-17
Figure 17-4:	Monte Negro Section 95600N Showing Mineralization and Resource Classification	17-18
Figure 19-1:	General Site Layout	19-3
Figure 19-2:	Final Mining Limits	19-9
Figure 19-3:	Simplified Flowsheet for Processing Pueblo Viejo Ore	19-11



Project No.: 147107		TOC v

1.0		SUMMARY

		Placer Dome Inc. (PDI) commissioned AMEC Americas Limited, formerly AMEC E&C Ltd (AMEC) to review mineral resource and mineral reserve estimates for the Pueblo Viejo Project in the Dominican Republic and determine if these estimates have been carried out in accordance with industry standard practices and are compliant with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves and Canadian National Instrument 43-101. The project is held under the subsidiary Placer Dome Dominicana Corporation (PDDC). AMEC prepared this Technical Report in support of the public disclosure of Mineral Reserve and Mineral Resource estimates for the project as of 27 September 2005. The format and content of the report are intended to conform to Form 43-101F1 of National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators. Susan Lomas, P.Geo., an employee of AMEC, served as the Qualified Person responsible for completing this Technical Report. Susan was assisted in a review of assay quality assurance and quality control by Dr. Ted Eggleston, R.Geo. in review of twin data, data validation and simulation studies by Kevin Francis, R.Geo., and in review of metallurgical information by Lynton Gormely, P.Eng. Susan was further assisted in review of environmental conditions and permitting by Dennis Krochak and review of capital costs by Ron Pearce. Tracy Barnes, P.E., was the Qualified Person responsible for review of the resources and reserves. Susan and Tracy visited the site between 10 and 14 January 2005.

1.1		Project Description

		The Pueblo Viejo Project is located in the central part of the Dominican Republic on the Caribbean island of Hispanola in the province of Sanchez Ramirez. The project is 15 km west from provincial capital of Cotui and approximately 95 km northwest of the national capital of Santo Domingo.

		The main port facility in the Dominican Republic is Haina in Santo Domingo. Other port facilities are located at Puerto Plata, Boca Chica, and San Pedro de Macoris.

		On 27 September 2005 Placer Dome announced the decision to proceed with development of the Pueblo Viejo project. Capital costs are estimated to be approximately $1 billion. In addition, a 140 MW, coal-fired power plant, and power line will be constructed at a cost of approximately $350 million to provide power to for the project.

		Gold ore will be mined by conventional open pit methods from two pits. Total mine life is anticipated to be 20 years. In the initial six years of operation, the mine will produce an average of 800,000 oz Au/a at cash costs ranging between $175 and $185/oz. The ore is refractory and will be treated at a rate of 15,000 t/d by whole ore pressure oxidation, followed by CIL cyanide leach.


Project No.: 147107		Page 1-1

Commencement of construction is subject to Placer Dome issuing a project notice to the Dominican Republic and this is conditional on receipt of permits for the mine and power plant and the Dominican Republic fulfilling all of its contractual obligations to the project.

1.2		Special Lease Agreement

		During the year 2000, the Dominican State invited international bids for the leasing and mineral exploitation of the Pueblo Viejo sulphide deposits. On 2 July 2001 Placer Dome America Holding Corporation presented a Technical and Financial Offer to the Dominican State committee. Subsequently PDDC won the bid, and the parties signed a Letter of Intent on 4 August 2001. The Special Lease Agreement (SLA) was published in the Official Gazette of the Dominican Republic on 21 May 2003.

		The SLA officially became effective on 29 July 2003 (the Approval Date), when PDDC expressed its satisfaction that all relevant conditions had been satisfied.

1.3		Environmental Review

		AMEC reviewed the existing environmental conditions at the mine and the plans PDDC has or will put in place to construct, operate, and close the mine.

1.3.1		Assessment of Existing Conditions

		PDDC commissioned a number of consultants to collect background baseline information on the existing biophysical and human environments from 2002 through 2004. Key findings indicated significant historical environmental impact, particularly with respect to ARD and metals leaching, which has degraded water quality in the Arroyo Margajita in particular and to a lesser extent in the Rio Maguaca. Water management infrastructure was found to be in poor condition and contributing to aquatic impacts. There were large amounts of hazardous wastes on the mine site (which remain). Pit walls, waste rock dumps, and stockpiles continue to erode, depositing sediments behind dams and water control structures. Stability of the Mejita and Las Lagunas dams was identified as a key issue. Under the concession agreement signed between PDDC and the Dominican government, existing liabilities are the responsibility of the Dominican government.

1.3.2		Future Liabilities and Risks

		Risks fall into three categories—permitting/approval, environmental, and social—and liabilities generally into two categories—environmental and social. The Dominican government has not previously permitted a mine separately from granting a concession (e.g., Falcondo and Rosario Dominicana), and therefore the time period involved in obtaining project approval and permits is not known; however, timeline guidelines suggest


Project No.: 147107		Page 1-2

approximately six months. In addition, a large number of permits must be obtained. There is a reasonable expectation that permits will be obtained.

A number of environmental risks are the responsibility of the Dominican government, but failures with any of these could reflect negatively on a new operator. These include:

	•		Failure of the existing tailing dams

	•		The inability of the Dominican government to satisfactorily relocate people currently living in areas that will contain mine facilities

	•		Acid-rock drainage. Pueblo Viejo is a significant ARD site. ARD is a major environmental issue because of the relatively high sulphide levels in the ore coupled with the wet tropical climate. While there are substantial limestone deposits in the mine area, post-closure exhaustion of limestone before all ARD is treated is an unknown risk but can be better quantified during the operations phase. Unplanned increased environmental treatment costs could occur if metals levels exceed those expected and natural degradation of cyanide is less effective than expected.

Currently, acidic drainage containing elevated metals is flowing from the mine site, particularly in the Margajita basin. Water management and treatment facilities are in need of repair and upgrading. With re-opening of the mine and construction of new facilities, this situation is expected to improve.

AMEC considers PDDC’s assessment of risks, liabilities, and costs associated with construction and operation of the Pueblo Viejo mine to have been conducted thoroughly. While more work remains, the information assembled to date represents a sound basis for further study. Environmental risks and liabilities centre on possible dam failure of the two existing dams which are government responsibilities.

Closure costs have been planned for, and additional post-closure costs will be covered by funds established during mine operation.

1.4		Property History

		The Pueblo Viejo property has a long history of exploration, development and production. The Dominican Government discovered the deposit in 1950 during a regional geological mapping program. Small-scale production was unsuccessfully attempted in subsequent years. Rosario Resources Corporation optioned the property in 1969 and commenced production of oxide ore from open pits in 1975. Production of oxide ores continued until 1991. Starting in 1986, Rosario examined options for treating refractory mineralization as oxide reserves diminished. Rosario rejected the available options as not being environmentally feasible and instead built a CIL plant and new tailings facilities to process transitional oxide-sulphide ores. Production continued at reduced rates until 1999


Project No.: 147107		Page 1-3

In 24 years of production, the Pueblo Viejo mine produced a total of 5.3 Moz of gold and 24.4 Moz of silver.

Starting in 1996, Rosario and the Dominican Republic invited bidders to participate in development of sulphide mineralization. GENEL JV, Mount Isa Mines (MIM) and Newmont each carried out technical work, including drilling, resource estimation metallurgy and economic assessments, in preparation of bids, but none concluded an agreement with the government.

Placer Dome carried out additional studies starting in 2002 and concluding in an internal feasibility study in 2005. Work included 3,039 m of core drilling in 18 holes during 2002 and 15,424 m of core drilling in 111 holes during 2004, revision of resource estimates, new mine designs and reserve estimates, metallurgy and process studies, capital and operating costs estimates and financial analyses. These concluded that mining of gold mineralization from the Moore and Monte Negro deposits would be economic and Placer Dome announced on 27 September 2005 a decision to proceed with development of the project.

1.5		Geological Setting and Mineralization

		Mineralization at Pueblo Viejo occurs within the Lower Cretaceous Los Ranchos Formation, a series of volcanic and volcaniclastic rocks that extend across the eastern half of the Dominican Republic. The Pueblo Viejo Member of the Los Ranchos Formation, the most important ore host, fills a restricted sedimentary basin measuring approximately 3 km north-south by 2 km east-west. Formation of the basin is interpreted to be either rift derived or a maar-diatreme complex that cut through lower members of the Los Ranchos Formation.

		The Moore deposit is located at the eastern margin of the Pueblo Viejo Member sedimentary basin. Stratigraphy consists of finely bedded carbonaceous siltstone and mudstone overlying horizons of spilite (basaltic-andesite flows), volcanic sandstone, and fragmental volcaniclastic rocks.

		The Monte Negro deposit is located at the northwestern margin of the sedimentary basin. Stratigraphy consists of interbedded carbonaceous sediments ranging from siltstone to conglomerate that are interlayered with volcaniclastic flows.

		The Moore and Monte Negro deposits are the largest of several high sulphidation gold deposits in the district. Mineralized structures at Pueblo Viejo are both sub-horizontal and sub-vertical. The deposits occupy funnel-shaped envelopes of advanced argillic alteration characterized by the presence of silica, pyrophyllite, pyrite, kaolinite, and alunite. Gold primarily occurs as solid solution and colloidal-sized microinclusions within the crystal


Project No.: 147107		Page 1-4

structure of pyrite and less commonly as micron-sized grains of native gold, electrum, sylvanite [AuAgTe₄], and aurostibite [AuSb₂]).

1.6		Drilling and Verification of Data

		Drilling campaigns have been conducted by most of the participating companies in the project over the years, including Rosario, GENEL, MIM, and PDDC. Resource estimates are supported with results from 573 RC and core drill holes (349 drill holes for Moore deposit, 224 drill holes for Monte Negro deposit) totalling 80,261 m.

		Approximately 65% of the drill holes used in the 2005 PDDC resource estimate were drilled vertically, with the remainder being inclined to intercept both vertical and subhorizontal structures.

		The history of the project database is well documented in the literature provided to AMEC for this study. This includes detailed descriptions of the data entry and validation procedures completed by both Fluor and AMS during their studies in the mid-1980s and 1992.

1.6.1		Rosario Drilling

		Rosario employed core, RC, and rotary percussion drilling methods during its exploration and development phases. Logged information was recorded on paper and later manually entered into an electronic database. AMEC verified 2.5% of the data generated by this program against original documents that could be recovered. Remaining holes were validated using nearest-pair comparisons with PDDC drill holes. PDDC paired assays from PDDC and GENEL drilling with assays from Rosario drilling using different search radii and then prepared QQ plots to show assay grade distributions for each. A nearest-pair analysis was necessary because most Rosario holes are vertical and nearly all GENEL, MIM and PDDC holes are inclined. AMEC constructed declustered QQ plots and confirmed conclusions by PDDC. In grade ranges below 2 g/t, Rosario drilling appears to be slightly biased high compared to PDDC drilling; however, above 2 g/t and below 6 g/t the assays fit a line x=y reasonably well (Figure 14-1). Above 6 g/t Rosario drilling appears generally biased high relative to PDDC drilling.

		This work generally shows the Rosario drilling to be reliable but represents a risk to the project where projected high-grade material may be lower than anticipated.

		Core recovery for Rosario drilling was poor, being about 50% in areas of mineralization and silicification. This resulted in globally offsetting positive and negative biases but these biases may vary significantly by recovery class. Loss of fines in sampling negatively affected Zn values.


Project No.: 147107		Page 1-5

1.6.2		GENEL Drilling

		In 1996, GENEL drilled 20 holes at the Pueblo Viejo project site, 11 in the Moore deposit and 9 in the Monte Negro deposit. All holes were drilled at an angle. AMEC verified the data generated by this program against original documents.

1.6.3		MIM Drilling

		In late 1996 and into 1997, MIM drilled 31 holes, 15 at the Moore deposit and 16 at the Monte Negro deposit. Five holes were vertical and 26 were drilled at an angle.

		Original records are not available for this drilling campaign, therefore these holes were demonstrated to be reliable by comparing them to PDDC and GENEL holes in a nearest-pair analysis and inspections of sections with PDDC and GENEL drill holes.

1.6.4		PDDC Drilling

		PDDC completed 3,039 m of diamond drilling in 18 holes during the months of October and November 2002 and 15,424 m in 111 drill holes in 2004. All but one of these holes were angle holes. This drill program largely confirmed the location and tenor or mineralization at Pueblo Viejo. AMEC verified these data against original documents.

1.6.5		Twinned-Hole Comparison

		Fluor undertook a study of twinned drill holes as part of its 1986 feasibility study. The results showed some differences between some of the drill holes. Fluor’s comments about the twin holes results include: “Analysis of Au, Ag, Zn, C, and S assays then showed no significant overall bias except for carbon and zinc. Carbon assays were consistently lower by 7% and zinc assays lower on average by 36% than in the original hole.”

		PDDC reviewed 20 twinned and closely spaced drill holes and compared the gold grades. The results of this analysis show good agreement between different drilling methods — rotary, reverse circulation, and core — when the holes are closely spaced.

1.7

Sampling Method and Approach

1.7.1		Pre-PDDC Drilling Programs

		Information about sampling strategies used by the Rosario, GENEL, and MIM companies is not available. Rosario samples were prepared and assayed by Rosario personnel at the mine site laboratory. AMEC does not know where the GENEL and MIM samples were


Project No.: 147107		Page 1-6

processed and assayed, but previous reviewers for this project have stated that the samples were processed at the mine site lab.

PDDC core logging and splitting met industry standard practices for this type of deposit and was performed by PDDC personnel. Sample splits were sent to a commercial laboratory with the remaining half retained for reference.

1.8		Sample Preparation, Analyses, and Security

		Minimal assay quality assurance and quality control (QA/QC) procedures were practiced in Rosario drilling, which was common for this period. PDDC validated historical drilling by Rosario by performing nearest-pair twin analyses with PDDC core holes, by comparing resource estimates with and without PDDC holes and by performing conditional simulation studies.

		Poor precision between sample duplicates in the 2002 PDDC drill program suggests that sample preparation methods could have been better. Duplicate sample analysis for the GENEL drilling show acceptable precision.

		Quality assurance-quality control (QA/QC) procedures have varied significantly during the history of work at Pueblo Viejo. QA/QC for the Rosario work consisted of two batches of check assays sent to a second laboratory and is considered by AMEC to be substandard. GENEL QA/QC involved insertion of standard samples and analysis of pulp duplicate samples and is considered by AMEC to be adequate.

		MIM sample data lack any QA/QC validation. MIM holes were validated by comparing these to nearby PDDC and GENEL holes.

		QA/QC by PDDC relied on two standards and check assaying and is substandard relative to current industry practices, but is generally acceptable to AMEC. No duplicate samples were analyzed, and the check analysis program included no certified reference materials or blank samples. AMEC recommends that, in addition to check assaying, future QA/QC programs include field duplicate samples, crusher duplicate samples, pulp duplicate samples, blind duplicate samples, standard samples, and blank samples.

1.9		Density

		AMAX Engineering and Mining Services derived a linear regression formula for density based on 152 pairs of density and sulphur samples from the GR series of diamond drill holes in 1985. AMEC validated the equation.


Project No.: 147107		Page 1-7

1.10		Mineral Processing and Metallurgical Testing

		Gold recoveries predicted from testing prior to PDDC’s involvement were in the range of 80% to 84%. Each of the concepts that were capable of yielding gold recovery in this range involved expensive destruction of the sulphide minerals, either by roasting or oxidative leaching, while rejecting 16% to 20% of the gold.

		After investigating a number of alternatives, PDDC has selected a fairly straightforward process based on pressure oxidation of the whole ore followed by carbon-in-leach cyanidation for recovery of the gold and silver. Autoclaving of the whole ore entails greater capital and operating costs than treating a concentrate, but the increment in operating cost is limited because of its dependence more on sulphur throughput rather than tonnage throughput. The increment in gold recovery with whole ore pressure oxidation is significant: gold recoveries are projected to vary with ore type from 88.7% to 94%. These projections are supported by an extensive test program completed at SGS Lakefield in Peterborough, including a 10-day continuous autoclave and cyanidation pilot-plant campaign, which reported 93.7% overall extraction.

		The extent of testing and pilot plant operation that underpins the Feasibility Study is appropriate for this level of project evaluation. The extent of investigation into metallurgical variability and the level of testing undertaken give AMEC a high level of confidence that the predicted recoveries for the various ore types will be very close to what would be experienced if the selected flowsheet were constructed

1.11		Mineral Resource Estimates

		Placer Dome Technical Services Limited (PDTS) in Vancouver, B.C., produced the resource models for the Pueblo Viejo Project for PDDC as part of an internal Feasibility Study completed for the project. The models were completed in early 2005 using the data available at that time. The database includes drilling from Rosario, MIM, GENEL, and PDDC. Two separate models were produced, one for each of the deposits at Pueblo Viejo—the Moore deposit and the Monte Negro deposit. The mineral resource was delineated by 314 core holes, 64 reverse circulation holes and 195 rotary holes, for a total of 573 drill holes totalling 80,261 m of drilling.

		As part of AMEC’s independent review, grade interpolation parameters were examined for suitability and the models checked for validity. AMEC supports the mineral resource declared by PDDC and believes it has been developed using methods consistent with accepted industry practices.


Project No.: 147107		Page 1-8

1.12		Mining Operations

		Placer Dome Technical Services (PDTS) has completed a number of studies investigating the Pueblo Viejo Project, including a Feasibility Study concluded in July 2005. These investigations have evaluated various throughput options and mining cutoff grade strategies. PDTS has shown that modestly positive economics can be achieved for the project. PDTS has shown that for a positive outcome for the project to achieved, the following are necessary:

	•		A minimum throughput rate of 15,000 t/d must be achieved.

	•		An aggressive cutoff grade strategy must be implemented to process the highest grade ore early in the mine life with lower grade ore stockpiled for processing later in the mine life. PDTS has demonstrated that this should be achievable.

	•		Proper blending of the ore will be necessary to maximize the utilization of the process facility and achieve required gold production. PDTS has demonstrated that the blending should be achievable.

The July 2005 Feasibility Study has evaluated the project to the level of detail that would be expected within the mining industry for this level of study. The work done has been very thorough and completed to a high professional standard. The potential problems with the project have been recognized and addressed.

The envisioned mining operations consists of mining from two open pits: Moore and Monte Negro; and from three limestone quarries: Plant, Las Lagunas, and Los Quemados. The mines will follow a conventional drill-blast-load-haul mining sequence. Mining operations (equipment and manpower) will be interchangeable between the mines and quarries. The ore from the mines will be hauled to a crusher near the plant site or to the low-grade stockpile. The waste from the mines will be hauled to the El Llagal tailings facility. It is anticipated that all of the waste mined will be acid generating. Therefore, all of the potential acid drainage will be contained within the tailings facility. The limestone will be mined and used as road construction material, rock for the construction of the El Llagal tailings dam, and for processing and water management operations. The process will require high quality limestone for the production of lime.

The ore reserve was developed using only the Measured and Indicated portion of the resource model. Mining limits, mine and quarry production schedules, mine operating costs, mining equipment requirements, mining capital, and mine sustaining costs were developed by PDTS based on first principles and experience with their other operating properties. The mine production schedule uses a staged mine development. The high-grade ore will be processed immediately as it is mined while the low-grade ore will be stockpiled and processed after the mine is exhausted. Gold cutoff grades vary by mining stage and were selected to optimize gold production given the ore tonnage and sulphur constraints in the plant.


Project No.: 147107		Page 1-9

AMEC has reviewed the mine designs and mine plans. AMEC believes that the designs and plans are suitable and generally achievable as planned.

The choice of equipment is appropriate for the deposit and mining schedule. The mining capital and operating costs have been developed in a thorough manner and appear to be reasonable.

1.13		Mineral Resource and Mineral Reserve Statements

		AMEC has reviewed the Mineral Resource and Mineral Reserve statements as of 27 September 2005 and accepts these statements. Proven and Probable Mineral Reserves (Table 1-1) total 129.7 Mt at an average grade of 3.22 g/t, containing 13.4 Moz Au. Average expected metallurgical recovery is 91.5%. Life of mine waste to ore strip ratio is 0.69:1.

		Table 1-1: Proven and Probable Mineral Reserves, 27 September 2005 at $400/oz Au Price


	Tonnes		Gold Grade		Contained Ounces
	(Mt)		(g/t)		(000s)

Proven Mineral Reserve		95.9		3.28		10,112
Probable Mineral Reserve		33.8		3.04		3,304
Total Proven + Probable Mineral Reserve		129.7		3.22		13,427

Note: Mineral Reserves are within open pits designed using a life-of-mine average cutoff grade of 1.7 g/t Au and a gold price of $375/oz. Mineral Reserves within this pit are tabulated as blocks with a positive value at a $400/oz Au price.

Mineral Resources exclusive of mineral reserves are listed in Table 1-2.

Table 1-2: Mineral Resources, 27 September 2005 at $450/oz Au Price


	Tonnes		Gold Grade		Contained Ounces
	(Mt)		(g/t)		(000s)

Measured Mineral Resources		22.9		2.20		1,618
Indicated Mineral Resources		11.0		2.38		846
Measured + Indicated Resources		33.9		2.26		2,464

Inferred Mineral Resources		2.2		2.72		190

Note: Mineral Resources are contained within a pit designed using an average cutoff grade of 1.5 g/t Au and a gold price of $450/oz.

1.14		Mill Operating Cost

		The mill operating cost is presented as formulas that permit the calculation of the operating cost per tonne as a function of power cost and ore sulphur content. Under the assumed unit power cost of US$0.03/kWh in the base case, with an 8% sulphur ore feed, power


Project No.: 147107		Page 1-10

accounts for 30% of the mill operating cost, labour for 12%, and consumables, including mill liners, mill media, and chemicals, for 41%. Consumables therefore represent a large component of the costs.

The build-up of these formulas is described in the July 2005 Feasibility Study. AMEC has reviewed these calculations and is in agreement that the derivation methods are appropriate for a feasibility level of evaluation.

1.15		Capital and Operating Cost Estimates

		The capital cost estimate of US$1.348 billion for the Pueblo Viejo project includes direct costs for general site preparation, the open pit mine, the processing plant, ancillary facilities, tailings disposal, site access road and offsite power generation and high voltage transmission line. Also included are indirect costs for design and engineering, project management, construction overheads, PDDC operations overheads, warehouse inventory and freight. A 10% contingency has been added to the above costs and an allowance for initial operating expenses has been added “below the line.” All costs are stated in 2^nd quarter 2005 U.S. dollars.

		AMEC did not review power plant and transmission line costs prepared for PDTS by SNC Lavalin because this is outside the expertise of AMEC cost estimators. AMEC assumes that the estimates were prepared by a Qualified Person. Otherwise, AMEC believes that the estimated capital costs are reasonable and complete and that they have been prepared in accordance with industry standards at levels of detail commensurate with a feasibility study.

1.16		Economic Analysis

		Economic analysis of the Pueblo Viejo project is based upon discounted cash flow analysis on a post-local tax basis. Projections for annual revenues and costs are based on data developed in a feasibility study dated July 2005 by Placer Dome Technical Services Limited for the mine, process plant, power plant and transmission line, capital expenditures and operating costs.

		Metal prices for the base case financial model (US$400/oz Au and US$6.00/oz Ag) provided by Placer Dome conform to the prices used to compute reserves. Based on the assumptions used, discounted cash flow analysis indicates that the project offers a positive return with a payback of less than half the mine life.

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


Project No.: 147107		Page 1-11

All other inputs are appropriate and, apart from the first few years of development, future cash flows are positive.

As with many projects of this type, the Pueblo Viejo Project is most sensitive to changes in metal price and rather less so to changes in operating cost and capital expenditure.

The mine schedule shows ever-decreasing grade material being processed over the mine life. While this may allow the discounted cash flow to be maximized, it also causes the operation to become more and more marginal as time passes. This condition, if unexpectedly combined with any significant drop in gold price, could result in the profitability being impaired.

In AMEC’s opinion, the level of detail used in the economic analysis is appropriate for a feasibility study


Project No.: 147107		Page 1-12

2.0		INTRODUCTION AND TERMS OF REFERENCE

		Placer Dome Inc. (PDI) commissioned AMEC Americas Limited, formerly AMEC E&C Ltd (AMEC) to review mineral resource and mineral reserve estimates for the Pueblo Viejo Project in the Dominican Republic and determine if these estimates have been carried out in accordance with industry standard practices and are compliant with Canadian Institute of Mining, Metallurgy and Petroleum (CIM) Standards on Mineral Resources and Reserves and Canadian National Instrument 43-101. The project is held under the subsidiary Placer Dome Dominicana Corporation (PDDC). AMEC prepared this Technical Report in support of the public disclosure of Mineral Reserve and Mineral Resource estimates for the project as of 27 September 2005. The format and content of the report are intended to conform to Form 43-101F1 of National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators. Susan Lomas, P.Geo., an employee of AMEC, served as the Qualified Person responsible completing this Technical Report. Susan was assisted in a review of assay quality assurance and quality control by Dr. Ted Eggleston, RGeo, in review of twin data, data validation and simulation studies by Kevin Francis, RGeo and Henrique da Silva, Mining Engineer, and in review of metallurgical information by Lynton Gormely, P.Eng. Susan was further assisted in review of environmental conditions and permitting by Dennis Krochack and review of capital costs by Ron Pearce. Tracy Barnes, P.Eng., was the Qualified Person responsible for review of the resources and reserves. Both Qualified Persons visited the site between 10 and 14 January 2005.

2.1		Terms of Reference

		The Pueblo Viejo Project is a gold deposit. All drill hole data used for the resource and reserve review were obtained from PDDC personnel both at site and from the PDI Vancouver office.

		All units in this report are metric:


Above mean sea level		amsl
Ampere		A
Annum (year)		a
Centimetre		cm
Cubic centimetre		cm³
Cubic metre		m³
Day		d
Days per week		d/wk
Days per year (annum)		d/a
Dead weight tonnes		DWT
Degree		°
Degrees Celsius		°C
Diameter		ø
Gram		g


Project No.: 147107		Page 2-1


Grams per litre		g/L
Grams per tonne		g/t
Greater than		>
Hectare (10,000 m²)		ha
Hour		h
Hours per day		h/d
Hours per week		h/wk
Hours per year		h/a
Kilo (thousand)		k
Kilogram		kg
Kilograms per m3		kg/m³
Kilograms per hour		kg/h
Kilograms per square metre		kg/m²
Kilometre		km
Kilometres per hour		km/h
Kilovolt		kV
Kilovolt-ampere		kVA
Kilowatt		kW
Kilowatt hour		kWh
Kilowatt hours per tonne (metric ton)		kWh/t
Kilowatt hours per year		kWh/a
Less than		<
Litre		L
Litres per minute		L/m
Megabytes per second		Mb/s
Megapascal		MPa
Megavolt-ampere		MVA
Megawatt		MW
Metre		m
Metres above sea level		masl
Metres per minute		m/min
Metres per second		m/s
Metric ton (tonne)		t
Micrometre (micron)		µm
Milliamperes		mA
Milligram		mg
Milligrams per litre		mg/L
Millilitre		mL
Millimetre		mm
Million		M
Million ounces		Moz
Million tonnes		Mt
Minute (plane angle)		'
Minute (time)		min
Month		mo
Parts per billion		ppb


Project No.: 147107		Page 2-2


Parts per million		ppm
Pascal (newtons per square metre)		Pa
Pascals per second		Pa/s
Percent		%
Percent moisture (relative humidity)		% RH
Phase (electrical)		Ph
Power factor		pF
Revolutions per minute		rpm
Second (plane angle)		"
Second (time)		s
Specific gravity		SG
Square centimetre		cm²
Square kilometre		km²
Square metre		m²
Thousand tonnes		kt
Tonne (1,000 kg)		t
Tonnes per day		t/d
Tonnes per hour		t/h
Tonnes per year		t/a
Variable Frequency Drive (Electrical)		VFD
Variable Speed (Fluid coupling)		VS
Volt		V
Watt		W
Week		wk
Weight/weight		w/w
Wet metric ton		wmt
Yard		yd
Year (annum)		a
Year (U.S.)		yr


Project No.: 147107		Page 2-3

3.0		DISCLAIMER

		Information regarding mining agreements and operational permits were acquired from PDDC staff and were not independently verified by AMEC. This information is assumed to be accurate.

		AMEC relied upon estimates of costs for a power plant and transmission lines that were prepared by SNC Lavalin because these facilities are outside AMEC’s area of expertise. AMEC assumes that the SNC Lavalin estimates were prepared by a Qualified Person.


Project No.: 147107		Page 3-1

4.0

PROPERTY DESCRIPTION AND LOCATION

4.1		Project Location

		The Pueblo Viejo Project is located in the central part of the Dominican Republic on the Caribbean island of Hispanola in the province of Sanchez Ramirez (Figures 4-1 and 4-2). The project is 15 km west from provincial capital of Cotui and approximately 95 km northwest of the national capital of Santo Domingo.

		The Pueblo Viejo property, referred to as the “Fiscal Reserve Montenegro,” is centred at 19°02’ N, 70°08’ W in an area of moderately hilly topography (Figure 4-2). The Fiscal Reserve Montenegro covers an area of 3,200 ha and encompasses all of the areas previously included in the Pueblo Viejo and Pueblo Viejo II concession areas, which were owned by Rosario Dominicana until 7 March 2002.

		Access from Santo Domingo is by a four-lane paved highway (Autopista Duarte), which is the main route between Santo Domingo and the second-largest city, Santiago. This highway connects to a single-lane, secondary highway #17 at the town of Piedra Blanca, approximately 70 km from Santo Domingo. This secondary highway is a two-lane paved highway that passes through the towns of Piedra Blanca and Maimon on the way to Cotui. The gatehouse for the mine is 22 km from Piedra Blanca.

		The main port facility in the Dominican Republic is Haina in Santo Domingo. Other port facilities are located at Puerto Plata, Boca Chica, and San Pedro de Macoris.

4.2		Special Lease Agreement

		During the year 2000, the Dominican State invited international bids for the leasing and mineral exploitation of the Pueblo Viejo sulphide deposits. On 2 July 2001 Placer Dome America Holding Corporation presented a Technical and Financial Offer to the Dominican State committee. Subsequently Placer Dome Dominicana Corporation (PDDC) won the bid, and the parties signed a Letter of Intent on 4 August 2001. The Special Lease Agreement (SLA) was published in the Official Gazette of the Dominican Republic on 21 May 2003.

		The SLA officially became effective on 29 July 2003 (the Approval Date), when PDDC expressed its satisfaction that all relevant conditions had been satisfied. These included “Publication and Promulgation of the SLA in the Gazette to assure free and clear title to the Fiscal Reserve; Opinion from a selected law firm; approval of the SLA by National Congress; and Recordation of the Approved Agreement in the Public Registry of Mining Rights.” Pertinent terms of the SLA are:


Project No.: 147107		Page 4-1

	•		The SLA will extend for 25 years following notice by PDDC to the Dominican State that PDDC has determined to develop a mine at the Pueblo Viejo site (Project Notice), with one extension by right for 25 years and a second 25 year extension at the mutual agreement of PDDC and the Dominican State, allowing a possible total term of 75 years.

	•		PDDC may exploit the Los Quernados Limestone Deposit and all other limestone deposits with the Fiscal Reserve at no additional charge.

	•		The Dominican State will acquire and lease to PDDC the lands and mineral rights necessary for the permanent disposal of tailings and waste.

	•		The Dominican State will mitigate all Historical Environmental Matters, except those conditions within areas designated for development by PDDC in the Project Notice.

	•		The Dominican State will relocate at its sole cost and in accordance with World Bank Standards, those persons dwelling in the Los Cacaos section of the Site.

	•		The Dominican State will provide a permanent reliable source of water necessary to conduct the operations at no additional charge to PDDC.

	•		PDDC shall make Net Smelter Royalty payments to the Dominican State of 3.2% of net receipts of sales, make a Net Profits Interest payment (with a rate that varies with the price of gold) after PDDC has recaptured its initial and ongoing investments and pay income tax under a stabilized tax regime.


Project No.: 147107		Page 4-2

Figure 4-1: Location of Pueblo Viejo Project, Dominican Republic


Project No.: 147107		Page 4-3

Figure 4-2:Pueblo Viejo Property Location Map


Project No.: 147107		Page 4-4

4.3		Environmental Review

		AMEC reviewed the environmental issues at the Pueblo Viejo mine in order to provide an independent assessment of environmental liabilities and risks for the purposes of this report. The scope of the review covered the existing conditions at the mine and the plans PDDC has or will put in place to construct, operate, and close the mine, should the company move forward with developing the deposit.

4.3.1		Information Sources

		Information for this review was obtained principally from the following sources:

•

“Special Lease Agreement of Mining Rights” between the Dominican State, The Central Bank of the Dominican Republic, Rosario Dominicana, S.A., and Placer Dome Dominicana Corp., specifically the following sections:

			- Section 7.2 Remediation of Historic Environmental Matters

			- Section 11.2 Assignment of Responsibility for the Management and Remediation of Environmental Conditions.

	•		Volume 4 of the Pre-feasibility Report, December 2004 (Placer Dome 2004).

	•		Interviews on site with Ana Maria Ibacache, Manager of Environment and Erling Villalobos, Manager of Engineering for PDDC.

	•		Site inspection of key facilities including the Mejita dam and tailings pond, the Las Lagunas dam and tailings pond, the UM-1 dam, the Montenegro and Moore pits, and the processing plant, as well as other minor ancillary facilities discussed in this report.

4.3.2		Assessment of Existing Conditions

		PDDC commissioned a number of consultants to collect background baseline information on the existing biophysical and human environments from 2002 through 2004:

	•		Rescan Environmental Services – water flows and quality, aquatic biota, terrestrial flora and fauna, preliminary hazardous wastes survey, wetland characterization

	•		Ambimet Ltda. – air quality (PM-10), climate

	•		Morwijk Enterprises – acid rock drainage (ARD)

	•		Harold Olsen Bogaert – archaeology

	•		PDDC (assisted by international consultants) – geology (including a comprehensive soil geochemistry survey to identify metal anomalies and sulphides


Project No.: 147107		Page 4-5

	•		Clearwater Consultants – hydrology

	•		Water Management Consultants – groundwater

	•		Dra. Amparo Chantanda (Dominican Republic) – socio-economic baseline.

Key findings indicated significant historical environmental impact, particularly with respect to ARD and metals leaching, which has degraded water quality in the Arroyo Margajita in particular and to a lesser extent in the Rio Maguaca. Water management infrastructure was found to be in poor condition and contributing to aquatic impacts. There were large amounts of hazardous wastes on the mine site (which remain). Pit walls, waste rock dumps, and stockpiles continue to erode, silting in dams and water control structures. Stability of the Mejita and Las Lagunas dams was identified as a key issue. Under the concession agreement signed between PDDC and the Dominican government, existing liabilities are the responsibility of the Dominican government. Specifically:

	•		The Dominican State will bear responsibility for the management and remediation of all historic environmental liability associated with the Pueblo Viejo site until Placer Dome makes a decision to move forward with development of the project as indicated through providing the Dominican State with a “Project Notice.”

	•		PDDC would assume responsibility for all environmental liability, including historic liability for facilities located within the boundaries of the Development Area as designated in the feasibility study, including drainage originating therefrom. This responsibility would be phased in, starting with the commissioning of first stage processing and attaching fully following the commencement of commercial production. In other words, PDDC would assume responsibility for remediation of the historic environmental liability within the area designated for development in its project feasibility study. The agreement suggests that this is likely to include the Monte Negro pit, the Moore pit, the plant site, and the El Llagal sites. Liability for historic areas outside its designated development area would remain with the State.

	•		From the outset, PDDC would be responsible for remediation of all environmental liability resulting from its activity on site.

4.4		Liabilities and Risks

		Risks fall into three categories—permitting/approval, environmental, and social—and liabilities generally into two categories—environmental and social. The Dominican government has not previously permitted a mine separately from granting a concession (e.g., Falcondo and Rosario Dominicana), and therefore the time period involved in obtaining project approval and permits is not known; however, timeline guidelines suggest approximately six months. In addition, a large number of permits must be obtained.


Project No.: 147107		Page 4-6

The largest environmental risk is failure of the existing tailing dams. While there are substantial limestone deposits in the mine area, post-closure exhaustion of limestone before all ARD is treated is an unknown risk but can be better quantified during the operations phase. Unplanned increased environmental treatment costs could occur if metals levels exceed those expected and natural degradation of cyanide is less effective than expected.

The most significant social risk is the inability of the Dominican government to satisfactorily relocate people currently living in areas that will contain mine facilities.

Risks and liabilities for pre-operation, operations, and closure/post-closure are discussed further below.

4.4.1		Pre-operation

		The water permit for obtaining freshwater from the Hatillo reservoir has been obtained to ensure an adequate water supply for the mill. This permit is a requirement for filing the EIA report.

		A number of water licences are required for commencement of mining and some remain to be obtained. Some additional land leases will be required for mine facilities. It is reasonable to assume that the remaining permits and leases will be obtained.

		*Permits*

		A relatively large number of permits and authorizations will be required and are listed in Volume 4 of the July 2005 Feasibility Study.

		Major environmental permits will require review by government and other stakeholders and will be a matter of time to obtain. As operation of the mine will not only provide government revenues and employment, but also improve environmental quality of the area and reduce a current long-term liability of the Dominican government (ARD), it is expected that permits will be granted. Rights to use ground and surface water, particularly sources already exploited will require negotiation with all the existing rights holders which could be a lengthy process in the absence of government expropriation of existing rights. For construction of some new mine facilities, such as the El Llagal tailings impoundment, trees will have to be cut. This will require a permit for cutting trees and a forestry management plan. Past experience in the Dominican Republic suggests cutting trees may be controversial, although it is expected that permits will be forthcoming after negotiations including commitments to plant trees in other areas.


Project No.: 147107		Page 4-7

Environment

The largest pre-existing environmental liability at Pueblo Viejo is acid generation and metals leaching, which is the responsibility of Rosario Dominicana. There is, however, no direct risk to PDDC, and upon commencement of mining and adequate water management and treatment, water quality off the site is expected to improve.

A second pre-existing environmental liability is the hazardous waste spread around the site, particularly in the area of the concentrator. Although soil contamination and any hazardous wastes left on the site from previous mine operations are the responsibility of the Dominican government to address, PDDC would most likely have to take the lead in the case of the concentrator site because much of the existing infrastructure would be dismantled to build a new concentrator, and much of soil clean-up would have to occur at that time.

A landfill operated by the previous mine owners was inspected for this report and, except for a number of large truck tires and some miscellaneous scrap, has been largely cleaned up. Full clean-up of this site is also the responsibility of the Dominican government.

Social

The largest social liability at Pueblo Viejo is the number of squatters who have moved onto the El Llagal site and will have to be moved off prior to commencement of mine construction. Relocation is to be undertaken by the Dominican government under terms of the SLA.

4.4.2		Operations

		*Tailings Dams*

		Failure of the existing tailings dams are the largest single risk at the Pueblo Viejo site. Any significant tailings dam incident would reflect on PDDC, regardless of their legal liability.

		Proposed El Llagal Tailings Dam

		BGC Engineering Inc. has produced a conceptual design for the proposed tailings dam (2004). The design follows conventional engineering practice, with the dam to be built to withstand probable maximum earthquakes and with excess capacity to hold a 200-year flood under a 24-hour maximum precipitation event. The spillway will be designed to pass a probable maximum flood. The valley to be used is underlain by faults, but geotechnical drilling has indicated that the faults are tight, i.e., tailings water will not seep into and contaminate groundwater.


Project No.: 147107		Page 4-8

Acid Generation and Metals Release

ARD from the site is the most significant environmental liability at Pueblo Viejo. Management of the ARD situation was initially evaluated for PDDC by Environmental Geochemical International (2003). PDDC accepted all findings, and all subsequent recommendations were incorporated into the 2005 Feasibility Study. Kevin Morin and Nora Hutt of the Minesite Drainage Assessment Group (MDAG) (Morwijk Enterprises) reported on environmental liability cost associated with ongoing treatment of contaminated drainage at the site in its September 2004 report entitled “Pueblo Viejo Project — Supplementary Geochemical Information and Predictions for Pre-feasibility Studies.” Costs for treating contaminated water are included in Feasibility Study operating costs and were based on an estimate of unit cost per tonne of acidity as CaCO3 equivalent, and as a unit cost per tonne of acidity as sulphur. These values were derived from metallurgical testing recently completed for Placer Dome by CEMI based on projected lime and limestone consumptions, with a factor of 5% added to cover other treatment consumables such as flocculant, labour, and power. Existing Rock Piles

Mine Drainage Assessment Group (MDAG 2003) reported that the existing run-of-mine rock piles are currently releasing approximately 69,000 t/a of acidity as CaCO3, which could eventually reach 2.7 Mt/a. MDAG’s estimates of liability for treating contaminated site drainage were confirmed by an independent review (EGI 2003) and appear reasonable.

Low Grade Stockpile

MDAG reported that PDDC plans to create a low-grade ore stockpile as part of its Feasibility Study mine plan. Water treatment of the drainage from this low-grade stockpile was estimated and, again, appears reasonable. If the low-grade ore were not milled, then the cumulative cost of treating from this stockpile would significantly increase, providing a strong impetus to ensure that the material is processed. Given the anticipated ARD and metal leaching from ore stockpiles, waste rock, and pit walls, significant resources will need to be expended on water treatment. ARD will be treated as part of process water. Based on baseline characterization and anticipated water treatment plant discharge quality, which will meet Dominican standards, sulphate will be above receiving environment quality (400 mg/L for drinking water) and will probably be about 1,800 mg/L downstream of the treatment plant discharge. To put this in context, however, 2004 background sulphate concentrations in Arroyo Margajita, the receiving waterbody for treatment plant discharge, ranged from 2,495 to 12,700 mg/L. Other than wetland


Project No.: 147107		Page 4-9

			treatment, which is not being considered at this time, there is no practical treatment for sulphate. Zinc may also be above the criterion (0.05 mg/L) for drinking water because of high natural background levels (Placer Dome, pers. comm.). One possibility that will be examined is to pump the treatment plant water into the Rio Maguaca, where dilution would drop concentrations below the criterion levels (Placer Dome, pers. comm.).

			*Social Costs*

			No exceptional social costs outside of normal employee and family services are expected during operations. Any residual costs from relocation would be the responsibility of the Dominican government.

4.4.3

Closure and Post Closure

			*Environmental*

			Normal costs associated with closure include reclamation, revegetation, and removal of buildings. Cost estimates appear reasonable.

			At closure an environmental risk assessment will be required to determine whether any pathways and receptors for contaminants are being generated at the mine (other than the known ARD/metal leaching). Some liability maybe associated with site clean-up on closure, but PDDC should be able to incorporate this into normal closure costs. ARD generation from pit walls on closure is largely unpredictable at present but is not expected to be a large percentage of the total ARD generated. The Feasibility Study costing provides for ARD treatment for 75 years after closure. In any event, a fund will be established during mine operation to handle ARD treatment costs on closure for the full 75 years when all runoff water is predicted to meet discharge standards. There is residual risk; however, that the local limestone resources for acid neutralization may be exhausted before ARD from the site ceases (Placer Dome, pers. comm.).

			The largest post-closure risk is the long-term stability of the El Llagal tailings dam. Because water will be retained against an upstream “beach” of tailings and not the dam itself, the risk of failure would be considerably lower than for the existing tailings dams. Existing dams (Mejita and Las Lagunas) would remain the responsibility of the Dominican government.

			*Social*

			Severance costs to be incurred by the mine on closure are estimated in the Feasibility Study and are appropriate.



Project No.: 147107		Page 4-10

4.5

Conclusions

			AMEC considers PDDC’s feasibility-level assessment of risks, liabilities, and costs associated with construction and operation of the Pueblo Viejo mine to have been conducted thoroughly. While more work remains, the information assembled to date represents a sound basis for further study. Environmental risks and liabilities centre on possible dam failure of the two existing dams, which are government responsibilities. .

			Pueblo Viejo is a significant ARD site. ARD is a major environmental issue because of the relatively high sulphide levels in the ore coupled with the wet tropical climate. Currently, acidic drainage containing elevated metals is flowing from the minesite, particularly in the Margajita basin. Water management and treatment facilities are in need of repair and upgrading. With re-opening of the mine and construction of new facilities, this situation is expected to improve.

			Closure costs have been planned for, and additional post-closure costs will be covered by funds established during mine operation.

4.6

Permits

			The Dominican Republic Environmental Law No. 64/2000 establishes the process of obtaining environmental permits. There are a number of other technical permits that will have to be assessed by corresponding agencies.

			Agencies involved include:

	•		Secretaría de Estado de Medio Ambiente y Recursos Naturales – SEMARN (Ministry of Environment)

	•		Instituto Dominicano de Recursos Hidráulicos – INDRHI (Water Resources)

	•		Secretaría de Estado de Industria y Comercio – SEIC (Ministry of Industry and Commerce)

	•		Subsecretaría de Recursos Forestales – SFR (Undersecretary of Forestry Resources)

	•		Dirección General de Minería – DGM (General Mining Agency)

	•		Ayuntamiento (Municipalities)

PDDC’s legal review indicates that there are 29 permits required to build and operate the Pueblo Viejo project. These are of three types: those needed before the environmental license, those included in the environmental license, and those needed after receipt of the environmental license.



Project No.: 147107		Page 4-11

Permits obtained to date include:

	•		Presentation of “Analisis Previo” Form.

	•		Definition and acceptance of Terms of Reference for the Environmental Impact Assessment, which was developed in conformance with local, PDI and World Bank Guidelines.

	•		Soil use permit for mining reserve and waterline corridor.

	•		Environmental authorization for soil use, mining reserve.

	•		Water supply permit for the Hatillo reservoir for mill water use.

The EIA is anticipated to be submitted to the Secretaria de Estado de Medio Ambiente y Recursos Naturales (SEMARN) during the fourth quarter of 2005. Upon submission of the EIA, other permit applications will be initiated.



Project No.: 147107		Page 4-12

5.0

ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE, AND PHYSIOGRAPHY

5.1

Climate and Physiography

			The central region of the Dominican Republic is dominated by the Cordillera Central mountain range, which runs from the Haitian border to the Caribbean Sea (Figure 4-1). The highest point in the Cordillera Central is Pico Duarte at 3,175 m. The Pueblo Viejo concession is located in the eastern portion of the Cordillera Central where local topography ranges from 565 m at Loma Cuaba to approximately 65 m at the Hatillo reservoir.

			Two rivers run through the concession, the Margajita and the Maguaca. The Margajita drains into the Yuna River upstream from the Hatillo reservoir, while the Maguaca joins the Yuna below. The flows of both rivers vary substantially during rainstorms.

			The Dominican Republic has a tropical climate with little fluctuation in seasonal temperatures, although August is generally the hottest month and January and February are the coolest. Temperatures at the project site range from daytime highs of 32°C to night-time lows of 18°C. Annual rainfall is approximately 1.8 m, with May through October typically being the wettest months. The Dominican Republic is in a hurricane channel; the hurricane season is typically August to November.

			As a result of previous mining and agriculture, there is little primary vegetation on the Pueblo Viejo site and surrounding concessions. Secondary vegetation is abundant outside of the excavated areas and can be quite dense. Rosario Dominicana, the previous owner of the concessions, also aided the growth of secondary vegetation by planting trees throughout the property for soil stabilization.

5.2

Access

			The mine site is approximately 95 km northwest of the national capital of Santo Domingo. Access is by a four-lane paved highway (Autopista Duarte), which is the main route between Santo Domingo and Santiago, the second-largest city of Dominican Republic. This highway connects to secondary Highway 17 at the town of Piedra Blanca, approximately 70 km from Santo Domingo. This secondary highway is a two-lane paved highway that passes through the towns of Piedra Blanca and Maimon on the way to Cotui. The gatehouse for the mine is 22 km from Piedra Blanca.

			The main port facility in the Dominican Republic is Haina in Santo Domingo. Other adequate port facilities are located at Puerto Plata, Boca Chica, and San Pedro de Macoris.



Project No.: 147107		Page 5-1

6.0

HISTORY

6.1

Pre-1969

			The earliest records of Spanish mine workings at Pueblo Viejo are from 1505, although Spanish explorers sent into the interior of the island during Columbus’ second visit in 1495 probably found the deposit being actively mined by the native population. The Spaniards mined the deposit until 1525, when the mine was abandoned in favour of newly discovered deposits on the American mainland.

			There are few records of activity at Pueblo Viejo from 1525 to 1950, when the Dominican Government sponsored geological mapping in the region. Exploration at Pueblo Viejo focused on sulphide veins hosted in unoxidized sediments in streambed outcrops. A small pilot-plant was built, but economic quantities of gold and silver could not be recovered.

6.2

Rosario/AMAX (1969-1992)

			During the 1960s several companies inspected the property, but no serious exploration was conducted until Rosario Resources Corporation of New York (Rosario) optioned the property in 1969. As before, exploration was directed first to the unoxidized rock, where sulphide veins crop out in the stream valley and the oxide cap is only a few metres thick. As drilling moved out of the valley and onto higher ground, the thickness of the oxide cap increased to a maximum of 80 m, revealing an oxide ore deposit of significant tonnage.

			In 1972 Rosario Dominicana S.A. was incorporated (40% Rosario, 40% Simplot Industries, and 20% Dominican Republic Central Bank). Construction of the mine started in 1973 and open pit mining of the oxide resource commenced on the Moore deposit in 1975. In 1979 the Dominican Central Bank purchased all foreign-held shares in the mine. Management of the operation continued under contract to Rosario until 1987. Rosario was merged into AMAX Inc. in 1980.

			Rosario continued exploration throughout the 1970s and early 1980s looking for additional oxide resources to extend the life of the mine. The Monte Negro, Mejita, and Cumba deposits were identified by soil sampling and percussion drilling and were put into production in the 1980s. Rosario also performed regional exploration, evaluating much of the ground adjacent to the Pueblo Viejo concessions with soil geochemistry surveys and percussion drilling. An airborne EM survey was flown over much of the Maimon Formation to the south and west of Pueblo Viejo.

			With the oxide resources diminishing, Rosario initiated studies of the underlying refractory sulphide resource in an effort to continue the operation. Feasibility-level studies were



Project No.: 147107		Page 6-1

conducted by Fluor Engineers Inc. in 1986 and Stone & Webster Engineering/American Mine Services (SW/AMS) in 1992.

			Fluor began its study of the resources at Pueblo Viejo with a detailed review and compilation of the drilling database. Rosario personnel had entered all the drill hole assay data into an electronic format. Fluor reviewed the drilling data from the different campaigns and used the drill holes that begin with DDH, R, RC, and RS but excluded the P series, which represented percussion drilling of the oxide portion of the deposit. Fluor discussed the data handling and checked the data, but gave no details of the data sources used for its work or the extent of verification that was completed.

			Fluor found that silver values were elevated at the sulphide/oxide interface area due to secondary enrichment. The analysis results for silver in these areas were changed to - -1.00 to negate their weight in the block model where a -1.0 code will result in the software ignoring these values. Based on sulphur content, bulk density was calculated for each assay according to the following formula: SG= 0.0322*%Sulphur+2.617. This formula was developed by linear regression between the SG and sulphur grade from 152 pairs of data. The assay data were composited to 10 m lengths, and values for gold, silver, sulphur, and carbon were interpolated into the model through ordinary kriging. The resources resulting from this modelling are historical and do not conform to current CIMM guidelines for Mineral Resource Classification. The resource was reported as a Geological Reserve without separation of the resource into categories and is summarized in Table 6-1.

Table 6-1: Historical Geological Reserve, Fluor 1985 (3.0 Au g/t cutoff)


	Tonnage		Au		Ag		S		C
Deposit	(t x 000)		g/t		g/t		%		%

Moore		44,200		3.881		19.92		8.36		0.57
Monte Negro		27,941		4.327		25.48		8.32		0.39

Total		72,143		4.053		22.07		8.35		0.50

			Fluor concluded that developing a sulphide project would be feasible if based on roasting technology, with sulphuric acid as a by-product. Rosario rejected this option due to environmental concerns related to acid production.

			SW/AMS prepared a feasibility study for a sulphide project, with a resource model using ordinary kriging (OK) and 20 m x 20 m x 10 m blocks. The drill hole data were composited to 10 m lengths, and a bulk density value of 2.62 was applied to all blocks for tonnage calculations. SW/AMS classified the resource into three categories of Proven, Probable, and Possible, the use of these categories is not in accordance with CIM Definitions but appear to be similar to the current categories of Measured, Indicated, and Inferred. Criteria for classification were based on the block centre to the nearest drill hole: Proven



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was 0 m to 15 m, Probable was 15 m to 30 m, and Possible was over 30 m. In the summary section of the SW/AMS feasibility study (the only section that was available for AMEC to review), the geological resources were tabulated with all classes grouped together. The historical geological resource was also reported at various cutoffs, as listed in Table 6-2.

Table 6-2: Historical Geological Resource Results, SW/AMS 1992


Cutoff	Tonnage
Au g	(t x 000)		Au g/t		Ag g/t

1.00		461,666		1.955		12.644
1.50		269,847		2.471		15.145
2.00		156,715		3.008		18.156
2.25		119,175		3.289		19.792
2.50		90,132		3.587		21.435
2.75		69,658		3.871		23.088
3.00		54,299		4.154		24.599
3.25		41,944		4.459		26.416
3.5		33,296		4.742		27.973

			SW/AMS concluded that a roasting circuit would be profitable at 15,000 t/d using limestone slurry for gas scrubbing and a new kiln to produce lime for gas cleaning and process neutralization.

			Rosario continued with mining of oxide material until approximately 1991, when the oxide resource was essentially exhausted. A CIL plant circuit and new tailings facility at Las Lagunas were commissioned to process “transitional” sulphide ore at a maximum of 9,000 t/d. Results were poor, with gold recoveries varying from 30% to 50%. Selective mining continued in the 1990s on high-grade ore with higher estimated recoveries. Mining in the Moore deposit stopped early in the 1990s owing to high copper content (which resulted in a high cyanide consumption) and ore hardness. Mining ceased in Monte Negro in 1998, and stockpile mining continued until July 1999, when the operation was shut down.

			In 24 years of production, the Pueblo Viejo mine produced a total of 5.3 Moz of gold and 24.4 Moz of silver.

6.3

Privatization (1996)

Lacking funds and technology to process the sulphide ore, Rosario Dominicana attempted two bidding processes to joint venture the property, one in about 1992 and the other in



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1996. In November 1996, Rosario selected Salomon Brothers (Salomon Smith Barney) to co-ordinate a process to find a strategic partner to rehabilitate the operation and to determine the best technology to economically exploit the sulphide resource. Three companies were involved in the privatization process: GENEL JV, Mount Isa Mines, and Newmont. Privatization was not realized, but each of the three companies conducted work on the property during their evaluations.

6.3.1

GENEL JV

The GENEL JV was formed in 1996 as a 50:50 joint venture between Eldorado Gold Corporation and Gencor Inc. (later Gold Fields Inc.) to pursue their common interest in Pueblo Viejo. The GENEL JV expended US$6 M between 1996 and 1999 in studying the project and advancing the privatization process. Studies included diamond drilling, developing a new geological model, mining studies, evaluation of refractory ore milling technologies, socio-economic evaluation, and financial analysis.

Geology-related work included:

	•		reviewing, copying, and digitizing the geological database

	•		preparing a geological resource model (by SRK Consulting, June 1996)

	•		collecting surface sulphide samples for pilot-plant and metallurgical studies

	•		conducting a 20-hole, 3,151 m diamond drilling program

	•		preparing a metallurgical sample from drill core

	•		surface mapping

	•		evaluating limestone resources

	•		assessing exploration potential

	•		completing a new geological resource (by Giroux Geostatistical Consultants, April 1997)

	•		initiating a new resource estimate (by MRDI, March 1998).

			In 1996, GENEL contracted SRK Consulting to estimate a resource model for the deposits. The database was the same one used by SW/AMS in its 1992 feasibility study. GENEL (on behalf of Rosario and completed by G.H. Giroux of Giroux Geostatistical Consultants) modified this model in 1997, but no documentation on the update was available for AMEC’s review.

			In 2001, Pincock, Allen & Holt (PAH) reviewed the 1997 model and had a number of comments and concerns. Initially, PAH found that block locations did not match the



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			existing block plans and sections and so reconfigured the blocks. PAH was also concerned that the block size, 20 m x 20 m x 10 m, while not inappropriate, did not adequately reflect short to intermediate grade variations and recommended that 10 m x 10 m x 10 m blocks be used in the future. PAH checked the data for 19 randomly selected drill holes, found no errors, and agreed with SW/AMS that the data were of good quality for resource estimation purposes. PAH reviewed the sample composite data, which were normalized to 10 m lengths with gold grades capped at 10 g/t Au, and found two grade populations for gold and zinc and only one for silver.

			PAH reviewed the geology model used for the grade estimation and found it to be too generalized for the complex nature of the deposit; a more comprehensive geology model was recommended for future work. When PAH inspected sections of the block model, plotted with the drill hole composites, the blocks were found to reasonably reflect the composite grades.

			It is of note that Giroux had used all the available drill hole data, including the ST series of holes rejected by subsequent modellers, for the 1996 resource model; the ST data set was excluded when the model was updated in 1997. Both the Giroux and GENEL models had been constructed by ordinary kriging.

			PAH created a nearest-neighbour model and compared the results to the kriged values in the 1997 model by GENEL. As would be expected, a general smoothing effect was found for the kriged grades relative to the nearest-neighbour grades, but no obvious areas of bias were found in the kriged model.

			GENEL used a fixed density value of 2.8 for its tonnage calculations. PAH reported 8% more tonnes when an estimated variable density was used to report the resource tonnage. PAH also found that zinc grades were 24% lower in its resource tabulation compared to GENEL’s. PAH could not resolve the difference satisfactorily.

			Giroux had originally classified the blocks based on the following criteria: Measured Resources for blocks within 0 to 15 m of a composite, Indicated Resources from 15 m to 100 m, and Inferred Resources at more than 100 m. PAH could not find any classification for the GENEL model and so used the following parameters to classify the resource estimate: Measured Resources were assigned to blocks when a composite was between 0 and 20 m from the block centre, Indicated Resources for ranges of 20 m to 40 m, and Inferred Resources for composites more than 40 m distance. The historical resource estimate results as classified by PAH are presented in Table 6-3.



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Table 6-3: Historical Resource Estimate, GENEL (Rosario) 1997 / PAH 2001
(2.0 g/t Au cutoff grade)


		Tonnage		Au		Ag		Zn		Cu		C		S
Deposit	Class	(t x 000)		g/t		(g/t)		(%)		(%)		(%)		(%)

Moore	Measured		35,316		3.09		19.47		0.82		—		0.46		8.38
	Indicated		54,208		2.98		18.70		0.63		—		0.48		8.02
	Total M&I		89,524		3.02		19.00		0.70		—		0.47		8.16
	Inferred		25,143		2.83		14.90		0.39		—		0.43		7.80

Monte Negro	Measured		22,193		3.23		18.68		0.78		0.05		—		—
	Indicated		32,694		3.05		17.68		0.77		0.06		—		—
	Total M&I		54,887		3.12		18.08		0.77		0.06		—		—
	Inferred		19,331		2.57		13.16		0.53		0.07		—		—

Total	Measured		57,509		3.14		19.17		0.80		—		—		—
	Indicated		86,902		3.01		18.32		0.68		—		—		—
	Total M&I		144,411		3.06		18.65		0.73		—		—		—
	Inferred		44,474		2.72		14.14		0.45		—		—		—

			In February 1998, GENEL contracted MRDI to complete a resource estimate for the Pueblo Viejo Project based on data provided by GENEL. The database was used as received, with no review of the QA/QC data for the drilling or work done to check the validity of the data. Two geological models were provided, one for Moore and one for Monte Negro, both of which were used without validation. GENEL also specified the pit limits (El Dorado pit) for the resource summary. The database contained 399 drill holes (35,154 m) from Monte Negro and 537 drill holes (45,515 m) from Moore. MRDI estimated the resource by multiple indicator kriging (MIK) with 14 grade classes (13 thresholds) and 10 m fixed-length composites, and used a minimum of 9 and maximum of 64 composites for gold kriging. An indirect lognormal approach was used for the change of support, with an SMU of 10 m x 10 m x 10 m. Density values were derived from the formula currently adopted by Placer Dome.

			MRDI’s resource estimate is summarized in Table 6-4.

			PDDC purchased the GENEL data files from Eldorado Corporation in November 2002.



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Table 6-4: Historical Measured + Indicated Resource Estimate,
El Dorado Pit, MRDI 1998


Cutoff		Tonnes		Au		Ag		Cu		Zn		S		C
Au g/t	Deposit	(Mt)		g/t		g/t		%		%		%		%

1.0	Monte Negro		65.0		2.92		18.23		0.064		0.65		6.94		0.34
	Moore		145.9		2.59		19.90		0.122		0.68		7.72		0.48

	Total		210.9		2.69		19.39		0.104		0.67		7.48		0.44

1.2	Monte Negro		61.4		3.03		18.96		0.066		0.66		6.98		0.34
	Moore		136.8		2.70		20.73		0.126		0.69		7.76		0.48

	Total		198.1		2.80		20.18		0.107		0.68		7.51		0.44

1.5	Monte Negro		54.8		3.24		20.38		0.070		0.66		7.01		0.34
	Moore		119.0		2.90		22.44		0.136		0.72		7.81		0.48

	Total		173.8		3.01		21.79		0.115		0.70		7.56		0.44

2.0	Monte Negro		42.8		3.67		23.41		0.074		0.67		7.10		0.35
	Moore		90.4		3.27		25.65		0.155		0.76		7.88		0.48

	Total		133.2		3.40		24.93		0.129		0.73		7.63		0.44

2.5	Monte Negro		31.7		4.16		27.08		0.078		0.68		7.24		0.36
	Moore		62.2		3.73		29.91		0.177		0.80		7.96		0.49

	Total		93.9		3.87		28.95		0.144		0.76		7.72		0.45

6.3.2

MIM

In 1997, Mount Isa Mines (MIM) conducted a due diligence program as part of its effort to win Pueblo Viejo in the privatization process. MIM proposed to carry out a pilot-plant and feasibility study using ultra fine grinding/ferric sulphate leaching.

Geology-related work carried out by MIM included:

	•		conducting a 31-hole, 4,600 m diamond drilling program

	•		collecting a metallurgical sample from drill core

	•		detailed pit mapping by geological consultant Carl Nelson

	•		completing IP geophysics over the known deposits

	•		organizing aerial photography over the mining concessions to create a new (1997) surface topography (shared expense with GENEL).



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6.3.3

Newmont

In 1992, and again in 1996, Newmont proposed to carry out a pilot-plant and feasibility study for ore roasting/bioheap oxidation. Newmont collected samples for analysis, but no results are available. Both of Newmont’s attempts to privatize or joint venture the property failed.



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7.0

GEOLOGICAL SETTING

7.1

Regional Geology

			Pueblo Viejo is hosted by the Lower Cretaceous Los Ranchos Formation, a series of volcanic and volcaniclastic rocks that extend across the eastern half of the Dominican Republic, generally striking northwest and dipping southwest. The Los Ranchos Formation consists of a lower complex of pillowed basalt, basaltic andesite flows, and dacitic flows, tuffs, and intrusions, which are overlain by volcaniclastic sedimentary rocks and is interpreted to be a Lower Cretaceous intraoceanic island arc; one of several bimodal volcanic piles that form the base of the Greater Antilles Caribbean islands. The unit has undergone extensive seawater metamorphism (spilitization), and lithologies have been referred to as spilite (basaltic-andesite) and keratophyre (dacite).

			The Pueblo Viejo Member of the Los Ranchos Formation is confined to a restricted sedimentary basin measuring approximately 3 km north-south by 2 km east-west. The basin is interpreted to be either due to volcanic dome collapse causing a lake or a maar-diatreme complex that cut through lower members of the Los Ranchos Formation. The basin is filled with lacustrine deposits that range from coarse conglomerate deposited at the edge of the basin to thinly bedded, carbonaceous sandstone, siltstone, and mudstone deposited further from the paleo-shoreline. In addition, there are pyroclastic rocks, dacitic domes, and diorite dykes within the basin. Plant fossils collected from the sedimentary basin and surrounding volcanic debris flows are considered to be of Neocomian age (121 to 144 Ma). The Pueblo Viejo Member is bounded to the east by Zambrana Member fragmental volcaniclastic rocks and to the north and west by Platanal Member basaltic-andesite (spilite) flows and dacitic domes.

			To the south, the Pueblo Viejo Member is over thrust by the Hatillo Limestone Formation by means of a low-angle, southwest-dipping thrust fault. Fossils collected from the base of the limestone indicate that the assemblage is likely Cenomanian (93 to 99 Ma) or possibly Albian (99 to 112 Ma) in age.



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Figure 7-1: Regional Geology


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Figure 7-2: Property Geology


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7.2

Local Geology – Moore and Monte Negro Deposits

7.2.1		Lithology and Stratigraphy

		*Moore*

		The Moore deposit is located at the eastern margin of the Pueblo Viejo Member sedimentary basin. Stratigraphy consists of finely bedded carbonaceous siltstone and mudstone (PV Sediments) overlying horizons of spilite (basaltic-andesite flows), volcanic sandstone, and fragmental volcaniclastic rocks. The entire sequence has a shallow dip to the west (Figures 7-3 and 7-4).

		Figure 7-3: Moore Section 94600N – Lithology (Looking North)


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Figure 7-4: Moore Section 94850N – Lithology (Looking North)

		Fragmental Dacite Porphyry (FDP) that outcrops north of the plant site intrudes the stratigraphic sequence. FDP is best described as a vent breccia with a volcaniclastic appearance (quartz eyes and lithic fragments) and intrusive phases such as local breccia dykes and intrusive contacts. Propylitically altered porphyry has been intersected in core with intrusive textures and appears to form a NNE-striking root-zone to the FDP. FDP appears to have been emplaced prior to mineralization with local zones of disseminated pyrite and anomalous gold mineralization. The eastern margin of the sedimentary basin hosting the Moore deposit is defined by fragmental volcaniclastic rocks (Zambrana Member) and non-carbonaceous sedimentary rocks (Mejita Sediments).

		There are indications that an internal sub-basin exists at Moore below the PV Sediments. The sub-basin is partially filled with a mixed sedimentary sequence consisting of interfingering PV Sediments and fragmental volcaniclastic rocks. Graded bedding and slump folding textures are often observed in core. The south and west margins of the sub-basin are defined by pinching of the Spilite and Volcanic Sandstone horizons (Figures 7-3 and 7-4).


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		*Monte Negro*

		The Monte Negro deposit is located at the northwestern margin of the sedimentary basin (Figure 7-2). Stratigraphy consists of interbedded carbonaceous sediments ranging from siltstone to conglomerate that are interlayered with volcaniclastic flows. These volcaniclastic flows become thicker and more abundant towards the west. This entire sequence has been grouped as the Monte Negro (MN) Sediments. In the eastern part of the Monte Negro deposit area, the bedding dip is shallow to the southwest; in the west, the dip is shallow to the northwest.

		The Monte Negro Sediments overlie a horizon of spilite and spilite-derived conglomerate (Figure 7-5). The conglomerate consists of pebble to boulder-size clasts of spilite that are often silicified and a light pink colour. Silicification is likely volcanogenic, occurring prior to the sedimentation of the basin. The conglomerate horizon represents either a basal conglomerate channelled into the margin of the basin or a reworked brecciated flow-top of the spilite below. The horizon ranges in thickness from tens of metres to non-existent and is likely filling channels in the uneven spilite surface below.

		Spilite that forms the basement of the Monte Negro deposit is the Platanal Member of the Los Ranchos Formation. In the west part of the deposit, porphyritic textures, and massive andesitic flows, often separated by brecciated flow-tops, are observed. The brecciated textures become more abundant towards the east.

		Thin-section work on the porphyritic spilite indicates a composition of either a high-silica andesite or a low-silica dacite. Primary textures observed are consistent with an intrusion, indicating that either a dome or a near-surface plug may exist under the west hill of Monte Negro. The dimensions of this possible intrusive have not been determined because core drilling is limited. Dykes that intrude the Monte Negro stratigraphy include a steeply dipping NNW-striking mafic (diorite/andesite) dyke approximately 10 m wide. The dyke typically follows the F5-fault through the deposit area but occasionally splays to the north. The dyke is propylitically altered and is barren of gold mineralization. Similar dykes have been intersected in core in the west part of the deposit, but they are much thinner. Thin breccia dykes (pebble dykes) have also been mapped in the pit walls.


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Figure 7-5: Monte Negro Section 95800N – Lithology (Looking North)

7.2.2		Laboratory Studies

		Several investigations were undertaken to gain a better understanding of the hydrothermal fluids and mineralization at Pueblo Viejo. These included studies in deportment of gold, petrography, and X-ray diffraction. The deportment of gold is discussed in Section 9.2.1 of this report.

		*Petrography*

		Nine drill core and surface rock samples were sent to Mason Geoscience Pty of Greenhill, Australia, for petrographic and mineralogical descriptions. Subsequently some of the samples were forwarded to Amdel Limited (Thebarton, Australia) for X-ray diffraction analysis. The purpose of these studies was to:

	•		describe and identify the primary rock types

	•		determine the paragenetic evolution of wallrock alteration and veining

	•		consider why mineralization changed from replacement style to vein style


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	•		determine the genesis of silicification

	•		describe the nature of the Moore Fragmental Dacite Porphyry

	•		identify minerals that could potentially be used for dating.

		This work provided some of the basis for the present geological model of mineralization.

		Placer Dome has a reasonable understanding of the geological setting of mineralization has developed reasonable geological models to support resource estimates.


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8.0		DEPOSIT TYPES

		The Pueblo Viejo Project is classed as a high-sulphidation, epithermal gold and silver deposit. Similar deposits occur at Summitville, Colorado, El Indio, Chile, Lepanto, Philippines, and Goldfield, Nevada. These deposits are characterized by veins, vuggy breccias, and sulphide replacements ranging from pods to massive lenses, generally in volcanic sequences associated with high-level hydrothermal systems. Acid leaching, advanced argillic alteration, and silicification are characteristic alteration styles in these deposits. Grade and tonnage varies widely with this type of deposit. Pyrite, gold, electrum, and enargite/luzonite are typical minerals in these deposits. Minor minerals include chalcopyrite, sphalerite, tetrahedrite/tennantite, galena, marcasite, arsenopyrite, silver sulphosalts, and tellurides (Panteleyev, A., 1996).

		Two distinct types of mineralization are included in this classification; enargite-pyrite (El Indio for example) and quartz-alunite-gold (Pueblo Viejo). These deposits are typically related to high-level intrusives.

		The geological setting of the deposit is controversial at this time. Sillitoe and Bonham (1984), Muntean and others (1990), Kesler and others (2005) have described the setting as a maar diatreme complex with the various deposits around the margins of the diatreme. The coarse grained fragmental rocks that occur at depth in the deposit are considered by these investigators to be the product of an explosive volcanic eruption that fragmented the rocks and partially filled the crater. The crater was then completely filled with shallow marine sedimentary rocks with variable amounts of fragmental rocks from nearby volcanoes. This sequence was then crosscut by later dikes and small dacite and andesite lava domes.

		Nelson (2000) describes the setting as volcanic dome complex emplaced in a shallow marine environment and attributes the coarse fragmental rocks to collapsing carapaces on those domes. The sedimentary rocks were deposited in depressions between the domes.

		In both cases, mineralization was controlled by structures that controlled emplacement of the lava domes.

		This controversy is largely academic at this point in time and has no practical impact on exploration of the deposit at the levels that may be mined by open pit methods. The aerial extent of the deposits has been defined by drilling and the vertical extents are reasonably well known, although additional drilling is required to define the deepest parts of the deposit. Exploration within the area of known mineralization will be following previously identified trends.


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9.0

MINERALIZATION

9.1		General Description

		Metallic mineralization in the deposit areas is predominantly pyrite with lesser amounts of sphalerite and enargite. Pyrite mineralization occurs as disseminations, layers, replacements, and veins. Sphalerite and enargite mineralization is primarily in veins, but disseminated sphalerite has been noted in core.

		Studies have determined that there were three stages of advanced argillic alteration associated with precious metal mineralization. Stage I alteration produced alunite, silica, pyrite, and deposited gold in association with disseminated pyrite. Stage II overprinted Stage I and produced pyrophyllite and an overlying silica cap. The third stage of mineralization occurred when hydro-fracturing of the silica cap produced pyrite-sphalerite-enargite (Stage III) veins with silicified haloes. Syntaxial vein growth preserves evidence for pyrite ® enargite ® sphalerite ® grey silica paragenesis.

		Individual Stage III veins have a mean width of 4 cm and are typically less than 10 cm wide. Exposed at surface, individual veins can be traced vertically over three pit benches (30 m). Veins are typically concentrated in zones that are elongated NNW and can be 250 m long, 100 m wide, and 100 m vertical. Stage III veins contain the highest precious and base metal values and are more widely distributed in the upper portions of the deposits.

		Veins tend to be parallel to follow a number of local structures that crosscut the deposit. Those structures have a northerly trend at Monte Negro and Moore with a northwest-southeast also present at Moore.

		The most common vein minerals are pyrite, sphalerite, and quartz with lesser amounts of enargite, barite, and pyrophyllite. Trace amounts of electrum, argentite, colusite, tetrahedrite-tennantite, geocronite, galena, siderite, and tellurides are also found in veins.

		The abundance of pyrite and sphalerite within veins varies across the deposit areas. Veins in the southwest corner of the Monte Negro pit are relatively sphalerite rich and pyrite poor when compared to veins elsewhere in the Moore and the Monte Negro deposits. The sphalerite in these veins is darker red in colour, possibly indicating that it is richer in iron. The abundance of dark red sphalerite in these veins may also be indicative of the outer margins of a system of hydrothermal-magmatic ore-bearing fluids.

		Late massive pyrophyllite veins, that probably represent the last stage of veining and alteration, cut the Stage III veins. All stages of veining are cut by thin white quartz veins associated with low-angle thrusts that post-date mineralization.


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9.2

Metal Occurrence and Distribution

9.2.1		Gold

		Gold is intimately associated with pyrite veins, disseminations, replacements, and layers within the zones of advanced argillic alteration. Gold values generally are the highest in zones of silicification or strong quartz-pyrophyllite alteration. These gold-bearing alteration zones are widely distributed in the upper parts of the deposits and tend to funnel into narrow feeder zones.

		Stage III sulphide veins also have higher gold values than replacement-style mineralization. In the Moore deposit, a high-grade structural feeder zone within an alteration funnel was intersected by a GENEL core hole (GEN_MDD6). The hole intersected an intensely silicified shear zone that returned gold values of 9.1 g/t over 40 m (30 m true width). The shear is steeply dipping and appears to strike either N or NW. While the shear is open to depth, it possibly has a strike-length of less than 100 m. This style of mineralization differs from the upper zones of the deposit, where high-grade gold is associated with sulphide veins. This feeder zone also contains a higher concentration of lead (Pb) in the form of Pb-sulphosalts and galena.

		In the Monte Negro deposit, a high-grade feeder zone has not been identified. A potential target is the Monte Negro Fault, which is intensely silicified and bounds high-grade mineralization at surface. A second possibility is a deeper zone of mineralization intersected by MIM (Hole MIM_MN013). There, the vertical core hole was testing an IP-chargeability anomaly approximately 100 m east of the main deposit. This zone of mineralization may represent an east-dipping feeder structure following the base of the sedimentary basin.

		The most common form of gold, is in a submicroscopic form within pyrite, where it is present as both solid solution within the crystal structure of the pyrite and as colloidal-size micro inclusions (<0.5 µm). The quantity of gold depends on the morphological and textural form of the pyrite.

		AMTEL Laboratories (London, Ontario) conducted a study to establish the deportment of gold in four separate composites from Pueblo Viejo. These composites represented four of the five metallurgical rock-types established at for the deposit: Sedimentary rocks (MN-BSD) and volcanic rocks (MN-VCL) at Monte Negro and sedimentary rocks (MO-BSD) and volcanic rocks (MO-VCL) at Moore. Spilites at Monte Negro were not sampled. The objectives of this study were to:

	•		identify all forms of gold

	•		characterize free gold in terms of grain size, composition, and association


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	•		determine the submicroscopic gold content of the different morphological types of pyrite and other sulphides

	•		determine the gold content of the binaries, size by size

	•		determine the gold content of the rock mineral fraction, size by size

	•		determine the gold in the slimes (<5 µm) fraction

	•		establish the deportment of gold based on the acquired data

	•		characterize sphalerite in terms of liberation and associated gold.

		Procedures used in this study included using gravity separation methods on sized fractions to pre-concentrate free gold and separate sulphide from binary and mineral particles; electron probe microanalysis (EPMA) to determine the composition of gold; and ion probe microanalysis (SIMS) to quantify the submicroscopic gold content of the different morphological types of pyrite and other sulphides.

		Gold occurs as native gold, sylvanite [AuAgTe₄], and aurostibite [AuSb₂]. The principal carrier of gold is pyrite, where the submicroscopic gold occurs in colloidal-size microinclusions (<0.5 µm) and as a solid solution within the crystal structure of the pyrite. The abundance of the gold minerals (native gold, sylvanite, aurostibite) varies significantly between the different composites (see Table 7-1).

		Microcrystalline pyrite particles with a contain the highest concentration of gold. This type of pyrite is also the most arsenic-rich, which renders it the most prone to oxidation. It is also the most difficult to liberate, as it forms complex intergrowths within the rock and with sphalerite.

		All significant gold values carried by sphalerite are due to intergrowths of pyrite that carry submicroscopic gold.

		Table 9-1: Mineralogically Determined Deportment of Gold


Form and Carrier	MN-BSD		MN-VCL		MO-BSD		MO-VCL
Gold Minerals	(%)		(%)		(%)		(%)

Free gold		1.8		25.3		22.4		68.6
Free sylvanite		20.6		0.8		5.1		—
Free aurostibite		—		—		—		0.2
Rock-sulphide binaries		8.3		13.1		8.5		0.9
‘Clean’ rock		4.0		6.1		9.6		0.2

Submicroscopic Gold
Microinclusions		51.7		33.4		28.0		19.8
Solid solution		13.6		21.5		26.4		10.3


Project No.: 147107		Page 9-3

		Studies have shown that there are four major forms of pyrite: microcrystalline, disseminated, porous, and coarse grained. The microcrystalline pyrite tends to have the highest gold concentration. This morphological form of pyrite also has the highest arsenic concentration, making it the most susceptible to oxidation. The coarse-grained form of pyrite has the lowest gold concentration and also has a well-developed crystal habit making it less susceptible to oxidation.

		A less common form of gold is gold minerals such as native gold, electrum, tellurides (sylvanite, calaverite, petzite), and locally aurostibite. Most grains are less than 10 µm in diameter and are largely associated with growth zones of pyrites. Gold minerals are also found to a lesser extent as inclusions in pyrite, enargite, quartz, and Pb-sulphosalts (primarily geocronite). Gold may also exist in the crystal structure of sulphosalts such as enargite and geocronite, but additional research is required.

		The proportions of the different forms and carriers of gold vary significantly throughout the Moore and Monte Negro deposits. Generally the majority of the gold is found as submicroscopic gold in microcrystalline, disseminated, or porous pyrite. There are local exceptions where gold tellurides or native gold are a significant contributor to the quantity of gold. The distribution of these ‘local exceptions’ is not truly understood at this time.

		While there is a strong correlation between gold and zinc, and zones with sphalerite veins tend to have the highest gold grades, sphalerite carries gold only as intergrowths of gold-bearing pyrite. The quantity of gold carried by the sphalerite depends on the percentage of gold-bearing pyrite encapsulated and the amount of submicroscopic gold within the pyrite.

9.2.2		Silver

		Assays for silver consistently have the strongest correlation with gold. Silver has a strong association with Stage III sulphide veins where it occurs as the native silver, Sb-sulphides (pyrargyrite), Ag-tellurides (hessite), Au, Ag-tellurides (sylvanite, petzite), and Ag-bearing tetrahedrite.

9.2.3		Zinc

		The majority of the zinc occurs as sphalerite, primarily in Stage III sulphide veins and secondarily as disseminations. The sphalerite is beige to orange coloured and is relatively iron free. An exception is the dark red veins found in the southwest corner of the Monte Negro deposit that may represent a discontinuous halo surrounding the alteration zone.

		Sphalerite commonly contains inclusions and intergrowths of pyrite, sulphosalts, galena, and silicate gangue. The encapsulated pyrite is often host to submicroscopic gold mineralization.


Project No.: 147107		Page 9-4

Trace amounts of zinc can be found in zinc-bearing tetrahedrite and enargite.

9.2.4		Copper

		Most of the copper occurs as enargite hosted in Stage III sulphide veins. Only trace amounts of chalcocite and chalcopyrite have been recorded in SEM work. Enargite-rich vein zones typically are confined laterally and vertically within the larger sphalerite-rich vein zones. The enargite is difficult to identify in hand-specimen and is easily confused with tennantite-tetrahedrite.

9.2.5		Lead

		Lead minerals include galena, geocronite, boulangerite, and bournonite, most of which are present as fine inclusions or within fractures in pyrite, sphalerite, and enargite. Geocronite and boulangerite are the most prevalent.

		There are a limited number of lead assays in the database. Assaying completed by GENEL shows a strong correlation between gold and lead. Elevated lead values were found in the structural feeder zone that GENEL intersected in the Moore deposit, and lead may provide clues on where to search for other feeder zones.

		Outside of the deposit areas, saprolite (where rock textures are pseudomorphed by clay) is as much as 25 m thick in the valleys but is negligible on the hilltops. Fresh rock and partially clay-altered rock can often be found on the tops of hills.

9.3

Mineralization

9.3.1		Moore

		*General*

		Bedding generally dips shallowly westwards (<25°), but locally, steep faults with NNE and NNW strikes have rotated bedding into steep orientations. The NNE faults preserve evidence for east-side-up and left-lateral sense of movement subsequent to mineralization. The NNE faults appear to link with a NNW-trending fault that controls the eastern margin of the Moore dacite porphyry and is a boundary to a gold-bearing pyrite vein zone at North Hill (Figure 9-1). Westward-dipping thrust and bedding plane faults offset pyrite veins but with only minor displacement. They are associated with an intense cleavage and bedding-parallel quartz veins with gold mineralization.


Project No.: 147107		Page 9-5

Figure 9-1 Moore Deposit Structure and Mineralization


Project No.: 147107		Page 9-6

		Pyrite-rich, gold-bearing veins have a mean width of 4 cm and are steeply dipping with a trend commonly NNW (~330°). Sub-dominant pyrite vein-sets trend N-S and NNE. The orientation of pyrite veins and steep faults is similar, albeit with different dominant sets (NNW for veins and NNE for faults). This indicates a probable genetic link between steep faulting and vein development.

		*West Flank Zone*

		Thinly bedded carbonaceous siltstones and andesitic sandstones in the West Flank dip shallowly westwards (280°/-16°). Dips increase towards the west, where northerly trending thrusts displace bedding (Figure 9-2).

		Pyrite and limonite-rich veins with gold mineralization are sub-vertical and trend commonly NNW (239°/-81°). The veins are oblique to the general NNE strike of bedding and do not appear to have been rotated. Quartz veins with gold trend NW (~303°) oblique to the pyrite veins but have a similar strike to the interpreted contact with the overlying Hatillo limestone. They also occur as tension-gash arrays in centimetre-scale dextral shear zones that trend NNW.

		Faults create centimetre-scale displacement of bedding, and pyrite-sphalerite veins occur along steep NNE-trending faults and westerly dipping thrusts. Two main NNE faults were mapped across the West Flank, sub-parallel with the Moore dacite porphyry contact. Displacement of veins preserves evidence for a lateral sinistral component of movement.

		*North and South Hills Zones*

		Bedding to the north of the Moore dacite porphyry dips shallowly westwards (344°/-24° to 210°/-18°). Bedding has been rotated about both NNW and NNE axes. The change in bedding orientation reflects movement associated with NNW and NNE trending faults.

		There are three gold-bearing, pyrite-rich vein sets: NW (230°/-82°), NE (296°/-80°), and N-S (086°/-85°). NW-trending veins generally contain enargite and sphalerite, while NE-trending veins are more pyrite ± pyrophyllite rich. The average vein width is 3.5 cm.

		The fault pattern is dominated by steep NNE-trending faults that appear to link with NNW-trending faults (Figure 9-1). A NNE-trending steep fault along the western margin of the Moore dacite breccia has rotated bedding from shallow to steep dips, indicating an east-side up sense of movement. The sense of movement along NNW faults could not be determined. Bedding parallel thrusting is common and is evidenced by intense cleavage and quartz veins parallel to bedding. Bedding plane displacement is minor, generally less than 20 cm.


Project No.: 147107		Page 9-7

Figure 9-2: Alteration and Mineralization, Section 94600 N, West Flank and Vein Zone, Moore Deposit


Project No.: 147107		Page 9-8

9.3.2		Monte Negro

		*General*

		Interbedded carbonaceous siltstones, sandstones, and volcanic rocks in the Monte Negro Central Zone generally dip shallowly (19°) towards the southwest. In the Monte Negro South Zone (Figure 9-3), andesitic volcanic and volcaniclastic rocks generally dip shallowly (13°) towards the northwest. A steep NNW-trending fault (Monte Negro Fault) with a west-side up sense of movement is interpreted to separate the sediments in the east from the volcanic rocks in the west. The fault is interpreted to have been a focus for silicification, breccia dyke emplacement, and mineralization.

		Bedding in the hanging and footwalls of the Monte Negro Fault has been folded into upright open folds in close proximity to the fault. The axial trace of the folds trends NNW sub-parallel to the strike of the NNW conjugate vein set.

		Thrust faults displace veins and have brought sedimentary rocks into contact with andesitic volcanic and volcaniclastic rocks. A disconformable thrust contact is well exposed at the southern end of Monte Negro West.

		*Monte Negro Central Zone*

		Along the western margin of the main Monte Negro pit are thinly bedded carbonaceous siltstones, andesitic sandstones, and andesitic flows that dip shallowly towards the southwest (225°/-19°). Towards the centre of the pit, bedding has been folded into a series of shallowly NNW plunging open folds.

		Pyrite-rich veins with gold mineralization are subvertical and have bimodal trends, which are interpreted to form conjugate sets. The mean width is 2 cm. The NNW-trending set (244°/-85°) is sub-parallel to the strike of bedding and fold axes, indicating a possible genetic relationship between folding and mineralization. Enargite- and sphalerite-bearing veins with gold dominantly trend NNE (303°/-86°) and have a mean width of 3 cm. The combination of vein trends forms a high-grade gold zone (Vein Zone 1) which extends 500 m NNW, and is 150 m wide and up to 100 m thick between the F5 Fault to the east and the Main Monte Negro Fault to the west (Figures 9-3 and 9-4).

		The fault pattern is dominated by steep NNW-trending faults sub-parallel to the dominant pyrite vein set. The main Monte Negro Fault is a zone of silicification, brecciation, mineralization, folding, and faulting approximately 25 m wide and 500 m long. It is interpreted as a major fault that was active during and subsequent to mineralization.


Project No.: 147107		Page 9-9

Figure 9-3 Monte Negro Deposit Structure and Mineralization


Project No.: 147107		Page 9-10

Figure 9-4: Cross Section of Alteration and Mineralization, Monte Negro Central Zone at 95650N (Looking North)


Project No.: 147107		Page 9-11

		*Monte Negro South Zone*

		Andesitic volcanic and volcaniclastic rocks with minor intercalations of carbonaceous sediments dip shallowly northwards (331°/13° to 055°/29°). Close to the interpreted Monte Negro Fault, bedding dips more westerly and strikes NNW.

		NNW-trending steep faults displace bedding and dip towards the southwest. Displacement of marker agglomerate beds indicates a metre-scale west-side up sense of movement. The faults are sub-parallel to the interpreted Monte Negro Fault, which also has an apparent west-side up sense of movement.

		Mineralized veins at the Monte Negro South Zone are relatively pyrite poor, sphalerite rich, and wider (5 to 6 cm). The veins are subvertical and trend NW. The episodic vein fill demonstrates a clear paragenesis (massive pyrite ® enargite ® sphalerite ® grey silica).

		Shallow-dipping bedding and subvertical sphalerite-silica veins on the southern margin of Monte Negro South are cut by a westerly dipping thrust. The thrust has brought thinly bedded pyritic sedimentary rocks into contact with andesitic volcanic and volcaniclastic rocks. The fault dips 35° and was mapped across the top of the Monte Negro South hill. The overthrust sedimentary rock package contains asymmetric folds and bedding-cleavage relationships that indicate a reverse (west-side up) sense of movement. An upper thrust has brought a massive volcanic unit into contact with the underlying folded sediments.

		The main zone of gold mineralization that results from this combination of structures extends for approximately 150 m along the West Thrust Fault (Figure 9-3).

9.4		Hydrothermal Alteration

		The Pueblo Viejo District comprises several high-sulphidation (or acid-sulphate) epithermal deposits of which Moore and Monte Negro are the largest (Figure 4-2). The deposits form funnel-shaped envelopes of advanced argillic alteration where hydrothermal fluids migrated upwards and laterally along permeable horizons. Alteration zones are typically characterized by silica, pyrophyllite, pyrite, kaolinite, and alunite. Silica is predominant in the core of the alteration envelope and occurs with kaolinite in the upper zones where a silica cap is often formed. Unlike typical high-sulphidation deposits, where silicic alteration is residual and a result of acid leaching, silicification at Pueblo Viejo represents silica introduction and replacement. Silica-enriched zones are surrounded by a halo of quartz-pyrophyllite and pyrophyllite alteration.

		Advanced argillic alteration is easily distinguished from the chlorite-albite-calcite-epidote assemblage typical of the seawater metamorphosed (spilitized) Los Ranchos Formation.


Project No.: 147107		Page 9-12

		Limits of the alteration zones are marked by a rapid change (over a few metres) in mineralogy. Outside of alteration zones, finer-grained sedimentary rocks are pyritic (framboids) or sideritic with diagenetic conditions suggesting an anoxic, restricted basin. Within orebodies, siderite is completely replaced by pyrite.

		In the Moore deposit, silica, and kaolinite are more common in the upper parts of the system. In the now-depleted oxide ore, silicification was closely associated with gold mineralization and caused ore zones to form hills with relief of about 200 m (Figure 9-2). In areas of intense silicification, jasperoid masses were produced, original sedimentary textures completely destroyed and carbonaceous material removed. Locally, veins and masses of pyrophyllite cut the jasperoid bodies.

		In Monte Negro, silica and kaolinite are again more abundant in the upper portions of the deposit, and a silica cap is present (Figure 9-4). Silicification is more widespread at Monte Negro and not as closely associated to gold mineralization. Regardless, gold content is typically higher in silicified or partially silicified (quartz-pyrophyllite) rock.

9.5		Weathering

		Past mining operations have stripped the deposit areas of any surface oxidation, and the oxide mineralization is now depleted. The oxide was formed where surface oxidation removed sulphide minerals and carbon from the host sediments, leaving silicified host rock and massive jasperoid with jarosite, goethite, and local hematite mineralization. The thickness of the oxide mineralization ranged from 80 m at North Hill in the Moore deposit to 50 m in the South Hill and East Mejita deposits to nothing in the stream valleys. The thickest oxide ore was developed in intensely silicified, thinly bedded, and well-fractured sedimentary rocks. In contrast, areas underlain by intensely pyrophyllitized sedimentary rocks only had a few metres of oxidation. Soil cover and saprolite were negligible over the oxide mineralized zones.

		Gold mineralization was largely immobile in the oxide mineralization. No gold enrichment occurred, but free gold existed. Fine specks of gold (<100 µm) could be panned from only the highest-grade zones. Silver was depleted in the near-surface parts of the oxide ore and enriched at the oxide-sulphide interface. Zinc and copper were leached from the oxide with the destruction of the sulphides.

9.6		Ore Controls Used in Resource Estimates

		The primary controls on the geometry of the gold deposits at Pueblo Viejo are strong quartz-pyrophyllite alteration and quartz-pyrite veining along subvertical structures and stratigraphic zones. The stratigraphic shape of some zones may be controlled by subhorizontal structures that contain pyrite veins. The veins are tens of centimetres wide


Project No.: 147107		Page 9-13

		but are most commonly less than 2 cm wide. Narrow veinlets occur along bedding planes and along fracture surfaces. These veins are commonly highly discordant to bedding, but locally branch out along shallow-dipping bedding planes, linking high-angle veins in ladder-like fashion without obvious preferred orientations. These veins served as feeders to the layered and disseminated mineralization that occurs in shallower levels in the deposit. The result is composite zones of mineralization within fracture systems and stratigraphic horizons adjacent to major faults which served as conduits for hydrothermal fluids (Figures 9-1 to 9-4).

		Gold is intimately associated with the pyrite veins, disseminations, replacements, and layers within the zones of advanced argillic alteration. Gold values generally are the highest in zones of silicification or strong quartz-pyrophyllite alteration. Sphalerite is largely restricted to the veins, with pyrite lining the vein walls and sphalerite occurring as botryoidal aggregates. Galena, enargite, and boulangerite occur in small quantities in the center of the veins.

		These gold-bearing alteration zones are widely distributed in the upper parts of the deposits and tend to funnel into narrow feeder-zones at depth. Mineralization is generally contained within the boundaries of advanced argillic alteration. The outer boundary of advanced argillic alteration, combined with lithological and veining zones, were used to generate domains for resource estimation. Recent exploration by PDDC has used this model

		PDDC’s understanding of ore controls is reasonably detailed and suitable to support resource estimates.


Project No.: 147107		Page 9-14

10.0		EXPLORATION

		In addition to drilling programs in 2002 and 2004 (see Section 11), PDI has also completed the following studies.

10.1		Structural Pit Mapping

		Tony Norman of GeoCentric Exploration conducted mapping of the Moore and Monte Negro open pits in October and November 2002. The focus of the mapping was to develop a structural framework that could be used for a resource model and to identify the main controls on mineralization. Mapping was undertaken at 1:1,000 and 1:500 scales and compiled onto 1:1,000 factual geology sheets. An Excel database was created that consists of 1,000 structural measurements and contains co-ordinates and descriptions for all structural data. Interpretation of bedding trend lines, fault traces, vein zones, and alteration zones were compiled onto 1:1,000 scale overlay interpretative maps. These maps were digitized into Mapinfo and converted into ArcGis shape files.

10.2		Regional Mapping and Sampling

		As a part of the ongoing environmental baseline study at Pueblo Viejo, PDDC conducted mapping and sampling over a 105 km² area around the concessions. The main objective of this program was to identify acid rock drainage (ARD) sources outside of the main deposit areas. Mapping was conducted at 1:5,000 scale, and traverses were completed by following stream drainages. A total of 42 stream sediment samples and 125 rock samples were collected and sent for gold and trace element analysis. ALS Chemex from Vancouver, Canada, conducted all of the analyses.

10.3		Tailings Evaluation

		Part of the regional mapping program was dedicated to evaluating the potential for mineralization in the proposed El Llagal tailings storage area. Mapping and stream sediment sampling were conducted in the El Llagal valley and in the adjacent Maguaca and Naranjo river valleys. Stream sediment samples were collected approximately every 300 m.

		BGC Engineering of Vancouver, Canada, drilled 20 core holes and collected numerous outcrop samples as a part of its geotechnical evaluation of the El Llagal valley. The core holes were re-logged by PDDC geologists to identify samples with the most favourable mineralization. Select samples from the core and surface rocks were sent for gold and trace element analysis.


Project No.: 147107		Page 10-1

11.0		DRILLING

		Drilling campaigns have been conducted by most of the participating companies in the project over the years, including Rosario, GENEL, MIM, and PDDC.

11.1		All Drilling Campaigns

		The mineralized structures at the Pueblo Viejo project are both sub-horizontal and sub-vertical. Approximately 67% of the drill holes used in the PDDC resource estimate were drilled vertically (Table 11-1). PDDC has drilled almost exclusively angle holes and AMEC recommends that angled holes continue to be drilled in both deposits. The PDDC 2004 drill program included a number of infill drill holes that help support the resource estimation in areas where all the data for the grade model are interpolated using only vertical drill holes from previous drilling campaigns such as those executed by Rosario Dominica. Figures 11-1 and 11-2 show the locations of all historical drill holes in the Moore and Monte Negro deposits, respectively.

		Table 11-1: Drill Hole Dip Angles by Deposit


		Drill Hole	Moore Deposit						Monte Negro Deposit
Company	Drill Type	Series	No. Vertical Holes			No. Angled Holes			No. Vertical Holes			No. Angled Holes

Rosario	Rotary	R Pre-1975		57			—			—			—
Rosario	DDH	DDH < 159		31			20			—			—
Rosario	DDH	DDH > 159		40			—			43			—
Rosario	Rotary	RS		88			—			52			—
Rosario	RC	RC		20			—			44			—
Genel	DDH	GENEL					11						9
MIM	DDH	MIM		1			14			4			12
PDDC-2002	DDH	PD02		—			11			—			7
PDDC-2004	DDH	GT04		—			7			—			6

PDDC-2004	DDH	PD04		1			51			—			49

		Total		238			114			143			83
				(68	%)		(32	%)		(63	%)		(37	%)

11.2		Rosario Drilling

		Rosario employed several drilling methods during the exploration and development phases at Pueblo Viejo; these are summarized in Table 11-2. Geological information was recorded on paper log-forms or graphic logs for all core, RC, and percussion drill holes.


Project No.: 147107		Page 11-1

Figure 11-1: Location of All Historical Drill Holes – Moore Deposit


Project No.: 147107		Page 11-2

Figure 11-2: Location of All Historical Drill Holes – Monte Negro Deposit


Project No.: 147107		Page 11-3

Table 11-2:Rosario Drill Hole Summary

Drill Hole

Drilling

Geology

Total in

Included in

Moore

Monte Negro

Series

Period

Recorded

Database

Resource Estimation

Deposit

Backup Available for Review

Description

P Pre-1975

Pre-1975

Yes

343

Pre-production. Shallow percussion holes tested oxide mineralization.

R Pre-1975

Pre-1975

Yes

No drill logs. Copies of hand-written original lab results.

Pre-production. Shallow rotary holes tested oxide mineralization.

Late 1970s

Shallow percussion & rotary holes in exploration phase at Monte Negro tested oxide mineralization.

DDH < 159

Pre-1975

Yes

Hand-written original drill logs. Copies of hand-written original lab results.

Pre-production. Shallow NQ diamond drill holes in Moore. Some angle holes. Mainly tested oxide mineralization.

DDH > 159

1980-91

Yes

103

Yes

Hand-written original drill logs. Copies of hand-written original lab results.

Deeper NQ & PQ diamond drill holes. Vertical holes. Tested sulphide mineralization.

1978-90

Yes

176

Yes

Copies of hand-written original lab results.

Deeper rotary holes tested sulphide mineralization. Some holes removed from resource.

1984-85

Yes

Copies of hand-written original lab results.

Deeper reverse circulation holes tested sulphide mineralization.

1987-93

552

Shallow rotary holes tested transitional oxide/sulphide mineralization.

Early 1980s

252

Shallow rotary holes in Cumba deposit tested oxide mineralization.

Post 1990

Shallow rotary holes.

Post 1990

Shallow rotary holes in Moore deposit drilled to estimate recovery prior to excavating.

Shallow rotary exploration holes.

400

Shallow rotary holes in Arroyo Hondo tested oxide mineralization.

Total

2,166

256

139



Project No.: 147107		Page 11-4

		Detailed geology logs are available for only the “RS-series” rotary holes. Assay intervals and results were hand-written onto separate paper forms. The combination of drill hole name and “From-To” assay interval were used to identify the samples. Unique sample numbers were not used. No photos of the core were taken. This was common for the period in which the drilling was undertaken.

		The majority of the Rosario drill holes were vertical. Drill hole spacing ranges from 20 m to 80 m, with a nominal overall spacing of 50 m. No downhole surveys were performed. The type of instrumentation used for surveying collar locations is not documented.

		Core recoveries were reported to be approximately 50% in areas of mineralization and within silicified material. Poor drill core recovery in gold deposits is a common source of recovery-induced grade biases. Fluor evaluated this by calculating the average metal accumulation (grade x thickness) for core intervals that had been classed as having poor, fair, good, and excellent recoveries. Fluor reported that the grade for Au, Ag, Zn, and S samples in poor to fair recovery categories were elevated compared to the samples in the good to excellent classes. These findings suggest that the grade of samples may be affected by the recoveries.

		Fluor completed a second study where the data were composited to 10 m intervals and plotted on scatterplots against the recorded recovery values. The following observations were made:

	•		Gold grades varied with different recovery classes. In areas of 80% to 100% recovery, the gold values decreased with decreasing core recovery values. When recoveries were between 60% and 80%, the gold grade increased with decreasing recovery values. For recoveries less than 60%, the gold grades were generally low.

	•		Silver values were not affected by the recovery.

	•		Zinc grades exceeding 1.5% in value decreased in grade with decreasing core recovery values. Zinc grades below 1.5% appeared to be unaffected by core recovery.

		Fluor concluded that the poor core recovery did affect gold grades in samples, but in both positive and negative ways. Fluor also concluded that in the context of the whole deposit, “statistical noise” was apparent, but the data were not biased.

		AMEC is of the opinion that compositing on large intervals may mask the original relationship between recovery and grade and that the intervals used should be the as drilled intervals or regularized to 1 m intervals. It is AMEC’s opinion that there may be remaining concerns in the recovery versus grade relationships and these should be addressed in any future resource estimate update by evaluating gold recovery versus grade with both composited on short intervals.


Project No.: 147107		Page 11-1

		Fluor’s report discusses a potential sampling problem with regard to the rotary and reverse circulation drill holes. During a site visit to review the sampling, it was observed that the sample barrels used to catch samples from these holes were allowed to overflow when high water flows were encountered. This is a concern because the grade of these intervals can become elevated due to the loss of non-mineralized fine clay material. Fluor reviewed the problem in four separate ways and concluded that, with the exception of the P-series RC holes and the RS series of holes below the 250 m elevation in the West Flank of the Moore deposit, there was no systematic high bias in RC gold values versus core gold values. Fluor did determine that zinc was subject to placering in overflowing RC sampling devices and zinc values are biased low in RC holes. Fluor used RS series holes for gold estimates.

		The potential for high bias in some RS series holes, as revealed in portions of Fluor’s work, argues for review of each RC hole for inclusion in the resource estimation database. Biases may exist only locally and therefore not be apparent in global comparisons between RC and core holes. PDDC staff indicate that they are indeed performing those inspections and that holes RS40, RS25, RS60, RS60A, and RS69 have been excluded from the resource estimate because of possible problems with those data.

11.3		GENEL Drilling

		In 1996, GENEL drilled 20 holes at the Pueblo Viejo project site, 11 in the Moore deposit and 9 in the Monte Negro deposit (Table 11-3, Figures 11-1 and 11-2). Swiss-Boring was contracted to do the drilling using HQ core size. All holes were drilled at an angle. Downhole surveys were performed, but there is no record of the type of instruments used for the surveys. GENEL used a GPS system to locate drill holes and to survey the existing pits.

		AMEC was able to verify 5% of the assay data from these holes and found no errors in the database.

		Table 11-3: GENEL Drill Hole Summary

Drill Hole

Total in

Included in

Moore

Monte Negro

Series

Database

Resource

Deposit

MDD

11
(2,098 m)

—

MND

9
(1,053 m)


Project No.: 147107		Page 11-2

11.4		MIM Drilling

		In late 1996 and into 1997, MIM drilled 31 holes at the Pueblo Viejo project site, 15 at the Moore deposit and 16 at the Monte Negro deposit (Table 11-4, Figures 11-1 and 11-2). Geocivil was contracted to do the drilling. Core size was HQ with occasional reductions to NQ as necessary to complete the holes. Five holes were vertical and 26 were drilled at an angle. There were apparently no downhole surveys performed on these holes. There is no record of instrumentation used to survey collar locations.

		Original data documentation is not available from this drilling campaign for database confirmation, thus, which lab analyzed the samples or the methodology used cannot be confirmed. Source certificates for confirmation of the database results are not available. Drill logs were entered into Excel, and assays are presented as printouts.

		PDDC personnel found some of the core to relog but could not take advantage of it for re-assaying. The core was described as being in very poor condition, with serious degradation of the core boxes, especially the row dividers in each box, thus it is difficult to resample the core. AMEC recommends that PDDC re-log this core to confirm the logging in the database and send 5% of the samples for re-analysis to confirm the assay results.

		Table 11-4: MIM Drill Hole Summary

Drill Hole

Total in

Included in

Moore

Monte Negro

Series

Database

Resource

Deposit

Backup Available

MDD

15
(2,535 m)

Print-out of Excel database for drill log and assay sheets

MND

16
(2,065 m)

Print-out of Excel database drill log and assay sheets

11.4.1		Pre-Placer Dome Drill Hole Surveying

		Surveying methods for GENEL holes are of an accuracy suitable to support resource estimates. The accuracy of collar and downhole surveys for Rosario and MIM drill holes cannot be confirmed, but is assumed to be accurate enough to support resource estimates

11.5

PDDC Drilling

11.5.1		Summary

		PDDC completed 3,039 m of core drilling in 18 holes during 2002 and 15,424 m of core drilling in 111 holes during 2004. The drilling was a part of the resource confirmation program conducted by the PDDC geological team. The following criteria were set for selecting drill targets:


Project No.: 147107		Page 11-3

	•		Must be within the first three pit-stages from the May 2001 Pit Optimization Study (stage 1 = Monte Negro, stages 2&3 = Moore)

	•		In areas of higher-grade mineralization as defined by the December 2001 block model

	•		In areas with a limited number of core holes

	•		In high-grade mineralization defined by rotary or RC drill holes

	•		In high-grade mineralization defined by a limited amount of drilling.

		The drilling was completed using thin-walled NQ rods that produce NTW (57 mm) core. All but one of the core holes drilled by PDDC were angle holes, allowing for a better representation of the vertical sulphide veining.

		PDDC drilled with oriented core in order to calculate the true orientations of bedding, veining, and faulting in the deposit areas. A down-hole crayon-marking system was used to mark the bottom of the core. The system proved to be successful, and down-hole structures showed similar orientations to those measured at surface. PDDC was the only company to attempt oriented core at Pueblo Viejo.

		All the core was sawn in half, with half sent for analysis and the other half kept as a record at the project site. Samples were normally collected at 2 m intervals. Sample intervals were shortened at lithological, structural, or alteration contacts.

		Figures 11-3 and 11-4 are plan maps showing all the core holes in the Moore and Monte Negro deposits.

11.5.2		Surveying

		Drill pads were marked with wooden pegs after using a GPS to find the pre-selected locations. In areas where the GPS signal was weak, the Rosario bench map and IKONOS satellite images were used. Holes were aligned using foresite and backsite pegs.

		Two or three down-hole surveys were completed in all drill holes using a Sperry-Sun single-shot survey camera. Surveys were spaced every 60 m to 75 m, and deviation of the drill holes was minimal. Azimuth readings were corrected to true north by subtracting 10°. After completion, a wooden post marked with the drill hole number was placed in the collar of every hole. Final drill hole locations were then surveyed in UTM coordinates by a professional surveyor, translated into the mine coordinate system (truncated UTM), and entered into the drill hole database.

		AMEC personnel did not observe any of the drill hole markers because non-company personnel had removed them from the site, and no permanent (cement) monuments mark the locations of the drill holes. AMEC recommends that PDDC erect permanent drill hole markers so that independent verification of the collar locations is possible.


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Figure 11-3: Core Holes in Moore Deposit


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Figure 11-4: Monte Negro Drill Hole Location Map


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Surveying methods for Placer Dome holes are of an accuracy suitable to support resource estimates.

11.5.3		Geological Logging

		All drill holes were logged on paper forms using codes, graphic logs, and geologists’ remarks to capture data. Geological information related to assay intervals was recorded on a geology log. A second structural log was used to record individual structure information. Coded data and remarks were keypunched into Excel spreadsheets and edited on site by geotechnicians. Coded data were later imported into Gemcom to generate sections for resource modelling.

		The following geological data were recorded on the geology log with reference to assay intervals:

	•		Lithology – type, interval in metres

	•		Assay – interval, sample number (interval normally 2 m, but intervals were also cut at lithology changes or major structures)

	•		Oxidation – oxide, transitional, or sulphide facies

	•		Alteration – type, intensity

	•		Veining – type, estimated percentage

	•		Disseminated Sulphides – type, percentage.

		*Oxidation*

		The drilling was done in the pit area where most of the oxidized material had previously been excavated. Only a few of the drill holes intersected significant surface oxidation. Oxidized material intersected at depth was often related to groundwater movement along structural zones.

		*Alteration*

		Pervasive alteration was recorded as propylitic, pyrophyllite, quartz-pyrophyllite, or silicified. Intensity of alteration was recorded on a 3-point scale (1 = weak, 2 = moderate, 3 = strong). Where an assay interval contained more than one alteration type, the dominant alteration was recorded.

		*Veining*

		Visual estimates of quartz, pyrite, sphalerite, enargite, calcite, and pyrophyllite contained in veins within a sample interval were recorded.


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		*Disseminated Sulphide*

		Disseminated Sulphide was estimated as a visual estimate of the percent pyrite in a sample interval that is not contained within a vein.

		*Structural Logging*

		On the structural log, the geologist recorded the following information:

	•		Oriented Interval – core interval oriented by crayon mark

	•		Structure Interval – down-hole depth of structure

	•		Structure Description – type, true thickness (mm), oxidized (y/n)

	•		Structure Angle – alpha angle to core axis (0-90°), beta angle from bottom of the core to the down-hole apex of the structure (0-360°)

	•		Vein Composition / Dominance – minerals in vein listed in order of abundance.

		*Geotechnical Logging*

		Most of the geotechnical information was collected and recorded by technicians under the supervision of a geologist. The data were recorded on custom-made forms and later entered into Excel spreadsheets.

		The following geotechnical data were recorded with reference to drill intervals:

	•		Drill Interval – from-to and length in metres of block to block intervals; 1.5 m under normal drilling conditions

	•		Core Recovery

	•		Sum of Core Pieces > 10cm – RQD calculated by Excel

	•		Fracture Count – number of natural fractures per interval

	•		Oriented – whether or not drill interval was successfully marked with orienting crayon.

		Prior to making geotechnical measurements, the entire core interval was removed from the core box and placed in a long trough made of angle-iron. The fractures in the core were lined up, and man-made fractures were identified. This process allowed the technician to mark the orienting line on the core for a better estimate of core recovery and RQD.

		*Core Recovery*

		Recovery was measured in metres from core block to core block. In ideal drilling conditions this was 1.5 m, the approximate the length of the core barrel. Recovery percentage was calculated by the following formula:

		Recovery % = [Measured Core (m) /Core Block Interval (m)] * 100


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		*Rock Quality Designation (RQD)*

		RQD, a quantitative measure of fracture intensity, was calculated by summing the length of all core pieces more than 10 cm long within a core block interval. The following formula was used:

		RQD = [Sum of Core > 10 cm (in cm) / Core Block Interval (m)]

		*Oriented Core*

		Drill core was oriented using a down-hole crayon-marking system. The marking apparatus was dropped down the drill hole after every second core barrel was pulled (essentially every 3 m). Geotechnicians in the core logging area evaluated each mark and disregarded ones that did not appropriately mark the bottom of the core. They then marked the core bottom by drawing a single line down the length of the core.

		Data were collected by geologists to determine the true orientation of structures such as bedding, veining, and faults. Structural data were collected by measuring two angles on the core, the angle to core axis (alpha) and the angle to the down-hole apex (beta). True orientations were calculated in GeoCalculator V4.5.

		Geological and geotechnical logging by Placer Dome is reasonably comprehensive in regards to attributes that influence resource estimates and mine designs. The logged data is suitable to support resource estimates.

11.6

Discussion of Drilling Programs

In order to evaluate the possible biases between drill types and to validate the historical Rosario and MIM drilling information, PDDC and AMEC performed two tests. The first was to compare assays from PDDC and previous drilling programs. The second was a cross section review.

11.6.1		Rosario Pseudo-Twin Assay Pairing

		The pseudo-twin assay pair test compared results from nearby holes by searching for Rosario samples near the PDDC holes. Search radii were selected by PDDC to pair assays of different drilling campaigns to assess the similarity of assay grade distributions of the pairs. After examination PDDC concluded the QQ plots indicate that the Rosario and PDDC drilling campaigns reflect similar assay distributions.

		AMEC constructed declustered QQ plots and confirmed the conclusions made by PDDC. In grade ranges below 2 g/t, Rosario drilling appears to be biased slightly high compared to


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PDDI drilling; however, above 2 g/t and below 6 g/t the assays fit a line x=y reasonably well. Above 6 g/t Rosario drilling appears generally biased high relative to Placer Dome drilling.

11.6.2		Cross Section Review

		AMEC reviewed the assays on cross-sections and assessed the similarity of the MIM, GENEL, Rosario, and PDDC drilling. In general, there is close agreement of the orientation, tenor, and thickness of mineralization between drilling campaigns. PDDC drillholes with prefix PD and GT were drilled preferentially across mineralization. Six Rosario drillholes are crossed by seven PDI drillholes and indicate similar Au grade and thickness of mineralization. Collectively, the drillholes indicate a sharp, inclined top of mineralization. The Rosario drillhole indicates similar grade and mineralization thickness when compared to the MIM and PDI drillholes.

11.6.3		Conclusions

		Approximately 2.5% of the Rosario data have been verified against original documents. Extensive evaluations of the possible bias introduced by various drilling procedures have been undertaken by Fluor, PAH, PDDC and AMEC, and after reviewing the drill data, AMEC is of the opinion that the Rosario core, RC and some Rosario conventional rotary (pre-1975 and some Rosario RS-series) data are generally reliable. There may be some bias in the RC data, but those holes have been individually evaluated and obvious problems have been eliminated. The risk involved in using those data is judged by AMEC to be acceptable. Drilling types that have produced questionable results, such as the P-series percussion holes, ST-series rotary holes and select RC holes, have been excluded from the database and are not used in the resource estimate. AMEC recommends that the possible bias indicated between the RC, conventional rotary and core holes be investigated in the future and quantified if possible.

		GENEL data have been verified against original documents and are believed by AMEC to be reliable.

		MIM data have not been verified against original documents and there is some risk involved with using those data. AMEC compared those data to nearby PDDC data and found that the MIM holes indicated mineralized zones with very similar tenors and thicknesses as the PDDC and Rosario data. AMEC is thus of the opinion that the risk involved with using the MIM data is acceptable but recommends that every effort be made to obtain the original documentation in the future and verify those data.

		PDDC data have been verified against original documents and are believed by AMEC to be reliable.


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12.0

SAMPLING METHOD AND APPROACH

12.1		Pre-PDDC Drilling Programs

		No information is available concerning the sampling strategies used by Rosario during their drilling programs. The record suggests that Rosario generally sampled core on 2 m intervals with some samples based on lithology. RC holes were generally sampled on 2 m intervals. This approach is appropriate for the style of mineralization and mining methods that will probably be used.

		GENEL sampled on 2 m intervals. The core was split into thirds and one-third was used for the analytical sample. The remainder could be archived or split again for metallurgical test work. This approach is appropriate for the style of mineralization and mining methods that will probably be used.

		From the record, it appears that MIM samples were collected on 2 m intervals with adjustments for lithological boundaries. There is no documentation of the approach.

		Averaged sample intervals for the different drilling campaigns are summarized in Table 12-1.

		Table 12-1: Sample Interval Data for Rosario, GENEL and MIM Drill Holes


		Avg. Sample		Min Sample		Max Sample				Avg. Au
DDH		Interval		Interval		Interval		No. Samples		Grade
Series	Company	(m)		(m)		(m)		Taken		(g/t)
R	Rosario		2.18		0.20		4.60		1,489		2.49
RS	Rosario		1.99		1.00		6.00		9,959		1.79
RC	Rosario		2.00		1.00		2.00		5,003		1.77
DDH	Rosario		2.20		0.08		14.41		8,910		2.02
GEN	GENEL		2.00		1.40		2.30		520		2.51
MIM	MIM		1.97		0.20		8.00		2,309		2.21

		AMEC reviewed the possibility of a grade bias due to use of different sample interval lengths. The average gold grade for samples of less than 1 m, between 1 and 3 m, and larger than 3 m are tabulated in Table 12-2. From these generalized results it is apparent that grade may be biased by the sample length. PDDC removed the effect of length biases in its resource estimate by compositing assays on 2 m intervals


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Table 12-2: Sample Length and Grade Comparison Results


	Moore				Monte Negro
Interval Class	Avg. Au g/t		No. Samples		Avg. Au g/t		No. Samples
0 to 1 m		2.94		117		2.52		43
1 to 3 m		2.15		15,064		2.39		10,472
>3 m		2.07		84		2.75		5

12.2		PDDC Diamond Drilling

		AMEC did not observe core sampling procedures during the site visit and has relied on PDDC documents for this description (Keech and O’Brien, 2004). Sample intervals were normally 2 m long but were shortened at lithological, structural, or major alteration contacts. Prior to marking the sample intervals, geotechnicians photographed and geotechnically logged the core, and a geologist “quick-logged” the core, marking all the geological contacts. Geotechnicians then marked the sample intervals and assigned sample numbers. After the sample intervals were marked, the geologist logged the core in detail and the core was sent for sampling where it was split into two halves using a core saw.

		PDDC procedures are consistent with the type of mineralization and the type of mining envisioned for the project.


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13.0

SAMPLE PREPARATION, ANALYSES, AND SECURITY

13.1

Summary

		This section describes the sampling, sample preparation, and analytical procedures used by PDDC in the 2002 and 2004 drilling program. Quality assurance/quality control (QA/QC) protocols for all of the drilling programs (where known) are included in this section.

		AMEC finds the current QA/QC data pertaining to all the historical drill programs except the GENEL program, to be inadequate for proper validation of the assay results. The PDDC data set is acceptable but future drill programs could benefit from improved QA/QC protocols. Additional comments specific to each dataset are as follows:

	•		The number of check assays completed for the Rosario drill holes is limited but provides a level of confidence for specific drill holes. AMEC tabulated these check samples by drill hole and found that duplicate samples from 37 drill holes had been sent to a second lab. Because no standard reference materials (SRM) were submitted with either batch of samples, it is not possible to comment about possible bias in sample preparation or analytical procedures at either lab. The lack of quality assurance and quality control information, which is common for drilling programs up until the mid 1990s, requires alternative methods of validating the assays for use in resource estimates. AMEC and PDDC have largely validated these data by comparing the Rosario drilling results to results from a relatively large number of recent PDDC drill holes with QA/QC support. AMEC considers these data to be adequate for the current resource estimate, but recommends that PDDC continue in the future to drill confirmation holes in areas where they may suspect the results of individual Rosario holes.

	•		The GENEL drill holes were submitted to Chemex along with pulp duplicates and standard samples. The duplicate data indicate that 90% of the time, duplicate Au sample pairs have a relative difference of less than 14%, slightly above the target of 10%. This indicates that the sample preparation protocols were generally adequate. The standard samples were prepared by GENEL using reject material from metallurgical samples. The standard dataset includes many results that exceed the accepted limits, and it is not known if these samples were re-assayed. AMEC considers the GENEL QA-QC to be adequate for the current resource estimate.

	•		MIM samples were submitted for analysis without any known QA/QC protocols. This is a significant deficiency with regard to these data. PDDC’s and AMEC’s nearest pair analysis and comparisons of resource estimates derived from all data and only PDDC drill hole data suggest that the MIM holes are generally reliable. AMEC recommends


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			that PDDC drill confirmation holes in the future to check individual MIM holes where the results may be suspect.

	•		In 2002 PDDC submitted commercially purchased SRM samples (Au only) along with its samples to the primary lab, ALS. These SRM samples performed well, and only one batch had to be re-assayed. PDDC submitted 13% of the pulp samples to an umpire lab (Acme) but did not submit SRM samples with those samples. Results for Au, Cu, and Zn indicate no significant biases between the two laboratories, although Ag exhibits a possible bias of about 12% (Acme is higher than ALS). These results are appropriate for resource estimation purposes; however, AMEC recommends that PDDC improve its QA/QC protocols to include submission of blanks, pulp duplicates, coarse reject duplicates to the primary laboratory, and submission of pulps and SRMs to a second laboratory for check assays. In addition, assays for elements other than Au need to be supported by a QA/QC program if they are significant to the project.

	•		In 2004, PDDC submitted SRM samples that included Au, Ag, C, S, Cu, Fe, and Zn. AMEC reviewed those data and found them to have performed well.

13.2

Sample Preparation Procedures

13.2.1

Rosario

		The fire-assay and other analytical procedures in use up to the time of the Stone & Webster (1992) study were reported to be of industry standard. No details are available for AMEC to review regarding the sample preparation and analytical procedures used. It is only known that the samples were analyzed using the fire-assaying technique for gold and silver and, the Leco combustion furnace method for carbon and sulphur, and AA (atomic absorption) for copper and zinc. No details are available on crush sizes, subsample sizes, or final pulp sample weights used during sample preparation.

		For the sulphide drilling program, which started in 1984, two assay labs were present at site: a “mainline” lab responsible for Au, Ag, Cu, Zn, and Fe analyses and a “sulphide” lab responsible for carbon and sulphur analyses. Sample preparation methods are not documented for this period of sample analysis.

		Security of the samples after removal from the hole is not documented.

13.2.2		GENEL

		GENEL samples were apparently prepared on site by GENEL personnel. This is not explicitly stated in the GENEL report, but is inferred by AMEC from the discussion in the report. A one-third split of the core was crushed to -10 mesh, homogenized by passing the sample through a Gilson splitter 3 times. The sample was then “down-sized” to about


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		400 grams using a Gilson splitter. This sample was then packaged and sent to Chemex Labs Ltd. in Vancouver, BC, Canada (Chemex) where, AMEC assumes, that the final pulverization was completed. In the GENEL report, the final pulp grain size is not stated.

		The samples were then assayed at Chemex for Au, Ag, Zn, Cu, S, and C. The procedures are not stated in the report. A 32 element ICP analysis (G-32 ICP) was performed on each sample.

		Security measures utilized by GENEL are not documented.

13.2.3		MIM

		No details are available on the sample preparation, analytical procedures, or security measures for the MIM samples.

		Core from Rosario, MIM and GENEL drilling was previously stored in inadequate storage facilities that have led to severe oxidation of the remaining core, rendering it relatively useless.

13.2.4		PDDC

		During the 2002 and 2004 program, drill core was sawn in half with a diamond blade saw at site. All of the second half of 2002 core was consumed in metallurgical testwork. The archived half of 2004 core was stored on site for future reference in suitable storage conditions. The other half was placed in plastic sample bags marked with the appropriate sample number and sealed with a numbered security tag (‘zap-strap’). The manager of the drilling company drove the samples from the site to the airport unaccompanied by a PDDC employee. The core samples were sent to Vancouver using airfreight and were received by ALS. No record was kept of the state of the zip tags when logged into ALS. AMEC recommends that if any security tags or bags are found to have been opened, then PDDC should track this in the assay database. Customs services in the U.S. or Canada were reported to have occasionally opened a bag for inspection purposes.

		Initial splitting of core on site was the only aspect of sample preparation performed by employees of PDDC. The split samples were then sent to commercial laboratories for sample preparation and analysis.

		The following description of sample preparation and analytical procedures was based on PDDC and ALS documents (Keech and O’Brian, 2004). PDDC used Acme Analytical Laboratories Ltd. (Acme) in Vancouver, British Columbia, Canada, as a check assay laboratory. AMEC did not visit the laboratories or observe the procedures during the 2002 drilling program.


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		The samples received by ALS were prepared following ALS’s ‘Prep-31’ procedure. This includes marking all bags with a bar code, drying and weighing the sample, crushing the entire sample to greater than 70% passing 2 mm (10 mesh), and splitting off 250 grams. The split is pulverized to better than 85% passing 75 µm (200 mesh) and is used for analysis. The remaining sample material (‘reject sample’) is stored at WestCoast Mineral Storage in Aldergrove, BC, Canada.

		The sample preparation procedure used by PDDC is primarily designed for Au assays using screen fire assay procedures and not for conventional fire assay procedures. The first crush and pulverization passing limits are quite low compared to similar projects in the industry. Gold deposits, especially those with even small amounts of coarse gold, typically need deposit-specific sample preparation protocols involving much finer crushing and pulverization. Gold deportment studies by AMTEL do not suggest a common presence of coarse gold, therefore the first crush specifications may not be biasing the assays. This should be confirmed in the future and, as standard procedure, AMEC recommends a 90% pass rate for the first minus 10 mesh crush.

		The core samples were assayed for Au, Ag, Cu, Zn, C, S, and Fe. Table 13-1 shows the assay techniques used by ALS. In addition to these elements, multi-element analysis was performed on drill hole PD02-003 (80 samples) using ALS’s ME-MS61 procedure. In 2004, every other sample from all drill holes was also analyzed using ALS’s ME-MS41 procedure.

		All drill core samples from the PDDC drilling campaign were analyzed for total carbon by ALS’s C-IR07 LECO furnace procedure. To ensure that the total carbon values represented organic carbon, a suite of 114 samples were re-analyzed by the C-IR6 procedure, which removes all inorganic carbonate by leaching the sample prior to LECO analysis. The sample suite represented all of the lithologies found in the deposit area; all exhibited advanced argillic alteration or silicification with varying intensities. The results showed that the total carbon analysis is representative of organic carbon in samples with advanced argillic alteration or silicification. A similar study should be completed with samples having propylitic alteration. There is an indication, especially from the Rosario drilling programs, that total carbon assay values represent both organic carbon and inorganic carbonate in propylitic altered zones.


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Table 13-1: Summary of Pueblo Viejo Assaying Procedures


	ALS-Chemex
Element	Method Code	Description	Range
Au	Au-GRA21	30 g fire-assay, gravimetric finish	0.05-1,000 ppm
Ag	Ag-GRA21	30 g fire-assay, gravimetric finish	5-3,500 ppm
Cu	AA46	Ore grade assay, aqua regia digestion, AA finish	0.01-30%
Zn	AA46	Ore grade assay, aqua regia digestion, AA finish	0.01-30%
C	C-IR07	Total Carbon, LECO furnace	0.01-50%
S	S-IR07	Total Sulphur, LECO furnace	0.01-50%
Fe	AA46	Ore grade assay, aqua regia digestion, AA finish	0.01-30%

13.3

QA/QC Procedures

13.3.1		PDDC 2002

		The PDDC QA/QC program consisted of submitting standards into the sample stream (Keech, 2004). Standards were purchased from CDN Resource Laboratories of Delta, B.C., Canada, and correspond to the approximate average gold grade and cutoff grade for the deposits. In 2002, the standards were alternatively inserted as every 20^th sample. PDDC did not routinely insert duplicate or blank samples for their 2002 drilling program. The QA/QC program was modified for the 2004 drilling such that a standard and blank were submitted with every batch of 20 samples (10% of the samples were control samples). AMEC recommends that for future sampling programs, PDDC should make a series of custom standards using material from the deposits. These SRMs need to be certified for all elements in the typical analysis, such as Au, Ag, Zn, Cu, Pb, and S. If the results for these elements are to be included in resource or reserve statements, then the quality of the analysis results need to be certified through SRMs.

		Plots of gold vs. batch number show that the majority of the standards returned values within 2 standard deviations of their established mean. Only SRM GS-2 returned a gold value outside of this range. To confirm the gold grade, the 21 samples associated with this standard were re-assayed. AMEC reviewed the pass-fail limits PDDC used for the two standards and is of the opinion that those limits are somewhat liberal and should be revisited and reduced.

		AMEC reviewed the 21 re-assayed duplicate pulp samples from the 2002 drilling campaign. The samples had been re-assayed because of the failure of a single standard. The set of duplicate data is very small and may not be representative of the whole, but no other duplicate data were available. The results indicate that the duplicates correspond reasonably well.


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		AMEC uses the relative error which is the difference of the samples divided by the average of the samples as an indication of precision. When the data are ordered and the cumulative frequency calculated, AMEC uses the relative error at the 90^th percentile as a standardized measure of the precision. In this case, the relative error is about 23% at the 90^th percentile, which is somewhat higher than the level of 15% anticipated from this type of deposit. The difference is probably due, in large part, to the sample preparation procedure. The relative error would probably decrease if the procedure were modified to produce 95% passing 2 mm and 95% passing 75 µm. As indicated above, these conclusions are based on a very small data set that is assumed to be representative of the whole, which may not be the case. Additional pulp duplicate data are required to produce an adequate estimate of precision, but these data indicate that the efficacy of the sample preparation procedure needs to be investigated and changed if it is found to be inadequate.

		In 2004, SRM samples GS-2, GS-4, and GS-9 were routinely inserted into the samples stream along with a blank limestone. All of these standards and the blank were assayed for Au, Ag, C, S, Cu, Fe, and Zn which provides a basis to evaluate the performance of those elements. AMEC calculated best values for all of the elements in each sample, based on the results from ALS-Chemex. Au is the only certified value and the best values for Au calculated from the ALS-Chemex data were indistinguishable from the certified values which indicates that ALS-Chemex generally performed well. One Au result for GS-2 fell outside the pass-fail envelope (25 analyses). All of the other elements fell within the expected range. GS-4 (174 analyses) showed three to six failures for most elements. GS-9 (120 analyses) performed well with only one or two failures per element for the 2004 data. One sample varied significantly for C, S, Cu, Fe, and Zn. The results for those elements suggest that SRM GS-4 was analyzed. Au and Ag, however, are consistent with GS-9. The source of this problem is uncertain. The blank (380 analyses) is a limestone. The data generally show blank values except for C which runs about 11% in the sample. Approximately 10 of these samples appear to have been miss-identified and were actually other SRMs. Overall, the standards indicate that laboratory performance in 2002 was adequate. AMEC has not verified that batches containing SRMs that failed were reassayed. AMEC has not reviewed any duplicate and check assay data for the 2004 drilling program.

13.3.2		ALS Chemex Quality Control

		ALS conducts analytical quality control in its laboratory by inserting blanks, standards, and duplicates into every sample run. Table 13-2 shows the frequency of insertion of quality control samples.


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Table 13-2: ALS Quality Control Sample Frequency

The laboratory staff analyzed quality control samples at the frequency specified in Table 13-2 as a minimum and, if necessary, may include additional quality control samples. All data gathered for quality control samples (blanks, duplicates, standards) are automatically captured, sorted, and retained in the QC database. PDDC should review ALS’s internal QA/QC procedures in the future as a second check on precision and accuracy and the effectiveness of the entire sample preparation protocol.

13.3.3		Acme Check Assay Program

		As part of the PDDC QA/QC program, sample pulps were sent to Acme Laboratories in Vancouver to be analyzed for Au, Ag, Cu, and Zn. Approximately 10 sample pulps were shipped from every drill hole, resulting in 187 samples, or 13% of the total samples from the 2002 drill program. An additional 247 sample pulps were shipped during the 2004 drilling program that were analyzed for Au only. AMEC did not review these data. Standard samples were not inserted into the check pulp shipments.

		Results for Au, Cu, and Zn indicate no significant biases between the two laboratories. There are one or two outliers in the Au data, but AMEC does not consider them to be important. Ag, on the other hand, exhibits a possibly significant bias of about 12% (Acme higher than ALS). The source of that bias is not immediately apparent, but if may be the result of differences in the analytical protocols. Without information from SRMs, the source of the bias cannot be determined.

13.3.4		GENEL QA/QC – 1996 to 1997

		GENEL used a combination of duplicate and standard samples to monitor the quality of its assays (Lockhart and Bowen, 1997) and restricted its QA/QC to Au. For the 1,565 assay results for the GENEL series of drill holes, 171 duplicate samples were submitted for analysis, representing 11% of the samples. Duplicate samples compare well, and the relative error of the sample pairs at the 90^th percentile is about 14%. This is within the expected range for this type of deposit and suggests that the analytical protocol was adequate. AMEC is of the opinion that the quality of the GENEL is adequate for incorporation in the resource estimate.


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		Results of GENEL standards STD A, STD B, and STD C show generally that assays of these standards were within acceptable limits. It is not clear if sample batches with failures were re-assayed. GENEL assay certificates should be used in the future to check that failed batches were actually re-assayed.

		The GENEL QA-QC results show that the quality of the GENEL data is adequate for inclusion in the resource estimate.

13.3.5

1985 Rosario Check Assays

		According to Appendix A of the Fluor Feasibility Study, Rosario sent samples to three labs in 1985 to validate assays for Au, Ag, C, and S. Rosario sent 392 samples to the Colorado School of Mines Research Institute for check assaying of the Au and Ag values in three batches of 188, 176, and 28 samples. Another 236 samples were sent to Hazen Labs and 154 to AMAX Research and Development Lab for sulphur and carbon analysis. Excel spreadsheets of the Au and Ag results were made available for review by AMEC. Results for the AMAX C and the S checks were not available digitally and no assay certificates were available to validate the data. Information concerning the sample sizes or the analytical methods used is also not available.

		The data file that AMEC received from PDDC included 216 results for Au and Ag from the Colorado School of Mines Research Institute. This represents results for the 188 and 28 sample shipments. The results for the other 176 samples were found and added to the database at a later date but were not included in AMEC’s study.

		For the data received by AMEC, Au results generally correspond well, but there are a number of outliers that are possibly sample swaps. There is a small, but acceptable, bias between the two laboratories. The Ag results generally agree well, but as with Au, there are number of unexplained outliers, some of which are probably also sample swaps. The bias between the laboratories is about 7%, which is slightly outside generally acceptable limits (5% bias is a general limit within the industry). The bias may be the result of differences between the analytical procedures, but it is not possible to accurately determine the cause at this time.

13.3.6

1978 Rosario Check Assays

Fluor reported that Rosario sent 1,586 samples to Union Assay Laboratory in Salt Lake City, Utah, U.S., for check assays in 1978. However, the Excel file supplied to AMEC by PDDC contained only 500 sample results. PDDC obtained these pairs of assay results from the handwritten assay sheets included with the paper copies of drill logs supplied by Rosario. After AMEC’s review, the missing check samples were located by PDDC and added to the database.


Project No. 147107		Page 13-8

		The samples were taken from 10 drill holes (DDH161, 163, 167, 169, 170, 171, 173, 175, 180, 185). The gold check assays exhibit substantial scatter including several obvious outliers. Some of the scatter may be due to sample swaps, but most of it is unexplained. There is a small bias in the data, just outside the general acceptance limit of 5%. AMEC does not consider this bias to be significant. Below 4 g/t, the data exhibit significant scatter that scatter is largely due to the analytical and reporting procedures used for the assays. Overall, excluding obvious outliers, the data correspond reasonably well. The Ag data are similar to the Au data in the significant amount of scatter about the X = Y line and the large number of outliers. There is a small (5%) bias between the laboratories, but this is probably not significant. Cu exhibits a small amount of scatter about the X = Y line and no appreciable bias between the laboratories. Some of the outlier samples may be sample swaps, but there is no way to confirm this at present. Zn exhibits more scatter than Cu, but less than Au and Ag. Some of the outliers appear to be sample swaps. There is about a 7% bias between the laboratories. Although the bias is outside the 5% limit generally accepted by the industry, AMEC is of the opinion that the bias is not significant to the resource estimate.

		Check assays for the 1978 and 1985 Rosario holes and Rosario holes of a similar age show generally that results are acceptable for resource estimation purposes. PDDC’s and AMEC’s validation of Rosario holes using nearest-pair comparisons to PDDC holes and comparison of resource estimates with and without PDDC holes generally confirm this conclusion..

13.4 Discussion of Sampling, Sample Preparation and QA/QC at Pueblo Viejo

		Sampling at Pueblo Viejo by PDDC has been performed appropriately for the style of mineralization present. Sampling of the pre-PDDC samples may have been adequate, but there is little in the way of documentation to confirm this.

		Sample preparation for the Rosario and MIM samples has not been documented and AMEC cannot comment on the adequacy of sample preparation for those programs. Sample preparation for the GENEL samples was adequate as demonstrated by the QA-QC data for those samples.

		Preparation of samples from the PDDC drill programs used initial crush specifications less than generally used for gold deposits. The 90^th percentile relative error for 21 duplicate samples assayed is about 23%, which is higher than is generally expected (15%) for pulp duplicates. This is a very small data set with which to draw conclusions, but if it is representative of the whole, much of the relative error is probably related to sample preparation procedures. A similar review of duplicate samples analyzed by GENEL indicates that that program had a 90^th percentile relative error of about 14%, which is within


Project No. 147107		Page 13-9

		the expected range. AMEC recommends that in the future a higher pass specification in the initial crush, such as 95% passing 2 mm, be used.

		QA/QC procedures have varied significantly during the history of work at Pueblo Viejo. During the time that Rosario operated the project, QA/QC consisted of two batches of check assays sent to a second laboratory without duplicate, blank, or standard samples. Although the QA/QC was substandard relative to current industry practice, it must be viewed in its historical context, and check assaying was the industry standard for QA/QC at that time. AMEC is of the opinion that the data for the time periods covered by the check assay samples are of adequate quality to include in a resource estimate.

		AMEC is of the opinion that the GENEL QA/QC program was adequate and that the data generated by that program are of adequate quality to include in a resource estimate.

		MIM sample data lack any QA/QC validation. The quality of those data is indeterminate. There is no reason to believe that there are any problems with those data, but the quality cannot be directly evaluated. Comparison of the tenor and thickness of mineralized zones defined by the MIM data with tenor and thickness of mineralized zones defined by the PDDC and GENEL indicate that the grades are similar and AMEC is of the opinion that this is adequate validation of the MIM data.

		PDDC relied on two standards and check assaying for QA-QC and AMEC is of the opinion that the amounts and types of control samples is less than required for confirmation of precision, accuracy and adequate sample preparation. No duplicate samples were analyzed, and the check analysis program included no certified reference materials or blank samples. AMEC recommends that, in addition to check assaying, future QA/QC programs include field duplicate samples, crusher duplicate samples, pulp duplicate samples, blind duplicate samples, additional standard samples, and blank samples.


Project No. 147107		Page 13-10

14.0 DATA VERIFICATION

14.1 Summary

The history of the project database is well documented in the literature provided to AMEC for this study. This includes detailed descriptions of the data entry and validation procedures completed Fluor, AMS, and PDDC. AMEC is of the opinion that these historical databases are not related directly to the current one and that the historical protocols for data entry and verification, as described in these documents cannot be used to validate the current database. To directly validate the current database, original data sources such as laboratory certificates, original drill logs, and survey data for both collar and down-hole surveys must be used. PDDC lacks some of this backup for the current database, which is a common problem with projects with the longevity of Pueblo Viejo. Alternative, indirect methods of validating the data must be used (such as comparisons with more recent drilling) where original documents are missing, or assumptions of data integrity must be made.

	•		For the Rosario drill holes, original drill logs and hand-written assay sheets are available for validation. The assay certificates do not state that they are directly from the lab, but it is a reasonable assumption that they are. No survey notes are available to validate the collar data. PDDC used the same surveyor employed by Rosario to complete collar surveys for the 2002 drill holes. PDDC found that of the 20 holes surveyed, the collars for about three were incorrect, with one hole being 1 km from its actual location. PDDC contracted a new surveyor to resurvey the holes. If Rosario was not as attentive as PDDC, then there is a likelihood of errors in the locations of the Rosario holes.

	•		For the GENEL drill holes, PDDC has copies of the original assay certificates which were used to validate the assay database and copies of drill logs printed from an Excel database. These were entered from the original logs, which have since been lost. Because these data are not original, they can be used to validate the geological database only if it is assumed that GENEL’s data entry was error free. Survey notes are not available to validate the collar data.

	•		For the MIM drill holes, original drill logs or assay certificates are not available for validation. PDDC’s drill hole data files are printouts from an Excel database.

	•		For the PDDC drill holes, both the original drill logs and the laboratory certificates are available to validate the geology and analytical results in the database. The assay data were checked and found to be of excellent quality. A small number of errors were discovered in the geological database. Those errors will have no significant impact on the resource estimate.


Project No. 147107		Page 14-11

14.2 Verification of Pre-PDDC Data

14.2.1

Database Development

		American Mine Services (AMS), as part of the 1992 Stone & Webster “Feasibility Study on the Sulphide Resources at Pueblo Viejo,” developed a computer database consisting of drill hole collar locations, assays and assay intervals, and geological data. The AMS database formed the foundation of the database provided to GENEL and MIM in 1995. In 2001, Rosario supplied PDDC with a computer drill hole database in Gemcom format. PDDC also purchased GENEL data files. PDDC personnel could not confirm that the database they received from GENEL is indeed the same as the one AMS built and validated.

		PDDC undertook a study to compare the GENEL drill hole database with that provided by Rosario and confirm that no changes had been made since AMS’s validation efforts. The study found that only minor changes had been made; these were examined and corrected based on original Rosario assay sheets and drill logs at the Pueblo Viejo site. PDDC compared drill locations and assay grades to original paper plans and sections at the mine site. Drill hole collar maps were plotted using the computer database and compared against hand-drawn maps and typewritten drill hole collar reports. A complete description of the validation work is contained in the PDDC report, “Report on the Comparison of the PDI02 and GENEL98 Drill Hole Databases for the Pueblo Viejo Project, Dominican Republic, 11 February 2003.”

14.2.2

Comparison of PDDC, GENEL, and Rosario Drill Hole Assays

		PDDC verified the assay grades of Rosario holes using the results from the independent confirmation drilling completed by PDDC in 2002 and 2004, and also using earlier drilling by GENEL. QA/QC data show the PDDC and GENEL holes to be generally reliable and thus suitable for this comparison.

		PDDC paired assays from PDDC and GENEL drilling with assays from Rosario drilling using different search radii and then prepared QQ plots to show assay grade distributions for each. A nearest-pair analysis was necessary because most Rosario holes are vertical and nearly all GENEL, MIM and PDDC holes are inclined. AMEC constructed declustered QQ plots and confirmed conclusions by PDDC. In grade ranges below 2 g/t, Rosario drilling appears to be slightly biased high compared to PDDC drilling; however, above 2 g/t and below 6 g/t the assays fit a line x=y reasonably well (Figure 14-1). Above 6 g/t Rosario drilling appears generally biased high relative to PDDC drilling.

		This work generally shows the Rosario drilling to be reliable, but represents a risk to the project where projected high-grade material may be lower than anticipated.


Project No. 147107		Page 14-12

Figure 14-1: AMEC Comparison of PDDC and Rosario Drill Hole Assays with 10 m Search Distance

14.2.3

Cross Sectional Review of MIM, Rosario, and PDDC Drilling

AMEC reviewed the assays from cross-sections on the computer screen and assessed the similarity of the MIM, GENEL, Rosario, and PDDC drilling in both the Moore and Monte Negro deposits. In general, there is close agreement of the orientation, tenor, and thickness of mineralization between drilling campaigns in both deposits where MIM, GENEL, Rosario and PDDC drill holes cross.

14.3 Verification of PDDC Data

AMEC compared one in twenty samples in the PDDC part of the assay database with original assay certificates and found no errors. Approximately 5% of the assay values in the database were checked against original assay certificates.

14.4 Historical Twinned-Hole Comparisons

Fluor Metals and Mining Ltd. (Fluor) undertook a study of drill holes twinned by Rosario as part of its 1986 feasibility study. Fluor compared closely spaced drill holes based on a metal accumulation approach (grade x thickness). The results showed some differences


Project No. 147107		Page 14-13

		between some of the drill holes. Hole RS-40 was removed from the study because it is believed to have been drilled down a near-vertical vein. Fluor’s comments about the twin holes results include: “Analysis of Au, Ag, Zn, C, and S assays then showed no significant overall bias except for carbon and zinc. Carbon assays were consistently lower by 7% and zinc assays lower on average by 36% than in the original hole.”

		PDDC reviewed 20 twinned and closely spaced drill holes and compared the gold grades using a profile plot and a scatterplot with summary statistics. The results of this analysis show good agreement between different drilling methods—rotary, reverse circulation, and core—when the holes are closely spaced. However, some R and RS holes are likely to have been contaminated by the drilling process or to have been drilled down a mineralized structure. As indicated earlier in this report, PDDC decided to remove drill hole RS40 from the drill hole database for resource estimation.

		AMEC reviewed the database and identified 40 holes that were twinned in part or in whole by Rosario, GENEL and MIM; these are listed in Table 14-2. None of the 2002 or 2004 PDDC holes were twins of previously drilled holes. The conclusions herein are predicated on the assumption that the collar locations and down-hole surveys for all holes are accurate. This review showed that some twin holes compare well and some do not.

		In the case of twin holes DDH161 and RS2 (core vs. RC comparison), the two holes are separated by a constant 7 m and compare very well in that the general grade trends are distinguishable in both holes, and the total metal content is similar in both. Twin holes RS94 and RC14 (RC vs. RC comparison) are separated by about 10 m throughout and show extremely good correlations of both mineralized zones and total metal.

		On the other hand, twin holes GEN_MNDD1 and GEN_MNDD1A (core vs. core comparison) show large differences in average grade (4.89 g/t and 5.82 g/t Au, respectively) and in total metal content. The two holes, even though they are at all points separated by less than 1 m, show high-grade zones that do not correlate.

		There is no clear relationship between core and RC drilling; however, there seems to be a tendency for RC holes to have higher grades than core holes. Twin holes DDH218 and RC3 (core vs. RC comparison) show that the RC drilling has substantially higher grades and more contained metal in holes separated by about 3 m throughout.

		Twin holes DDH219 and RS108 (core vs. RC comparison) are separated by about 1.5 m and show core with somewhat higher grades (1.78 g/t, vs. 1.59 g/t, respectively) and overall higher metal contents.

		Where core grades and metal contents are higher than the corresponding RC hole, the difference is typically much less than the difference between RC and core where RC grades are higher.


Project No. 147107		Page 14-14

Table 14-1: List of Twin Holes


Hole-ID	Easting	Northing	Elevation	Length	Hole-ID	Easting	Northing	Elevation	Length

AH367	73626.00	96332.70	472.40	34	AH369	73626.10	96332.80	472.40	26
AH533	75615.30	95618.90	311.70	16	DDH233	75615.32	95618.90	311.73	128.6
DDH131	75756	94552	214.5	38.2	RS83	75751.2	94546.53	215.65	174
DDH161	76159.75	94600.34	348.61	232.5	RS2	76165.87	94596.29	348.84	152
DDH162	76000.00	95302.20	338.39	242	RS11	76000.14	95301.24	329.35	130
DDH218	74992.03	95750.3	381.28	108	RC3	74992.97	95747.84	381.18	210
DDH219	74881.52	95808.29	363.06	116.1	RS108	74882.88	95808.97	362.81	184
DDH258	74312.19	95613.00	303.35	89.6	AH395	74314.00	95608.40	304.20	40
DDH259	75836.23	95092.43	310.96	187.85	RC47	75834.01	95097.23	311.25	112
GEN_MDD2	75871.76	94400.91	219.52	132.2	GEN_MDD2A	75871.76	94400.91	219.52	40
GEN_MNDD1	75210.71	95175.38	262.67	27.4	GEN_MNDD1A	75209.46	95175.25	262.41	74.1
GEN_MNDD4	75151.64	95600.18	318.77	140.2	GEN_MNDD4A	75151.64	95600.18	318.77	18.3
GT04-10	76251.35	94657.25	344.92	126.49	RC21	76251.48	94647.66	347.44	177
MIM_MN007	75175.61	95713.38	351.77	50.3	MIM_MN008	75175.61	95713.38	351.77	122
MIM_MO007	76006.03	94476.26	245.14	200.15	MIM_MO009	76005	94476.3	245.1	200
MIM_MO015	75903.42	94702.51	258.41	150	RC23	75899.97	94700.09	261.05	178
R117	76674.50	94570.10	328.70	18	R117B	76675.72	94571.16	329.03	44
R17	75992	94298	243.5	98.3	RS135	75997.24	94302.59	247.42	200
R29	75860.00	94455.60	226.60	18.3	R30	75860.90	94451.10	226.60	59.4
R42	76233	95003	332.8	77.7	RS90	76232.8	95005.17	329.34	194
R60	75856.00	94782.00	258.00	75.4	ST562	75856.42	94787.88	258.30	60
R70	75852	94832	264.6	91.4	RC51	75845.41	94837.84	265.8	164
RC14	75043.02	95746.28	380.13	204	RS94	75051.03	95752.06	379.84	237.58
RC15	75115.12	95394.38	309.67	114	RC9	75115.18	95397.14	309.79	208
RC16	75985.15	94802.1	281.68	84	RC18	75981.53	94808.59	281.36	146
RC16	75985.15	94802.1	281.68	84	RS40	75984.9	94800.8	288.9	180
RC9	75115.18	95397.14	309.79	208	RC15	75115.12	95394.38	309.67	114
RS111	74904.55	95592.19	344.56	56	RS111A	74901.39	95583.09	343.61	140
RS131	75290.86	95100.48	251.04	116	ST455	75291.20	95104.90	250.00	60
RS142	75195.50	95362.30	291.90	150	ST183	75196.20	95370.70	291.90	50
RS2	76165.87	94596.29	348.84	152	RS3	76169.54	94604.54	349.04	202
RS3	76169.54	94604.54	349.04	202	RS4	76175.83	94609.45	349.41	72
RS4	76175.83	94609.45	349.41	72	RS5	76184.67	94605.36	349.38	142
RS40	75984.90	94800.80	288.90	180	RC16	75985.15	94802.10	281.68	84
ST257	75947.72	94360.58	220.18	30	ST236	75947.78	94361.00	220.00	30
ST329	76252.18	94759.21	337.08	30	ST630	76252.18	94759.21	337.08	60
ST445	75073.19	95630.21	341.21	10	ST445A	75073.19	95630.22	341.21	50
ST543	74882.45	95652.65	344.62	12	ST543A	74882.45	95652.65	344.62	32
ST569	75633.84	95850.93	338.29	11	ST578	75633.84	95850.93	338.29	16
ST630	76252.18	94759.21	337.08	60	ST631	76252.18	94759.21	337.08	50


Project No. 147107		Page 14-15

AMEC concludes that the wide divergence in twin-hole behaviour allows no simple conclusions to be drawn from the data. There is a tendency for RC holes to return somewhat higher grades and metal contents than core holes; this may be due, in part, to localized down-hole contamination in the RC holes. There appear to be zones within the Pueblo Viejo deposits where the grades are extremely erratic, and holes separated by only a few metres return very different results. This probably explains many of the differences observed between twin holes.

14.5 Decay and Cyclicity (Down-hole Contamination) of RC Holes

14.5.1

Introduction

		AMEC investigated the possibility of down-hole contamination in the RC portion of the drilling at Pueblo Viejo. AMEC’s review focused on the two specific down-hole contamination problems that can occur in RC drilling: cyclicity and decay. Cyclicity is the tendency of metal to concentrate at the bottom of holes during pauses in drilling, which typically occurs when rods are changed but can happen at any time during the drilling process. Collapse of unstable zones in RC holes tends to occur when drilling is stopped. Decay is the tendency of material from soft, gold bearing zones to ravel down hole, contaminating samples from less mineralized material. This usually is expressed as a gradual diminishing of values downhole. This feature can also occur naturally due to halos of low-grade material around high-grade material. Typically, cyclicity and decay are linked. Gold grades can be enhanced by both factors.

		An idealized example of decay and cyclicity is presented in Figure 14-2. For this example, there are assumed to be three samples per rod and rod breaks are on intervals of 3 depth units. The grade spike at depth 6 is symmetrical and is followed by five numbered cycles. Each cycle has a grade spike at the rod break and is followed by monotonic decay prior to the next cycle. The five cycles also exhibit diminishing grade (decay) from depth 9 to depth 22 (red line connecting the grade spikes). The dashed magenta line shows the expected projection of the grade profile in the absence of contamination. To adjust the data, samples from depth 9 to depth 11 would be adjusted downward to a grade on the projected grade trend (or general background level if the uncontaminated grade spike is not symmetrical). Samples from depths 12 to 21 would be adjusted downward to the background value at 0.05 grade units. As can be seen in the figure, cyclicity and decay can cause significant increases in the grade and tonnage of a deposit and must be identified and grades adjusted, if possible.


Project No. 147107		Page 14-16

Figure 14-2: Idealized Example of Cyclicity and Decay

AMEC assumed that 6 m rods were used and that sample lengths are more-or-less uniform at 2 m. Table 14-2 lists the holes AMEC identified as having possible down-hole contamination problems. The RC and RS series holes exhibit some evidence of contamination based on this analysis, but the problem appears to be substantially less than in the ST series holes and generally acceptable given that the other Rosario RC holes will the most obvious problem (RS40, RS25, RS60, RS60A, and RS69) have been removed from the resource database. The ST series holes seem to be particularly prone to contamination. AMEC concurs with PDDC’s decision not to use the ST series holes for the 2003 or subsequent resource estimates.

14.5.2

Conclusions

		This review of decay and cyclicity as indicators of down-hole contamination in RC holes revealed that there are possible problems with a number of holes but, with the exception of the ST series holes, contamination is not believed to be a widespread problem. Table 14-2 lists the holes with anomalous behaviour in decay and cyclicity profiles. Each of these intervals should be reviewed in detail before use in future resource estimate updates. ST series holes appear to be particularly susceptible to down-hole contamination. AMEC concurs with PDDC’s decision not to use the ST series holes in the 2005 and subsequent resource estimates.

		Intervals containing more than 15 g/t Au should be individually investigated to detect possible contamination in intervals below the high-grade intervals.


Project No. 147107		Page 14-17

Table 14-2: Intervals with Possible Down-hole Contamination


Hole	From	To	Hole	From	To

RC1	30	54	ST159	12	42
RC2	156	174	ST163	24	54
RC4	24	48	ST167	6	24
RC5	36	54	ST178	30	54
RC13	114	132	ST197	6	30
RC23	30	42	ST206	12	36
RC23	84	108	ST214	6	30
RC35	96	114	ST233	0	24
RC41	34.00	72	ST241	12	36
RC51	120.00	138	ST262	6	24
RC51	144.00	162	ST265	12	30
RS1	102.00	138	ST319	12	42
RS23	54.00	72	ST320	6	30
RS49	72.00	90	ST344	0	24
RS53	126.00	156	ST345	0	18
RS64	36.00	66	ST410	6	24
RS70	138.00	156	ST478	30	54
RS73	24.00	36	ST494	40	56
RS73	62.00	54	ST502	30	54
RS95	42.00	78	ST509	6	30
RS113	66.00	84	ST514	24	42
RS128	24.00	48	ST536	6	24
RS132	60.00	84	ST547	18	36
RS139	42.00	66	ST555	40	60
RS141	12.00	36	ST556	18	42
RS147	54.00	84	ST562	6	36
RS149	36.00	54	ST599	25	36
RS150	126.00	144	ST600	42	60
ST66	6.00	36	ST603	6	36
ST113	18.00	36	ST606	24	36
ST125	12.00	36	ST637	48	60

14.6 Density

		Because the SG testing completed by other companies was well documented, PDDC did not conduct any further testing in this area.

		AMAX Engineering and Mining Services derived a linear regression formula for density based on 152 pairs of density and sulphur samples from the GR series of diamond drill holes in 1985. The regression formula is:

		Density = (0.0322 * sulphur %) + 2.617

		Fluor checked the regression equation during its 1986 Feasibility Study. Fluor used the equation to estimate the in situ bulk density of a bulk sample from the Moore deposit for which both density and sulphur content had been previously calculated by laboratory analysis. The regression equation predicted a very similar specific gravity. Fluor noted,


Project No. 147107		Page 14-18

		however, that all but 13 of the density samples were from Pueblo Viejo sediments in the Moore deposit. At the request of Fluor, Rosario personnel measured specific gravity and sulphur content for 34 spilite, 13 volcaniclastic, and 20 conglomerate samples. The results showed that the regression difference for the other lithologies was minor, and Fluor accepted the density equation.

		In 1997, GENEL confirmed the accuracy of the density formula using 100 core samples from its own drilling program. The 20 cm to 30 cm long samples were weighed in air and then reweighed immersed in water. The samples were then sent to the Chemex laboratory in Vancouver for sulphur analysis. Samples were grouped into three sulphur ranges: 0-8%, 8-15%, and >15%. Within each range, all of the principal lithologies were represented. The results showed that the density formula is accurate to within 5% for all sulphur ranges and lithologies.

		AMEC compiled all of the available density data and confirmed the above equation.


Project No. 147107		Page 14-19

15.0 ADJACENT PROPERTIES

There are no adjacent properties of significance to the Pueblo Viejo Project.


Project No. 147107		Page 15-1

16.0

MINERAL PROCESSING AND METALLURGICAL TESTING

16.1		Introduction

		The information provided and reviewed in this section and also in Sections 19.2 and 19.3.8 is substantially based on the PDTS “Pueblo Viejo Project Feasibility Study, Volume 3, Metallurgy, July 2005.”

16.2		Metallurgical Studies Prior to Placer Dome

		The metallurgical history of the Pueblo Viejo project has been summarized by Pincock, Allen & Holt (PAH). Defining an economic processing method for the sulphide phase of this project has been challenged by the high sulphur to gold ratio in mineralization and by the wide distribution of gold in various mineralogical facies not easily concentrated with one metallurgical process.

		Table 16-1 is duplicated from PAH and summarizes the extensive efforts that have been made by previous owners and consultants to develop an economic processing concept for Pueblo Viejo. The valuable constituents of the ore are summarized for the average assay in Table 16-2, valued at October 2005 metal prices.

		The base metals have been studied as potential contributors to project economics but the cost to recover these (zinc in particular) has led to rejection of this option. Gold is the economic driver of the project. The focus of historical and present metallurgical testwork has been on gold recovery. From Table 16-1 (PAH), the highest gold recovery (88%) in pre-Placer Dome testwork was for a flotation concentrate. This is not a final gold product, however, and would accrue additional recovery penalties as well as various smelting charges if it could be sold to a smelter.

		The PAH table shows that other than production of a flotation concentrate, gold recoveries using the approaches tested prior to 2003 show gold recoveries in the range of 80% to 84%. Each of the concepts listed in the table that is capable of yielding gold recovery in this range involves expensive destruction of the sulphide content in concentrates, either by roasting or oxidative leaching, while rejecting 16% to 20% of the gold, or approximately $7.80 to $9.75/t of ore.


Project No.: 147107		Page 16-1

Table 16-1: Summary of Metallurgical Test Programs (from PAH)


				Indicated Recoveries
				(%)
Entity	Years		Processes Examined	Au		Zn		Ag
Lakefield Research		1973-1993	Differential flotation of pyrite and zinc concentrates; roasting of pyrite concentrates; cyanidation, CCD; Merrill Crowe (MC)		80		80		30
Hazen Research		1977-1981	Flotation; roasting; cyanidation; pressure oxidation		70		69		70
Fluor Engineering (Pre-feasibility) (four alternatives)		1983	Bulk flotation; bulk concentrate roasting; sulphuric acid; CCD, MC Bulk flotation; autoclave bulk concentrate, CCD, MC Bulk flotation; partial bulk concentrate roasting-autoclave; CCD, MC Bulk flotation concentrate		75 84 80 88		0 0 0 N/A.		33 30 80 80
Amax Extractive Research & Development		1984-1986	Whole ore roasting; sulphuric acid production; CIL, MC		82.5		0		26
Fluor Engineering (Feasibility)		1988	Whole ore roasting; sulphuric acid production; CIL; MC		82.5		0		26
Stone & Webster/AMS (Pre-feasibility)		1992	Whole ore roasting; SO₂ neutralization; CIL; MC		82.5		0		26
Davy International (Feasibility) (two alternatives)		1993	Whole ore roasting (Lurgi and Fuller methods); SO₂ neutralization: CIL; MC Bulk flotation; fine grinding/cyanidation of concentrate; zinc flotation; CIL, MC		83 64		0 1		35 50
MIM Holdings		1995-1997	Fine grinding, Albion Process N/A		N/A		N/A		N/A
GENEL		1996-1997	Biooxidation of bulk sulphide concentrate; CIL; MC		N/A		N/A		N/A
Resource Development Inc. (RDI) (Flow Sheet 1) ²		2001	Zinc flotation; fine grinding of zinc cleaner tailings; cyanide leaching of zinc cleaner tailings and rougher tailings; CIL, Merrill Crowe		55		80		55
Resource Development Inc. (RDI) (Flow Sheet 4) ²		2001	Zinc flotation; fine grinding of zinc cleaner tailings; biooxidation, and cyanidation leaching of zinc cleaner tailings and rougher flotation tailings; CIL, MC		70		80		70


¹		Not quantified in the report .

²		Recoveries shown do not include incremental recovery of ±3% when CIL slurry is heated.

Table 16-2: Valuable Constituents of Ore (October 2005)



	Current Price						Value at 100% Recovery
Metal	US$			Nominal Ore Content			US$/t

Gold		460.00	oz		3.30	g/t		48.80
Silver		7.60	oz		18	g/t		4.40
Zinc		0.65	lb		0.79	%		11.30
Copper		1.85	lb		0.09	%		3.65

Total								68.15


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16.3

Placer Dome Metallurgical Testwork

16.3.1		Introduction

		Motivated by the high costs and low recoveries associated with the concepts listed above, Placer investigated other options:

	•		concentrate ultrafine grinding (UFG)

	•		concentrate stirred tank bio-oxidation (BIOOX)

	•		whole ore heap bio-oxidation

	•		concentrate coated Geocoat® heap bio-oxidation

	•		concentrate pressure oxidation.

Each of these oxidative and/or liberation processes was to be followed by carbon-in-leach cyanide leaching to dissolve and recover the gold onto activated carbon. All of these alternative processes suffered from one or more of the following problems:

	•		in pre-concentration by flotation, poor gold recoveries to a concentrate with a high mass pull (30% to 40%)

	•		poor gold recovery in CIL

	•		ineffective and incomplete gold liberation

	•		slow return of revenues

	•		inconsistent process performance

	•		process royalty payments

	•		high technical risks.

		Ultimately, Placer Dome selected a fairly straightforward process based on pressure oxidation of the whole ore followed by carbon-in-leach cyanidation for recovery of the gold and silver. One innovation, a hot cure on the slurry from the autoclave, is designed to reduce lime consumption by solid basic ferric sulphate in the cyanide leaching circuit in favour of the neutralization of dissolved ferric sulphate in the liquid phase by limestone in the HDS circuit. Autoclaving of the whole ore entails greater capital and operating costs than treating a concentrate, but the increment in operating cost is limited because of its dependence more on sulphur throughput rather than total solid throughput. The increment in gold recovery with whole ore pressure oxidation is significant: gold recoveries are projected to vary with ore type from 89.4% to 95.5%. These projections are supported by an extensive test program completed at SGS Lakefield in Peterborough, including a 10-day continuous autoclave and cyanidation pilot-plant campaign, which reported 93.7% overall extraction (discussed below).


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16.3.2		Whole Ore Pressure Oxidation

		Pressure oxidation testwork was started on a batch scale at the Placer Dome Research Centre in July 2003. It was moved in September 2003 to SGS Lakefield Research, where additional batch-scale testwork and extensive pilot-plant campaigns were performed. All of the testwork was completed by 2004.

		To ensure a fast oxidation rate and so minimize the size of the autoclaves, the operating conditions selected for pressure oxidation were an operating temperature of 230°C, an oxygen partial pressure of 100 psi, and a retention time of 60 to 90 minutes.

		Cyanide consumption in CIL was reasonable, but in early testwork, lime consumption was extremely high, up to 90 kg CaO/t, especially when pressure oxidation was run at a high pulp density. The high lime consumption was reduced to 10 kg CaO/t or less in early 2004 by introducing a hot curing step after pressure oxidation.

		Significant preg-robbing remained after pressure oxidation for the MO-BSD (Moore sedimentary host rock) and MN-BSD (Monte Negro sedimentary host rock) ore types. Minor preg-robbing was evident with the MN-VCL (Monte Negro volcanic host rock) ore type. Silver recovery was poor, ranging from 14% to 41% without hot curing. Copper and zinc minerals dissolved completely in the autoclave and remained in solution after discharge. Most of the iron and aluminum in the ore dissolved and then precipitated within the autoclave. Free acid concentration in the autoclave discharge liquor was relatively constant, regardless of the feed pulp density, while soluble sulphate concentration increased with increasing pulp density.

		A number of short but continuous pilot-plant runs were operated in late 2003 with the objectives of determining the rate of sulphide oxidation, reducing the CIL lime consumption, and improving the silver recovery. A continuous integrated pilot plant comprising autoclaving, hot curing, CCD washing, CIL, and cyanide detox circuits was operated for two weeks in April 2004. Each of the five ore types was tested individually.

		The last two continuous pilot plant campaigns were performed in June 2004. A continuous CIL circuit was operated with blended feed produced from five ore types. An HDS circuit was operated to neutralize the acidic liquor generated during pressure oxidation, to determine the density of neutralization precipitate, and to confirm the lime and limestone requirements. The continuous pilot-plant campaigns confirmed the high gold recoveries obtained in batch scale testwork, confirmed the cyanide (1.0 kg NaCN/t) and lime (10 kg CaO/t) consumptions in CIL, delineated the limestone and lime requirements for neutralizing the autoclave acidic liquors and acid rock drainage, and developed a successful flowsheet for maximizing the density of the neutralization sludge prior to disposal in the tailing pond.


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16.3.3		Limestone Calcining and Slaking Testwork

		Four limestone samples collected in 2004 at Pueblo Viejo site were submitted to Maerz in Switzerland for calcining and slaking tests. Two additional limestone samples were collected in April 2005 from Los Lagunas Quarry, and were subject to calcining and slaking tests by Maerz. Assays of these limestone samples indicated calcium carbonate (CaCO₃)contents of 96% to 99%.

		Maerz observed that these samples were composed mainly of a fine crystalline grain structure. The following tests were completed on these samples:

	•		LOI to determine the loss of weight during burning

	•		shatter test to examine the mechanical strength of the burned lime

	•		decrepitation test to characterize the tendency of limestone decomposition during calcinations

	•		reactivity test according to ASTM-C110 (1993) to characterize the slaking reactivity of burned lime after burning at a temperature of 1,050°C.

		Based on the results of the shatter and decrepitation tests, the burned limestone samples had a high mechanical stability. The reactivity test results demonstrated that the burned lime was highly reactive with an available CaO content around 94% to 96%. Maerz confirmed that all six of these limestone samples would be a suitable feed to the Maerz PFR kilns.

		Testing was also completed to evaluate grindability and abrasiveness of the limestone for use in the process calculations.


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17.0

MINERAL RESOURCE AND MINERAL RESERVE ESTIMATES

17.1		Summary of Conclusions

		Placer Dome Technical Services Limited (PDTS) in Vancouver, B.C., produced the resource models for the Pueblo Viejo Project for PDDC as part of an internal Feasibility Study. The models were completed by Christopher Keech, PGeo, in early 2005 using the data available at that time. The database includes drilling from Rosario, MIM, GENEL, and PDDC. Two separate models were produced, one for each of the deposits at Pueblo Viejo—the Moore deposit and the Monte Negro deposit.

		A Measured and Indicated Mineral resource of 150.6 Mt grading 3.15 g/t Au and 17.1 g/t Ag at 1.7 g/t Au cutoff was estimated for the Monte Negro and Moore deposits. The Mineral Resource contains 15.3 Moz Au and 82.8 oz Ag. There is also a small additional Inferred Mineral Resource of 2.2 Mt grading 2.9 g/t Au and 12.6 g/t Ag. Mineral resources were estimated using assays from 314 core holes, 64 reverse circulation holes and 195 rotary holes, for a total of 573 drill holes totalling 80,261 m of drilling. All quoted Mineral Resources are inclusive of Mineral Reserves.

		As part of AMEC’s independent review, grade interpolation parameters were examined for suitability and the models checked for validity.

		*Comment*

		AMEC supports the mineral resource declared by Placer Dome and believes it has been developed using methods consistent with accepted industry practices. Furthermore, AMEC believes that the mineral resource is of sufficient quality to support the declaration of a reserve contained within a design pit and using appropriate economic factors.

17.2		Methodology

		The resource models for the Moore and Monte Negro deposits were developed separately and then later combined into a single model for pit optimization. The basic modelling methodology was virtually identical for each of the deposits. Three different metals were modeled: Au, Ag, and S.

		The rate at which the ore can be processed is directly proportional to the sulphur content, which makes sulphur assays critical to the mine plan and ultimately the cash flow. A significant number of sulphur determinations are missing on the samples in the Moore deposit. An adjustment was required to in-fill the missing sulphur data for Moore. A simple regression formula between Au and S was developed, per geological domain, to calculate the missing sulphur grades so that each gold grade had a measured or calculated sulphur


Project No.: 147107		Page 17-6

		grade. These “calculated sulphur grades” were kriged by PDTS independently, and the addition of calculated sulphur grades improved the Moore S grade model so that the Au-S conditional distribution better matched the original data. As the Monte Negro drill hole data has very few missing sulphur grades, no adjustments were required for this model.

		Each of the variables was estimated independently using geologically constrained ordinary kriging. The data analysis and resource estimation were done using a variety of software tools. These included the PDTS in-house software OP and GEOLOG for statistical analysis, variography, and kriging; Maptek’s Vulcan for 3D geological interpretation and visualization; and SAGE2001, by Ed Isaaks for calculating variograms.

		A total of 80,261 m of drilling in 573 holes was used for resource estimation. The holes are summarized by deposit and type in Table 17-1. Plan maps showing the drilling for Moore and Monte Negro are shown in Figures 11-1 and 11-2.

		Table 17-1: Number of Drill Holes and Metres by Drilling Type


Type	Moore				Monte Negro				Total
	No. of Holes		Metres		No. of Holes		Metres		No. of Holes		Metres

Core		185		27,271		129		18,231		314		45,502
RC		144		17,415		51		7,342		195		24,757
Rotary		20		2,821		44		7,181		64		10,002

Total		349		47,507		224		32,754		573		80,261

		Following estimation, PDTS validated the models statistically. The models were then used to generate a pit shell using a $450 per troy ounce gold price. PDTS summarized the resources within these pit shells.

		The topographic surface for the model is based on a 1997/98 set of aerial photographs taken by GENEL. The available contours were at 4 m intervals.

		The density model was based on a linear regression relationship between sulphur grade and density that was initially developed by Rosario but has been validated with data collected by GENEL. The following formula is used to calculate specific gravity:

		Density = (0.0322 * Sulphur Grade(%) ) + 2.617

		The block model estimate of the sulphur grade was used to calculate a density value for each block in the model.


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17.2.1		Geological Modelling

		Geological models were produced for both the Moore and the Monte Negro deposits. The geological models consisted of a lithological model, a structural model, and an alteration model. These were then superimposed upon each other to define the statistical domains. The geological models were developed in cross-section and rectified in three dimensions by overlaying the cross-sections. The interpretations were then wire-framed to create a three-dimensional model. The block model and drilling composites were then coded using the wire-framed model.

		*Moore*

		PDTS defined seven lithological units for Moore (Figure 17-1 and Table 17-2). These units were then further subdivided by five structural domains, primarily defined by faults. The structural domains were defined by PDTS using surface mapping and data from oriented core. To further define the ore controls, PDTS interpreted the contact between advanced argillic (silicification, quartz-pyrophyllite, pyrophyllite) and propylitic alteration. Resources were interpolated within the advanced argillic zone.

		The different lithology, structure and alteration zones were then combined together to create 15 unique estimation domains (Table 17-2). It can be seen that a majority of the mineralized blocks are defined by a sequence of six sedimentary units, split by the F3 Fault into the Central and Eastern domains.

		PDTS investigated the trend of grades across boundaries between different units with three different methods: contact plots, cross-correlograms, and geo-bound plots. As the result of these analyses, the boundaries between units 3, 4 and 5 (also 9, 10 and 11) were treated as soft boundaries while the remaining boundaries were treated as hard boundaries.

		*Monte Negro*

		PDTS defined seven lithological units for Monte Negro. Of these three contain the majority of the mineralization: Spilite, Conglomerate, and Pueblo Viejo Sediments. There are six structural domains which were defined at the Monte Negro deposit, three of which contain significant mineralization. The procedures to build the Monte Negro alteration and vein models are similar to those described above for the Moore Deposit. The boundary between advanced argillic and propylitic alteration was used to constrain the estimate. Combining the structural and the lithological domains produces fifteen defined statistical zones. The Monte Negro zones are listed in Table 17-3. A typical Monte Negro cross-section is shown in Figure 17-2



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Table 17-2: Zone Names in Monte Negro Deposit


Domain	Names	Litho	Struct	Alter	Description	Block Count

1	HW-PVS	6	1	2	Hanging Wall Sediments		228,405
2	MEJ	1	2+3	2	Southwest + Central Mejita Sediments		149,991
3	VCL	2	2+3	2	Southwest + Central Volcaniclastics		584,533
4	VSS	3	2+3	2	Southwest + Central Volcanic Sandstone		221,490
5	SPY	4	2+3	2	Southwest + Central Spilite		34,870
6	MXD	5	2+3	2	Southwest + Central Mixed Sediments		42,861
7	PVS	6	2+3	2	Southwest + Central Pueblo Viejo Sediments		232,051
8	EST-MEJ	1	4	2	Eastern Mejita Sediments		14,634
9	EST-VCL	2	4	2	Eastern Volcaniclastics		177,739
10	EST-VVS	3	4	2	Eastern Volcanic Sandstone		42,354
11	EST-SPY	4	4	2	Eastern Spilite		12,868
12	EST-MXD	5	4	2	Eastern Mixed Sediments		21,171
13	EST-PVS	6	4	2	Eastern Pueblo Viejo Sediments		125,017
14	FDP	7	all	all	Fragmental Dacite Porphyry		105,855
15	CHL-PP	all	all	1	Chlorite Propylitic Alteration		1,970,628

Figure 17-1: Moore Section 94600 N Lithology and Structure



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Table 17-3 Monte Negro Zone Names


Domain	Name	Litho	Struct	Alter	Description	Block Count

1	WBF	all	9	1	Western Boundary Fault		47,697
2	WMF-MNS	1	1	2	West Monte Negro Fault- Monte Negro		47,578
					Sediments
3	WMF-CGL	2	1	2	West Monte Negro Fault -- Conglomerate		75,345
4	WMF-SPY	3	1	2	West Monte Negro Fault -- Spilite		190,026
5	MNF-MNS	1	2-5	2	Monte Negro Fault - Monte Negro		20,314
					Sediments
6	MNF-CGL	2	2-5	2	Monte Negro Fault - Conglomerate		24,605
7	MNF-SPY	3	2-5	2	Monte Negro Fault - Spilite		115,003
8	EF5-MNS	1	6	2	East F5 Fault -- Monte Negro Sediments		47,861
9	EF5-CGL	2	6	2	East F5 Fault -- Conglomerate		44,599
10	EF5-SPY	3	6	2	East F5 Fault - - Spilite		448,929
11	Breccia	4	1-6	2	Hydrothermal Breccia		6,683
12	Dyke	5	1-6	2	Andesite Dyke		33,330
13	EBF	all	8	all	Eastern Boundary Fault		22,987
14	NEF	all	7	all	North East Fault		32,388
15	PP-ALTER	all	1-6	1	Chlorite-Calcite propylitic alteration (unaltered)		833,689

Figure 17-2: Monte Negro Section 95800N Lithology and Structural Domains



Project No.: 147107		Page 17-10

Geological domain boundary conditions were reviewed using the same graphical techniques as Moore, and the boundaries were chosen so that the CGL and SPY domain boundaries are soft, and the CGL-SPY and MNS boundary is hard.

Comment

The geological models defined by PDTS appear to do a very reasonable job of defining the appropriate ore controls. AMEC believes that the methodology followed by PDTS of independently defining lithological, structural and alteration models and then combining these models to define the estimation domains is strong practice that should lead to superior domain models.

17.2.2		Data Compositing

		The assay data was combined into fixed length 2 m composites. More than 95% of the samples are between 2 m and 3 m in length and 82% are exactly 2 m long. A composite length of 2 metres was chosen so that data would be regularized, but not be smoothed by the compositing.

17.2.3		Data Analysis

		*Top Cutting*

		Since ordinary kriging was used for all grade estimates, it was necessary to top-cut the assay distributions to reduce the influence of high-grade “outlier” assays, and control their influence during grade interpolation. To determine the most appropriate threshold to cut or cap grades, a variety of statistical graphs were used, including histograms, log probability plots, “cutting statistics” graphs; which include an indicator correlation graph, a coefficient of variation graph and a contained metal graph; and decile analysis graphs.

		The procedure to select a cutting threshold was to examine each graph independently and based on that graph alone make an appropriate selection as a possible cutting threshold. The results from each graph were then tabulated and reviewed to determine a single cutting threshold for each variable for each domain.

		Within Moore, a total of 94 out of 21,855 2 m gold composites grades were cut to the selected cutting limit, while 112 of 21,853 2 m silver composites were cut and 140 of 16,179 2 m sulphur composites were cut. This represents approximately 0.4 % of the gold composites, 0.5% of silver composites and 0.9 % of the sulphur composites were adjusted by cutting. The number of total composites adjusted by cutting is low and has not significantly affected the global mean grades of Au, Ag, and S.



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Within Monte Negro, a total of 92 out of 14,453 2 m gold grades were cut, while 150 of 14,453 silver composites were cut, and 112 of 14,156 sulphur composites were cut. The results of the cuts within Monte Negro are greater than in Moore. A little less than 2% of the gold and in excess of 5% of the silver was eliminated by the top cuts.

Comment

AMEC has reviewed the cutting statistics used by PDTS and the thresholds that were applied to the gold, silver and sulphur values. AMEC agrees with the thresholds used and believes that they are appropriate for the estimation method used to build the resource model.

17.3		Variogram Analysis

		PDTS examined the spatial continuity of the Au, Ag, and S using correlogram maps and directional correlograms. PDTS found it necessary to lump together a number of similar zones to acquire enough data to calculate a stable variogram.

		The correlogram models were initially fitted by the auto-fit function of SAGE2001 software to produce a single anisotropic exponential model with a nugget effect, three rotations, and three ranges. PDTS then confirmed the fitted models with its directional variogram software.

		The variogram models were validated by using a “jack-knife” technique to cross-validate the fitted models. The variogram models for gold, silver, and sulphur were adjusted to provide the best cross-validation statistics.

17.4		Interpolation Plans

		The basic interpolation plan was similar for gold, silver, and sulphur for both Moore and Monte Negro. In all cases, 2 m composites were used. A minimum of 8 and a maximum of 16 composites were used. The geological boundaries were generally treated as soft between the mineralized domains. For gold, silver, and sulphur estimation, anisotropic search radii with anisotropic distance calculations were used based on the variogram models. The search radii were set equal to the “effective range” of the fitted exponential variogram models.

17.4.1		Moore

		The Moore model limits (Table 17-4) were established to encompass the Moore mineralization. A standard 10 m x 10 m x 10 m block size was used.



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Table 17-4: Moore Model Limits


	Minimum		Maximum		No. of Blocks		Block Size

Northing		94,250.0		95400.0		115		10.0
Easting		75,500.0		76500.0		100		10.0
Benches		-100.0		490.0		60		10.0

17.4.2		Monte Negro

		The Monte Negro model limits (Table 17-5) were established to encompass the Monte Negro mineralization. The block size was a 10 m x 10 m x 10 m.

		Table 17-5 Monte Negro Model Limits


	Minimum		Maximum		No. of Blocks		Block Size

Northing		95,050.0		96,200.0		115		10.0
Easting		74,700.0		75,350.0		65		10.0
Benches		0.0		490.0		50		10.0

Comment

A small block size can produce a sense of high spatial resolution of grades that may not exist given the actual drill hole spacing. This can lead to overly optimistic expectations in mining selectivity that might not be achievable in practice. It is not uncommon within the mining industry for this to be done, but it is not considered the best practice.

The model was estimated with a minimum of 8 and a maximum of 16 samples. With 2 m composites, this means that the blocks were estimated with between 16 and 32 m of a sample. With a 10 m bench height, this would mean that each block was estimated with only 2 or 3 bench sized composites. This sample support is inadequate to provide reasonable local estimates. This does not appear to have biased the global estimate, but in combination with the small block size could produce some local anomalies in the mine design and mine plan that will need to be managed with good grade control.

The small block size may result in loss of some tons and grade in peripheral material in the pit where local mining limits will be naturally adjusted to mine the best material. The small block model is not globally biased, so as a large amount of material is mined the estimated global ratio of ore and waste should be correct. However, on a local basis, mining limits will be based on local values (not global values) and these, if optimistic, will lead to losses in tonnes and grade. The total effect cannot be measured without comparing the small block model to a pit designed with larger blocks. AMEC believes the only material at risk is mineralization in the periphery of the deposit where local positive and negative biases may



Project No.: 147107		Page 17-3

not be contained in enough tons to balance out. Losses of ounces in the peripheral zones may decrease cash flows in this part of the deposit.

17.5		Model Validation

		PDTS’s validation of the Au, Ag, S, and calculated S grade models consisted of five different checks of the results.

	•		global mean check

	•		global variability check

	•		trend check/swath plots

	•		comparison with ST holes

	•		visual inspection

All checks were done for gold, silver, and sulphur. All checks were summarized by estimation domain. The global mean check compared the ordinary kriged estimate with a nearest-neighbour estimate. The nearest-neighbour estimate will provide a declustered estimate of the global mean. In all cases, the resource estimate appears to be globally unbiased.

The next validation was to check the variance reduction imposed on the resource model by the ordinary kriging. This was done by comparing the variance of the nearest neighbour model to the variance of the estimated blocks. It is desirable that a degree of smoothing be imposed on the model such that the resource estimate will represent the distribution of tons and grade that would be expected to be found at the cutoff grades used in mining. With a small number of exceptions, the variance reduction seen with the estimates in the various domains was within the desired range.

The trend check used swath plots to examine the spatial smoothing horizontally and vertically. The swath plots indicate that there has not been any undue smearing of the grades spatially throughout the deposit.

A number of short transition (ST series) holes were drilled to test the near surface mineralization. Because of the spatial clustering of these holes, they were not used in the resource model estimate. These holes were compared to the model to provide an independent check of the model. The comparisons were reasonable and supported the modeling procedures used.

The final validation of the model was a visual inspection of the model results. A complete set of cross-sections was produced for both Moore and Monte Negro. The cross-sections were inspected comparing the estimates to the drilling used to produce the estimate. Overall, the model appears to fairly represent the drilling.



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Comment

Globally the Moore and Monte Negro models appear to be unbiased. They also appear to be conditionally unbiased within the range of operating cutoffs. This implies that the resource models should provide a reasonable estimate of what will be achieved from mining on a global basis. Ordinary kriging is a smoothing estimator and when used with skewed distributions, as seen at Pueblo Viejo, will often over-smooth the estimate, resulting in significant conditional biases and potentially very biased estimates at the operating cutoff. To prevent these problems, the degree of smoothing resulting from the ordinary kriging was controlled within these models by using short composites and by estimating with a very small total length of samples. These techniques ensure a properly smoothed model but may yield very poor local estimates.

The use of ordinary kriging with skewed sample distributions is not unusual within the mining industry; but it is difficult to achieve conditionally unbiased results. PDTS has gone to extraordinary efforts to ensure that the resource estimate is unbiased. However, due to the inherent weaknesses in ordinary kriging methods, there still is some risk that biases will be present on a local basis.

17.6		Resource Classification

		The resource has been classified as a Measured, Indicated, and Inferred by PDTS. Because of the uncertainty associated with the historical nature of the initial data, none of the resource was previously classified as Measured. With the additional drilling completed in 2004 by PDDC, the historical drilling has been confirmed and additional fill-in drilling added. A significant portion of the resource is now classified as Measured.

		Resource classification was based on the kriging estimation variance from the gold estimation. The estimation variance is generally a reasonable relative measure of how well-informed a block estimate is because it considers the spatial correlation between the data and the block being estimated, as well as the spatial correlation between the data being used in the estimate. In general, when estimates are interpolated between data points, the estimation variance will remain relatively stable. When the estimator begins extrapolating away from data, the estimation variance will rapidly rise. PDDC chose the estimation variance threshold to try to classify interpolated material as Indicated. The classification statistics are shown in Table 17-6. The table shows the average data support for the various classifications of the resource. For example, the Measured Resource in Moore was estimated on average with 12 to 13 composites from 2 to 3 different drill holes from 3 to 4 octants with an average distance of 27.2 m from the block being estimated.



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The classification is illustrated in Figure 17-3 for the Moore deposit and in Figure 17-4 for the Monte Negro deposit.

Table 17-6: Resource Classification and Estimation Statistics


	Class		No. Blocks		Avg Kvar		Avg dist		Avg ncmp		Avg nhol		Avg noct
Measured
Moore		1		42,367		0.333		27.230		12.678		2.396		3.908
Monte Negro		1		25,361		0.240		25.019		13.390		2.396		4.151

Indicated
Moore		2		14,562		0.596		43.596		11.929		2.362		3.497
Monte Negro		2		5,911		0.514		39.747		10.747		2.015		3.304

Inferred
Moore		3		1,528		0.811		63.610		9.696		1.696		2.705
Monte Negro		3		938		0.782		55.101		8.164		1.224		2.285

The resource classification was further validated by simulation studies that supported the estimation variance criteria used. The simulation studies examined multiple realizations of simulated panels representing quarterly and annual production. The criteria used in the simulation study were to be within 15%, 90% of the time on a quarterly basis for Measured and on an annual basis for Indicated. Using these criteria, the proportion of Measured and Indicated found were in approximately the same proportion as that of the resource.

Comment

AMEC agrees with the use of the estimation variance as a relative measure of the quality of the estimate. The simulation study lends quantitative support to the thresholds chosen for the classification of the resource. The higher-grade portion of the resource has a slightly higher drilling density than the rest. This lends some added confidence to the estimate of the economic core of the resource.



Project No.: 147107		Page 17-16

Figure 17-3: Moore Section 94500N Showing Mineralization and Resource Classification



Project No.: 147107		Page 17-17

Figure 17-4: Monte Negro Section 95600N Showing Mineralization and Resource Classification



Project No.: 147107		Page 17-18

17.7		Mineral Resource Statements

		The two resource models were combined into a single model used for the definition of pit limits and resources in the July 2005 Feasibility Study. A pit based on $425/oz Au was used to define the limit of material that meets the criterion of “reasonable prospects for economic extraction.” The corresponding cutoff grade for the $425 pit is 1.7 g/t Au. The PDTS resource summary at this breakeven cutoff is shown in Table 17-7. The Indicated Resource at various cutoffs is summarized in Table 17-8. Cross-sections of the Moore and Monte Negro pits resource are shown in Figures 17-3 and 17-4, respectively.

Table 17-7: Resource Summary at 1.7 g/t Au Breakeven Cutoff


	Tonnes		Gold		Gold		Silver		Silver
Category	(kt)		(g/t)		(Moz)		(g/t)		(Moz)

Measured		118,609		3.22		12.3		18.14		69.2
Indicated		31,965		2.90		3.0		13.28		13.7

M+I		150,574		3.15		15.3		17.11		82.8

Inferred		2,155		2.88		0.2		12.64		0.9

For the purpose of stating Mineral Resources exclusive of Mineral Reserves in its 27 September announcement of its investment decision, Placer Dome modified the criteria for resources to include material within a pit based on a $450 gold price and a 1.5 g/t Au cutoff grade. These resources are external to Mineral Reserves contained within a pit designed with a $375 gold price and tabulated as blocks with a positive value at $400/oz Au price. Table 17-8 lists Mineral Resources as of 27 September 2005.

Table 17-8: Mineral Resources, 27 September 2005, Exclusive of Mineral Reserves


	Tonnes		Gold Grade		Contained Ounces
Category	(Mt)		(g/t)		(000s)

Measured Mineral Resources		22.9		2.20		1,618
Indicated Mineral Resources		11.0		2.38		846
Measured + Indicated Resources		33.9		2.26		2,464

Inferred Mineral Resources		2.2		2.72		190

Note: Mineral Resources are contained within a pit designed using an average cutoff grade of 1.5 g/t Au and a gold price of $450/oz.

17.8		Mineral Reserves

		Mineral Reserves were estimated using Measured and Indicated Resources and appropriate design criteria. The mining limits, mine and quarry production schedules, mine operating costs, mining equipment requirements, mining capital, and mine sustaining costs were developed by PDTS based on first principles and experience with their other



Project No.: 147107		Page 17-19

operating properties. The mine production schedule uses a staged mine development. Cutoff grades vary by mining stage and were selected to optimize gold production given the ore tonnage and sulphur constraints in the plant. These aspects are discussed in more detail in Section 19 of this report.

Proven and Probable Mineral Reserves as of 27 September 2005 are listed in Table 17-9.

Table 17-9: Proven and Probable Mineral Reserves, 27 September 2005 at $400 Gold Price


		Waste Tonnes		Ore Tonnes		Gold Grade		Contained Ounces
Deposit	Category	(Mt)		(Mt)		(g/t)		(000s)

Moore	Proven		31.5		60.5		3.26		6,338
	Probable		14,5		24.3		2.99		2,335

Monte Negro	Proven		11.6		35.4		3.32		3,780
	Probable		5.2		9.5		3.17		970

Total	Proven		43.0		95.9		3.28		10,118
	Probable		19.7		33.8		3.04		3,305

	Proven + Probable		89.5		129.7		3.22		13,423

A total of 89.5 Mt of waste material will be mined, for a life-of-mine average strip ratio of 0.69 waste:ore.



Project No.: 147107		Page 17-20

18.0		OTHER RELEVANT DATA AND INFORMATION

		This section is not pertinent for this report.



Project No.: 147107		Page 18-1

19.0

REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT AND PRODUCTION PROPERTIES

19.1

Mining Operations

19.1.1		Introduction

		Placer Dome Technical Services (PDTS) has completed a number of studies investigating the feasibility of the Pueblo Viejo Project, culminating in a Feasibility Study completed in July 2005. These investigations have evaluated various throughput options and mining cutoff grade strategies. PDTS has shown that modestly positive economics can be achieved for the project at a $400/oz Au price, the pricing level PDI has chosen for its year-end reserve statements. The Feasibility Study shows that for a positive outcome for the project to achieved, the following are necessary:

	•		A minimum throughput rate of 15,000 t/d is necessary. Lesser throughput rates investigated did not achieve the desired results.

	•		An aggressive cutoff grade strategy must be implemented to process the highest grade ore early in the mine life with lower grade ore stockpiled for processing later in the mine life. The ability to implement this aggressive strategy is critical to success of the project. PDTS has demonstrated that this should be achievable.

	•		Proper blending of the ore will be necessary to maximize the utilization of the process facility and achieve required gold production. This will require much diligence in grade control and blending, but PDTS has demonstrated that it should be achievable.

The July 2005 Feasibility Study evaluated the project to the level of detail that would be expected for this level of study. The work done has been very thorough and completed to a high professional standard. Potential problems have been recognised and addressed within the study.

Comment

PDTS has thoroughly considered many of the potential pit-falls that may exist with the project, but no project is without risks. Those particular to Pueblo Viejo include the requirement for an aggressive cutoff grade strategy early in the mine life and the necessity of successful ore blending. The work that PDTS has done has gone far to minimize these risks. Given the project’s economics at the disclosed gold price assumption, there is very little cushion to absorb any shortfalls. PDTS has reasonably demonstrated that the aggressive cutoff grade strategy and the ore blending can be expected to be achievable, but it must be emphasized that project economics are highly dependent on doing so.



Project No.: 147107		Page 19-2

19.1.2		Mine Operations Summary

		Proposed mining operations consists of mining from two open pits: Moore and Monte Negro; and from three limestone quarries: Plant, Las Lagunas, and Los Quemados. The mines will follow a conventional drill-blast-load-haul mining sequence. Mining operations (equipment and manpower) will be interchangeable between the mines and quarries. The general layout of the mining and plant site is shown in Figure 19-1. The ore from the mines will be hauled to a crusher near the plant site or to the low-grade stockpile. The waste from the mines will be hauled to the El Llagal tailings facility. It is anticipated that all of the waste mined will be acid generating. Therefore, all of the potential acid drainage will be contained within the tailings facility. The limestone will be mined and used as road construction material, rock for the construction of the El Llagal tailings dam, and for processing and water management operations. The process will require high quality limestone for the production of lime.

		Figure 19-1: General Site Layout

As part of the Pre-Feasibility Study, completed in early 2005, PDTS developed two alternative mine plans with different throughputs of ore: one with an average milling rate of 15,000 t/d and the other at 10,000 t/d. The 15,000 t/d option demonstrated superior economics and was used as the basis of the Feasibility Study completed in July 2005.


Project No.: 147107		Page 19-3

The resource model, ultimate mining limits and mine plan were all updated as part of the Feasibility Study. The throughput for the process plant is limited by two constraints. The first is the crushing and milling capacity of the plant at 15,000 t/d on average and at a maximum of 18,000 t/d. The second is the oxidizing capacity of the three autoclaves at 1,200 t/d of sulphur. As is standard for this type of facility, it is necessary for the mine plan to balance both the ore tonnage and the contained sulphur tonnage to maximize the utility of the plant. This calls for careful blending of the ore over the life of the mine.

Mine plans were developed on an annual basis (except the first year, which was scheduled monthly). Within that timeframe, the blending of the ore was achievable without undue consideration. On a daily or weekly production basis, it can be anticipated that achieving a blend to maximize sulphur oxidation will be a key consideration. Optimization of mill tonnage and, in particular, sulphur oxidation, will require diligent mine planning during operations.

The Pre-Feasibility study also demonstrated that the use of elevated cutoff grades would be necessary throughout the mine life for the project to be economically viable. This strategy was pursued in the Feasibility Study. The mine operates at an elevated cutoff for the entire life of the active mining. After the mine is exhausted, the low-grade stockpiles are reclaimed and processed. The mine will be exhausted in a little less than ten years. The low-grade will be processed for an additional ten years. This yields a twenty year active project life.

The ore reserve was developed using only Measured and Indicated Resources (Section 17). The mining limits, mine and quarry production schedules, mine operating costs, mining equipment requirements, mining capital, and mine sustaining costs were developed by PDTS based on first principles and experience with their other operating properties. The mine production schedule uses a staged mine development. The cutoff grades vary by mining stage and were selected to optimize gold production given the ore tonnage and sulphur constraints in the plant. For the July 2005 Feasibility Study, Mineral Reserves were defined within a pit designed at a gold price of $375/oz. Mineral reserves defined in this manner are noted in Table 19-1.


Project No.: 147107		Page 19-4

Table 19-1: Mineral Reserves in Moore and Monte Negro Pits at $375 Gold Price


	Unit	HG		LG		Total

Ore Mined	Mt		54.0		68.9		122.9
Gold Grade	g/t		4.42		2.42		3.30
Sulphur Grade	%		7.74		7.50		7.61
Gold Recovery	%		93.2		90.2		91.5
Gold Recovered	Moz		7.1		4.8		11.9
Specific Gravity	t/m³		2.87		2.86		2.86
Waste Mined	Mt		—		—		96.3
Strip Ratio	—		—		—		0.78

Total Mined	Mt		—		—		219.2

Final reserve statements were developed in September 2005 to reflect more current gold prices. The pit shells defined using $375/oz Au was retained and the value of blocks internal to these shells were recalculated using a $400/oz Au price. This resulted in an additional 6.8 million tonnes being converted from waste to ore and a reduction of the life-of-mine average cutoff grade to 1.6 g/t Au. In addition, Measured Resource blocks estimated with one drill hole were reclassified to Indicated Resources, resulting in a reclassification of some Proven material to Probable Reserves.

Mineral Reserves for the project as of 27 September 2005 are summarized in Table 19-2.

Table 19-2: Proven and Probable Mineral Reserves for Moore and Monte Negro Deposits, 27 September 2005 at a $400 Gold Price


	Tonnes		Gold Grade		Contained Ounces
	(Mt)		(g/t)		(000s)

Proven Mineral Reserve		95.9		3.28		10,112
Probable Mineral Reserve		33.8		3.04		3,304
Total Proven + Probable Mineral Reserve		129.7		3.22		13,427

Note: Mineral Reserves are within open pits originally designed using a life-of-mine average cutoff grade of 1.7 g/t Au and a gold price of $375/oz. Mineral Reserves within this pit are then tabulated as blocks with a positive value at a $400/oz Au price.

The limestone resources have been estimated from drill core data for Los Quemados only. The Las Lagunas and Plant deposits were interpreted from regional geology maps, surface samples, and a few depth-probing drill holes. The data collected from Lagunas and Plant deposits confirms the Las Lagunas material is of kiln quality, but the Plant deposit is not. Further exploration will be required to finalise the reserve estimates and identify additional kiln quality resources. The resulting limestone Mineral Resource is summarized in Table 19-2.


Project No.: 147107		Page 19-5

Table 19-3: Inferred Limestone Mineral Resource by Quarry Site


Quarry	Tonnes (M)

Plant		16.6
Los Quemados		46.8
Las Lagunas		19.9

Total		83.3

Comment

The cutoff grades are varied by mining stage. It would be more logical for the cutoff grades to vary in time, so that at any point in time the cutoff being applied everywhere within the mine is the same. In practice this will likely happen. The rationalisation of the cutoffs with time would only improve the project economics, though the improvement would likely be marginal.

PDTS has tried a number of approaches to develop the cutoff strategy to maximize the profitability of the production. The approaches taken will not produce a truly “optimal” schedule but, as a result of the efforts taken, should be close. Any further optimization of the cutoffs would only improve the project economics, though the improvement would likely be marginal.

19.1.3		Mine Design and Mine Plan

		The final mining limits were derived using optimized pit shells based on an economic model such that all ore blocks had a minimum profit equal to the ore mining costs. The optimized pit shells were smoothed and designed to meet geotechnical slope constraints and haulage access requirements. The mining volumes were then broken into mineable stages, with the first stage representing the most profitable material progressing to the last stage being the least profitable material. There are six stages with Stage 1 and 5 in Monte Negro and Stages 2, 3, 4 and 6 in Moore. The final mining limits are shown in Figure 19-2.

		The mining stages were then scheduled using an elevated cutoff grade strategy. This is done to shift revenue to the early years of the mine life which will result in a more attractive discounted cash flow. The Pueblo Viejo project can accommodate this without significant consequences to the mining due to the relatively low strip ratio associated with Moore and Monte Negro. PDTS has determined that the best cash flow could be achieved by immediately processing the best 40% to 50% of the ore and stockpiling the remaining 50-60% for later processing. This was done on a stage by stage basis. Since the most profitable ore was contained in the first stage, the cutoff that split the ore in half was


Project No.: 147107		Page 19-6

naturally higher than that necessary for the later three stages. This produced a very reasonable progression of declining cutoff grades over the life of the mine.

The determination of the cutoff grade is complex because the operating cost and throughput are a function of the sulphur grade while the revenue is a function of the gold grade. The primary bottleneck in the process is the sulphur capacity of the autoclaves, secondarily is the capacity of the crushing and grinding circuit. The pre-feasibility study used a profit per sulphur percent metric for the cutoff grades. This proved reliable to put the most profitable material through the oxidation circuit but was problematical with lower grade, low sulphur material when the overall tonnage constraints were exceeded. The Feasibility Study changed to a “Profit-per-Day” metric. The advantage of this is that it considers both the 1,200 t/d sulphur production capacity and the 18,000 t/d maximum processing rate while getting the most profitable blocks through the process plant first.

Detailed mine plans were then built using the elevated cutoffs and the consideration of the production constraints. The ore was further subdivided into: low-sulphur ore, where the 18,000 t/d processing rate represented the bottleneck; average-sulphur, where both constraints are roughly balanced; and high-sulphur ore, where the 1,200 t/d of sulphur represented the bottleneck. The production schedules were then built to maximize the process profitability by blending the ores. Achieving proper blending of the ore will be a very important component of the mining to ensure the profitability of the project.

Comment

AMEC has reviewed the mine designs and mine plans. AMEC believes that the designs and plans are suitable and generally achievable as planned.

As is typical for this type of processing facility, PDTS realises that the proper blending of the ore is a critical component to the success of the project and have consequently expended quite bit of effort in analysing the distribution of the sulphur. Over the annual time frames used for scheduling, PDTS has demonstrated that achieving the blending requirements should not be a problem provided due attention is paid at the time of mining. But, they also note that the sulphur grade can change dramatically over short distances. This may require more aggressive blending than is currently anticipated which would increase operating costs or reduce the effective capacity of the process plant.

The primary risk associated with planning on operating with elevated cutoff grades is simply not being able to achieve a clean separation of the high-grade, low-grade and waste. If high-grade ore is interspersed within the low-grade or low-grade interspersed within the high-grade, it may be difficult to segregate the two without one significantly diluting the other. Alternatively, if the low-grade is a rind between the high-grade and the waste, it may be impossible to practically recover all of the low-grade. Most certainly, both of these situations will happen to some degree within the mine. PDTS has evaluated this


Project No.: 147107		Page 19-7

issue, primarily through simulation studies, and believes that the production schedules are achievable. AMEC agrees with this conclusion, but there is still a level of risk associated with this issue that is not normally associated with most mining projects.

19.1.4		Mining Operations

		Operating at elevated cutoffs will reduce the production life of the pits by accelerating the mining rate to maintain the 15,000 t/d processing rate with high-grade only and stockpiling low-grade ore. Over the life of actual mining, 44% or the ore will be processed and 56% stockpiled. The consequence of this is that the overall mining capacity will have to be increased to 225% of what would be necessary if all of the ore was processed at the time of mining. Since the starting strip ratio is low, capital requirements are not excessive to start with, so the impact on the overall project is not overwhelming. Also, because of economies of scale, the mining capital increase is less than the production capacity increase and the increased capital costs are partially offset by lower average operating costs.

		All pit waste material mined over Years 1 to 11 will be hauled +6.0 km to the El Llagal tailings dam for direct dumping along the Eastern shore the tailings pond. The waste mined to develop the access to Stage 2 in Year -2 will be used as ramp fill to connect the lower fixed haul roads with the primary crushers, the upper Moore pit benches with the waste hauls to El Llagal, and the sub-190 m benches in the Moore pit with the primary crushers.

		The mining strategy will allow the waste rock to be buried in the tailing pond and is an effective remediation of long term ARD issues associated with waste rock at this property.

		Open-pit production will be performed with conventional drilling, haul trucks, and loading equipment. Drill penetration studies have been completed to validate the drilling productivities. All of the haulage profiles have been examined for the life of the mine to determine truck and shovel requirements. Capital equipment requirements, including replacement and on-going capital, have been developed to sustain the production schedule. Expected costs of the capital equipment have been received from equipment suppliers. Operating costs for the mine operation have been developed in detail from a first principle’s approach. This includes life-of-mine workforce requirements, and equipment operating and maintenance costs.

		Mine dewatering will be an ongoing operating expense. During periods of extensive rainfall the lowest benches of any pit stage may flood. To ensure uninterrupted flow of ore to the mill, working faces will need to be maintained at higher elevations in a given pit, when possible. In general, as the mining stages progress, this is always possible until the end of the mine life.


Project No.: 147107		Page 19-8

Figure 19-2: Final Mining Limits


Project No.: 147107		Page 19-9

The pit waste rock is acid producing and is not suitable for road base material, road construction, or topping material. All such construction and road maintenance materials will be derived from the three limestone quarries. This material has been included in the calculation of the limestone requirements. The quarries will also supply limestone to the processing operation for neutralization purposes and the bulk construction material for the construction of the El Llagal tailings dam.

Costs were developed separately for the limestone quarried for processing and that to be used for construction of Llagal dam. The cost of limestone for processing was included in the operating costs for processing ore. These costs were also used for estimating the capital and sustaining costs of constructing the El Llagal tailings dam. The cost of hauling the quarry waste to the El Llagal dump was added to the mine waste cost.

Comment

The choice of equipment is appropriate for the deposit and mining schedule. The mining capital and operating costs have been developed in a thorough manner and appear to be reasonable.

19.2

Process Metal Recoveries

19.2.1		Overview

		The following process description is taken from the PDTS Feasibility study. These concepts have been extensively evaluated and are now fairly representative of the likely final design. A simplified block flowsheet of the current processing concept is provided as Figure 19-3.

		In situ sulphur content in the Pueblo Viejo ores varies widely, from 4% to 14%. Ore blending procedures in the mine will minimize this variation; it is planned to limit the sulphur grade of the pressure oxidation feed within the range of 6.67% S to 10% S. Ore delivery to the mill includes the option to stockpile run-of-mine ore at the primary crusher and blend the stockpiled material into the main ore stream as required to adjust the sulphur grade of the plant feed to a satisfactory specification. Some blending may occur in the fine ore stockpile, but in normal operation these facilities do not achieve much backmixing of the throughput. The autoclave feed slurry holding tanks are intended to supply additional short-term blending for the feed.


Project No.: 147107		Page 19-10

Figure 19-3: Simplified Flowsheet for Processing Pueblo Viejo Ore

The pressure oxidation circuit has been designed to treat 1,200 t/d of sulphur as a base case, which corresponds to a nominal ore throughput of 15,000 t/d based on the average sulphur content of 8% throughout the orebody. The crushing, grinding, and leaching circuits have been sized to treat 18,000 t/d of ore when processing the lowest anticipated sulphur content of the mill feed, namely 6.67% S. When the sulphur content in the ore increases to 10% S, ore throughput to the mill will be reduced to 12,000 t/d. An operating time of 93% has been assumed for the pressure oxidation circuit.

19.2.2		Process Description

		The crushing and grinding circuit uses high-pressure grinding rolls (HPGR) and ball mills, in preference to the more commonly seen semi-autogenous grinding technology in circuit design for new mills. The stated goal is to achieve lower operating and capital costs in the comminution circuit and improve circuit availability over what is achievable with SAG mills. While this selection should be feasible, it may not be optimum and may represent a higher technical risk to the project due to a lack of extensive experience with the technology in the


Project No.: 147107		Page 19-11

base and precious metals industries. However, it should be noted that HPGR technology is seeing increasing acceptance in the industry for crushing hard ores.

Primary Crushing

The primary crushing circuit is designed to crush 18,000 tonnes per calendar day based on a 65% operating time. The ore is delivered by 135 tonne haul trucks from the open pit and dumped directly into a 250 tonne dump pocket above the gyratory crusher. Plans include the capability to re-handle and blend in stockpiled ore at the primary crusher to provide a more constant feed sulphur grade to the process. The crusher discharge CSS setting is 125 mm. A hydraulic rock breaker is installed at the crusher dump pocket to break down any oversize rocks that are too large for the crusher feed opening. A belt feeder under the 200-t surge bin controls the rate of material fed to a conveyor that feeds two vibrating screens that have 50 mm square openings. The screen undersize fraction is conveyed to the 12,000 tonne live fine ore stockpile, while the screen oversize fraction is forwarded to a cone crusher MP800 for further size reduction.

Dust emissions at material transfer points are controlled by baghouses. Two belt scales are installed to measure the materials going into and leaving out of the 200 tonne surge bin.

As the ore is weakly acidic, about 5 kg/t limestone will be added to the crusher feed to minimize equipment corrosion. Over-addition of limestone is undesirable as carbon dioxide gas formed in the autoclave will waste oxygen and adversely affect solution chemistry in the autoclave. Alternatively, about 2 kg/t of lime can be substituted for the limestone to mitigate any problems in the autoclave.

Secondary Crushing

The secondary crushing circuit is designed to reduce 18,000 tonnes per calendar day to a minus 50 mm product based on a 65% operating time. The crusher feed is the plus 50 mm screen oversize fraction. One MP800 cone crusher will be installed to crush the plus 50 mm material to 100% passing 50 mm. This size reduction is achieved by setting the crusher CSS at 38 mm. The crushed product is conveyed directly to the 12,000 tonne live fine ore stockpile. The fine ore stockpile measures 60 m in diameter at the base and 22 m high when it is full.

Grinding

The grinding circuit has 98% availability and it is designed to grind 18,000 tonnes per calendar day based on a 93% operating time. A high operating time is necessary to provide a continuous feed to the autoclave circuit.


Project No.: 147107		Page 19-12

The grinding circuit consists of two HPGRs in parallel (2,400 kW each c/w VFD). Each HPGRs feeds in series two ball mills (7,000 kW each, c/w VFD) in series, two cyclopaks (6 of 838 mm diameter for the first ball mill, and 8 of 660 mm diameter for the second ball mill), two vibrating screens for the first ball discharge, and two vibrating trash screens. The transfer size from the HPGR to the first ball mill is 80% passing 5.5 mm at 18,000 t/d ore throughput, and the transfer size from the first ball mill to the second ball mill is 80% passing 275 ìm.

The fine ore stockpile is reclaimed by two (plus one standby) belt feeders and conveyed to a 200 tonne surge bin which is shared by two HPGRs. Dust emissions under the stockpile are controlled by a baghouse. A belt scale measures the mass of the fine crushed ore conveyed to the grinding circuit.

Each of the two belt feeders under the 200-t bin is capable of providing 80% of the total mill throughput. The HPGR feed size is 100% passing 50 mm and 80% passing 35 mm. The HPGR discharge is fed directly into the first ball mill. The first ball mill discharge is screened with the oversize (-25 ~ +5 mm) being recycled to the HPGRs, and the undersize fraction is pumped to the first cyclones. Fresh water is added to the first ball mill, the screens and to the pumpbox to adjust the cyclone feed pulp density. The first cyclone underflow returns to the first ball mill for further grinding. The first cyclone overflow at 30% pulp density goes to the second cyclone feed pumpbox. The second ball mill discharge is forwarded into the second cyclone. Both ball mills are operated in a closed circuit with the cyclones to produce a finished product grind size of 80% passing 80 to 90 ìm. A particle size monitor is installed to measure the second cyclone product size.

The second cyclone overflow, which becomes autoclave feed, is screened for trash removal. After the trash screens, the slurry flows by gravity to a 70 m thickener where it is thickened to 50% pulp density and pumped to four autoclave feed slurry holding tanks in the pressure oxidation area. Thickener overflow is recycled to the grinding circuit. Washdown and spills in the grinding area are collected in a trench served by floor sumps and pumped to the cyclone feed sumps.

Pressure Oxidation

The pressure oxidation circuit of the base case is designed to treat 1,200 tonnes of sulphur per calendar day based on a 93% operating time. There is a single stage of pre-heating to heat the autoclave feed slurry when it contains an insufficient amount of sulphur to reach the operating temperature autogenously. If autoclave feed slurry is pre-heated to a maximum 92°C and if there is 70% sulphide oxidation in the first compartment, the lowest permissible sulphur grade of the autoclave feed is 5.5%. It is projected that with the blending capacities designed into the circuit, the sulphur content of the autoclave feed should not fall below 6.67% S.


Project No.: 147107		Page 19-13

In order to reduce capital costs and maintenance costs, and to achieve 93% operating time, the autoclave will be made of titanium clad carbon steel. Sacrificial wear plates beneath the agitators may be provided to protect the shell from any erosion effects caused by the agitated slurry. It is more likely, however, that the internals will be well protected by the hematite scale formation. This was observed to be the case at Porgera, and scale formation was evident in the Pueblo Viejo pilot plant campaigns at SGS Lakefield Research.

Use of exposed titanium in the oxygen-rich environment will employ suitable control of the relative partial pressures of oxygen and steam in order to avoid a potential fire hazard.

To treat 1,200 tonnes of sulphur per calendar day with a minimum of 60 minutes retention time based on a 93% operating time, three autoclave trains will be installed. The autoclaves operate at 230°C and a gauge pressure of 500 psi (3,450 kPa). Each autoclave consists of six compartments, with the first compartment containing two agitators and having twice the volume of other compartments. Agitation is supplied by two 315 kW agitators in the first large compartment and one 150 kW agitator in each of other compartments. Cryogenic oxygen (99.5% O₂) is injected beneath the impellers via the nozzles located at the bottom of autoclave.

There are four autoclave feed slurry holding tanks ahead of the autoclaves. These tanks provide 10 hour surge capacity and also smooth the sulphur content of the autoclave feed. Three of these tanks are connected together and the slurry in these tanks flows by gravity to the fourth holding tank. When preheating is required, slurry in the fourth holding tank is pumped to a preheating tower where the cold slurry contacts the steam from the flash discharge tank. After being heated, the slurry is returned to the fourth holding tank. From the fourth holding tank, the slurry flows by gravity to three autoclave feed slurry tanks.

Ore slurry is fed to the autoclaves by high pressure diaphragm pumps. One feed pump is installed for each autoclave and there is a complete spare diaphragm pump that can replace any of the three installed pumps in the event of a catastrophic failure. Fresh water will be used as quench water to trim the excess heat generated from the oxidation of pyrite to maintain the operating temperature of 230°C. Oxygen is added relative to the amount of sulphur oxidized and the design criteria require the addition of two tonnes of oxygen for every tonne of sulphur in the feed. The oxygen plant has a capacity of 2,700 tonnes per operating day giving an oxygen to sulphur ratio of 2.09. Oxidized slurry is discharged from the autoclave through a choke control to a flash tank. Flashing of steam cools the slurry to 95 to 100°C. About 25% of water in the autoclave discharge evaporates to form steam during this step. Some of the flash steam will be used to heat the autoclave feed slurry when the low sulphur ore is processed. The flashed slurry flows by gravity to a stainless steel oxidized slurry tank, and is then pumped to the hot curing circuit.


Project No.: 147107		Page 19-14

When required, 2% of the 2,700 t/d of oxygen can be produced as liquid. The oxygen plant can capacity can be turned down by 30%, which will be applicable when one autoclave is down under scheduled maintenance.

Hot Curing

The oxidized slurry is maintained at an elevated temperature and ambient pressure for several hours tin four stainless steel tanks to dissolve the basic ferric sulphate that has formed in the autoclave. The hot curing step is necessary to achieve a CIL lime consumption of 10 kg/t or lower. The slurry suspension in the tanks during hot curing is maintained by 112 kW agitators. While the basic ferric sulphate dissolves during hot curing, free sulphuric acid is consumed and soluble ferric ion is produced. During hot curing about 15% of the solids will dissolve.

Oxidized Slurry Cooling

Hot cured slurry is pumped into a cooling tower. The hot slurry enters the cooling tower at the top and cold air is blown into the cooling tower from the bottom. The slurry will be cooled to below 40°C as a result of about 10% water evaporation from the slurry.

CCD Wash

The liquid phase of the oxidized slurry from the autoclave contains large amounts of free sulphuric acid and soluble metal sulphates. They are removed from the gold bearing solids before CIL leaching. The CCD washing circuit consists of three high capacity stainless steel thickeners 65 m in diameter. Over 99.0% of the free acid and soluble sulphate salts in the liquid phase are separated from the gold bearing solids at a wash ratio of 4.5 and thickener underflow pulp densities of 40%. The free acid and soluble sulphate salts are collected in No.1 thickener overflow and are then neutralized with limestone and lime in the HDS (high-density sludge) circuit. The washed solids that are discharged from No.3 thickener underflow become the CIL feed.

CIL Cyanide Leaching

The underflow from No.3 wash thickener is pumped to a vibrating trash screen sizing at 28 mesh (600 µm). The screen undersize is forwarded to a neutralization tank in the CIL cyanide leaching circuit. 10 kg of lime per tonne of CIL feed are added to the slurry to achieve pH 10.5. In the event of an upset in the CCD wash circuit leading to anomalously high acid concentrations in the CIL feed, there is provision to add limestone prior to lime to minimize reagent costs. The CIL circuit consists of six carbon steel tanks with a total of 18 h to 24 h retention time. Each tank is equipped with an agitator and two KEMIX interstage screens sizing at 28 mesh (600 µm). Cyanide dosage is 1.0 kg sodium cyanide


Project No.: 147107		Page 19-15

per tonne of ore feed, added to the first CIL tank. Air is sparged at a rate of 0.6 m³/h/m³ of slurry into the tanks to provide the oxidant for gold leaching.

Concentrations of activated carbon are maintained at 30 g/L in the first tank and 10 g/L in the subsequent tanks. A higher carbon concentration in the first tank will ensure a high gold recovery when treating the preg-robbing MO-BSD, MN-BSD and MN-VCL ore types. A relatively low final gold loading of 2,000 g/t on carbon will maximize gold recovery, when treating these feeds. Depending on the ore type, 88.7 to 94.0% of the gold in the ore is dissolved and loaded onto the activated carbon. Gold solution losses in the CIL tailing are below 0.01 ppm. Activated carbon is forwarded with recessed impeller pumps between CIL tanks, at a frequency of 45 minutes pumping every 6 hours. Total pumping time for one complete carbon forwarding for 6 tanks is 4.5 hours.

The loaded carbon from No.1 CIL tank is separated and washed on two vibrating screens, and after washing, is transferred to the acid wash vessels in the refinery circuit. At a maximum gold head grade of 7 g/t Au, 40 tonnes of activated carbon will be stripped in a 24 hour period. The CIL tailing is pumped to four safety screens with a 28 mesh (600 µm) aperture to capture any carbon particles that have escaped from the circuit due to operation upsets and/or holing of interstage screens. The captured carbon will be recycled to the CIL circuit.

About one third of the total 1.0 kg/t sodium cyanide addition that remains in the CIL tailing is not recovered. Before being discharged to the tailing pond, the cyanide in the CIL tailing is destroyed using Inco’s SO₂/air method. After cyanide destruction, the total cyanide concentration is less than 5 ppm.

Two 75 mm vertical slurry sump pumps are used for recovering any spillage in the CIL area. The recovered slurry is discharged to the CIL tanks.

Carbon Acid Washing

The loaded carbon is acid washed in four vessels. Each coned vessel is capable of treating 12 tonnes of carbon in a batch. Hydrochloric acid is used to remove scale from the carbon prior to stripping. Dilute hydrochloric acid is circulated from an FRP (fiberglass reinforced plastic) dilute acid holding tank through the acid wash vessel. After circulation, the spent hydrochloric acid is drained from the acid wash vessel and directed to the tailing discharge pumbox after being neutralized. After acid washing, the carbon is rinsed with fresh water, and then a dilute caustic soda solution is circulated through the carbon in the vessel to neutralize any residual hydrochloric acid. Spent caustic soda solution is drained and directed to the tailing discharge pumpbox.


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Carbon Stripping

Recessed impeller pumps transfer the acid washed carbon to one of four stripping vessels. Each vessel is capable of stripping 12 tonnes of carbon in a batch. A modified Zadra stripping process is used to strip the gold off the loaded carbon. The carbon stripping is operated at a temperature of 150°C for 9 hours. Barren and pregnant electrolytes are stored separately in two covered tanks.

The carbon stripping operation involves circulation of the barren electrolyte through one primary plate and frame heat exchanger, then one secondary shell and tube heat exchanger, and finally through the carbon stripping vessel. Barren electrolyte enters the primary plate and frame heat exchanger at 60°C and exits at 110°C. Low pressure steam is used in the secondary shell and tube heat exchanger to raise the barren electrolyte temperature to 150°C. The discharge solution, i.e., the pregnant electrolyte, from the stripping vessel, passes through the primary plate and frame heat exchanger, then the cooling shell and tube heat exchanger and finally to the pregnant electrolyte holding tank. Cooling water is used in the cooling shell and tube heat exchangers to cool the pregnant electrolyte to below 100°C. After stripping, the carbon is cooled by a water wash cycle, dewatered on a vibrating screen and then transferred to the carbon thermal regeneration area. The screen underflow flows by gravity to a strip vessel drain tank, and then is pumped to a screen to recover the fine carbon.

Carbon Thermal Regeneration

There are two carbon thermal regeneration circuits. Each circuit consists of one feed tank, one screw feeder, one electrically fired horizontal rotary reactivation kiln and one quench tank. The stripped carbon is dried and activated at a temperature of 700°C in the kiln. The kiln exhaust gases are vented to a common vent scrubber. The reactivated carbon is quenched in a quench tank. The quenched carbon from two regeneration kilns is advanced to a carbon attrition tank. After attrition, the carbon is pumped to a vibrating screen for separation and washing. The oversize fraction is returned to No.6 CIL tank to replace the forwarded carbon. The fine undersize carbon is forwarded to a settling pond. The settled fine carbon is filtered by a plate and frame filter press and is then periodically drummed and stored for sale.

A vertical sump pump is provided to recover any spillage and wash down.

Electrowinning

The pregnant electrolyte is pumped from the pregnant holding tank to four parallel electrowinning (EW) trains, each having three EW cells connected in series. Gold is plated on the woven wire stainless steel cathodes along with impurities such as Hg, Ag, Cu, etc.


Project No.: 147107		Page 19-17

The barren electrolyte that is discharged from the EW cell contains less than 2 ppm gold. It flows by gravity to a barren electrolyte collection tank, and is then pumped to a barren electrolyte surge tank for recycling to the strip circuit. Up to 15% of the barren electrolyte is bled for impurity control. The bled barren electrolyte is put to the CIL circuit. All EW cells are covered and ventilated. The EW cells are periodically removed from service and the cathodes are lifted out with a 1 tonne capacity overhead crane to harvest the gold. The cathodes are transferred to a cathode wash tank where the gold sludge is removed from the cathodes using high pressure water. The sludge is decanted and filtered through a plate and frame filter press.

A sump pump is provided to recover any spillage and washdown.

Refining

The cathode sludge will contain about 50% gold and 3.2% mercury. With an annual gold production of 700,000 oz, the daily mercury output is about 3.8 kg. The filtered gold sludge is loaded into boats, and then dried and heated to 1,000°C in an electrically dried mercury retort furnace. Mercury volatilized from the cathode sludge is condensed and recovered for sale.

The dried cathode sludge is then mixed with fluxes and smelted in three 500 kg capacity electric furnaces rated at 200 kW. The furnaces are hooded and ventilated to a dust collection system with the dust being collected and recycled. The furnace melt is poured to make gold doré. The slag is crushed manually and recycled to the grinding circuit.

HDS Circuit – Acid Neutralization

The overflow from No.1 CCD wash thickener is neutralized first with limestone and then with lime in a HDS (high density sludge) circuit. If ARD is available, it will be neutralized in the same circuit, at the same time. The HDS circuit consists of one stainless steel tank for mixing the recycled precipitate and limestone, six stainless steel tanks for limestone neutralization, two carbon steel tanks for lime neutralization and one 80 m high-rate thickener. Total limestone and lime dosage are 4.0 t at 100% CaCO₃ and 0.20 t at 100% CaO, respectively, for every tonne of sulphur in the solution. The limestone will be mined at site, with the high quality limestone being used to make lime in two vertical kilns. The limestone will be ground to 80% passing 60 µm for use in neutralization. All of the neutralization tanks are air-sparged to oxidize ferrous iron and other reducing species, and to remove the carbon dioxide gas.

In the recycle mixing tank, the recycled precipitate is mixed with limestone slurry. Retention time in this tank is about 2 minutes. The mixture then overflows to the first stainless steel limestone neutralization tank. Acid neutralization and metal precipitation will


Project No.: 147107		Page 19-18

progress with time as slurry pH rises slowly to about 6.4 in the sixth limestone neutralization tank. After a total of 3.3 hours retention time, 92% sulphate, 99.9% Al, 99.9% Cu, 100.0% Fe and 87% Zn in the original acidic solution will be neutralized and/or precipitated. In the remaining two lime neutralization tanks, the slurry pH will rise to 7.6 to 8.0. The final solution contains about 1.8 g/L SO₄, 0.10 ppm Al, 0.02 ppm Cu, 0.6 ppm Fe, 1.5 ppm Mn and 1.0 ppm Zn.

After neutralization with limestone and lime, the slurry is thickened in an 80 m high-rate thickener. A portion of the neutralized sludge at 40% solids is recycled to the neutralization circuit, with the balance being discharged into the tailing pond. The thickener overflow is acceptable for recycling to the CCD wash circuit. It is saturated with gypsum and contains elevated levels of chloride; it is therefore not suitable for recycling to the grinding circuit or the pressure oxidation circuit as quench water.

Cyanide Destruction

The SO₂/air method will be used to destroy the cyanide contained in the CIL tailing in a series of agitated tanks with a residence time of 90 minutes. Agitation power input is set at 0.45 kW/m³. A sulphur burner will be installed at site to produce SO₂ gas. In the feasibility study, partial cyanide destruction is selected to reduce the cyanide level down to 50 ppm CN. Nevertheless, the sulphur burner is sized to generate necessary SO₂ required for complete cyanide destruction. The residual cyanide level in the CIL tailing solution is typically 130 ppm CN. The SO₂/air method can readily reduce cyanide level to 2 ppm CN. After cyanide destruction, the slurry is pumped to the tailings pond together with the HDS sludge.

Tailing Disposal

A single cell tailing pond will be constructed to impound collectively the CIL tailing, HDS sludge and waste rock. The CIL tailing and HDS sludge are pumped to the tailings pond at 40% pulp density. The as-mined waste rock is truck dumped into the pond from beaches extending into the body of the pond. In the base case of 1,200 t/d sulphur processing rate, on average, 0.83 tonne of HDS sludge is produced for every tonne of CIL tailing. Since the waste rock should co-mingle well with the CIL/HDS tailing, sulphide oxidation from the waste rock should be minimal once under water.

Lime and Limestone

For processing 1,200 tonnes of sulphur per calendar day, the total limestone requirement is 7,400 tonnes per calendar day or 2.70 Mt/a. Limestone will be used directly in neutralization of waste rock in the tailing pond, site ARD and autoclave acid liquor. Some of the limestone may also be used in neutralization of CIL feed ahead of lime


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neutralization. Of the 7,400 t/d total, 1,300 t/d limestone will be used to make lime in two vertical kilns. Lime will be used in neutralization of CIL feed, cyanide destruction, tailing water treatment, millfeed and in the second stage neutralization of ARD and autoclave acidic liquor. The limestone deposits at Pueblo Viejo site are high quality typically containing 98% CaCO₃.

Primary Crushing of Limestone

The primary crushing circuit for limestone is designed to crush 7,400 tonnes of limestone per calendar day based on a 65% operating time for the base case of 1,200 tonnes of sulphur per calendar day. The ROM limestone is delivered by 135 t haul trucks from the quarry and dumped directly into a 250 tonne dump pocket above the gyratory crusher. The crusher discharge CSS setting of 100 mm produces a product size of P₈₀ passing 84 mm. A hydraulic rockbreaker is installed at the crusher dump pocket to break any oversize rocks too large for the crusher feed opening. A conveyor under the crusher transfers the crushed product to a 200 tonne surge bin at a nominal rate of 474 tonnes per operating hour. Dust emissions at material transfer points are controlled by baghouses. Two belt scales are installed to measure the materials going into and leaving the 200 tonne surge bin.

Secondary Crushing of Limestone

The secondary crushing circuit is designed to crush 6,100 tonnes of limestone per calendar day based on a 65% operating time for the base case of 1,200 tonnes of sulphur per calendar day. The limestone at the 200-t surge bin is reclaimed by one belt feeder, which forwards to a conveyor, and then to a vibrating double deck screen with a top deck of 100 mm and a bottom deck of 50 mm. The screen undersize fraction is conveyed to two screens which have square openings of 13 mm. The screen oversize fraction, (-100+50 mm), is forwarded to a splitter, from which sufficient limestone is conveyed to a stockpile that feeds two vertical lime kilns. The overall plant requirement for lime is estimated at 650 tonnes per calendar day which will require 1,300 tonnes of limestone being directed from the splitter to the stockpile. Any coarse limestone in excess of that required for burning to make lime will be conveyed to the secondary crushing circuit. A conveyor under the cone crusher transfers the crushed product to two vibrating screens with 13 mm openings. The screen undersize fraction with a P₈₀ of 9 mm is conveyed to a 6,000 t live capacity fine limestone stockpile at a rate of 391 tonnes per operating hour. The screen oversize fraction is recycled to the cone crusher which has a discharge CSS setting of 14 mm. The crusher product size P₈₀ is 13 mm. The screen product size, i.e. the ball mill feed size, has a P₈₀ of 9 mm. Dust emission is controlled by a baghouse which extracts dusts from the crusher discharge chute.


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Grinding of Limestone

The limestone grinding circuit is designed to grind 6,100 tonnes per calendar day to produce a product grind size of 80% passing 60 ìm. A high operating time is needed to provide a continuous supply of limestone slurry to the autoclave acidic liquor neutralisation circuit. Limestone at the fine limestone stockpile is reclaimed by two belt feeders and conveyed to two ball mills operating in parallel. A belt scale measures the mass of the fine crushed limestone conveyed to the grinding circuit. Each ball mill measures 4,110 mm x 6,707 mm, draws 2,000 kW and operates in a closed circuit with 8 mm to 375 mm cyclones to produce a finished product grind size of 80% passing 60 ìm or less. The cyclone underflow is returned to the ball mill for further grinding. The cyclone overflow at 40% pulp density flows by gravity to a limestone surge tank, and is then pumped to three limestone slurry storage tanks in the autoclave acid liquor neutralisation (HDS) area. A particle size monitor is installed to measure the cyclone product size. Washdown and spills in the grinding area are collected in a trench served by floor sumps and pumped to the ball mills.

Lime Plant

Based on the 1,200 t/d sulphur processing rate and 18,000 t/d ore processing throughput, the lime requirements for autoclave acidic liquor neutralisation and CIL cyanide leaching are 240 t/d and 180 t/d, respectively. There is an additional 200 t/d lime requirement to neutralize the ARD. About 10 t/d lime is needed in the cyanide destruction circuit, and a small amount of lime will also be used in tailing water treatment. Thus, the total lime requirement is around 640 tonnes per calendar day or 690 tonnes per operating day based on a 93% operating time.

There are two large high-quality limestone deposits conveniently located on the Pueblo Viejo site. Because about 80% of the quarried limestone is used as ground slurry to neutralize the autoclave acidic liquor and ARD and only 20% of the quarried limestone is burnt to produce lime, a narrow size range of the crushed limestone can be selected to optimize the kiln performance. Limestone sized at -100+50 mm has been selected to feed the Maerz vertical kilns.

Two 350 t vertical twin-shaft PFR (parallel flow regenerative) lime kilns will be installed. The vertical kiln has been selected due to 30% fuel savings compared to the horizontal kiln. The kilns will be fired with No. 2 diesel fuel; -100+50 mm limestone is loaded intermittently to the top of the kiln. The outgoing hot gas is used to heat the incoming limestone feed. Inside the kiln, the limestone is heated and burnt up to a maximum temperature of 1,050°C. At high temperature, limestone decomposes to form lime and carbon dioxide gas. Total retention time inside the kiln is about 12 hours.


Project No.: 147107		Page 19-21

Particle loading in the off-gas is between 5 and 10 g/m³ before the baghouse, which is equivalent to 2.5% of the daily production, and <20 mg/m³ after the baghouse. The dusts collected in the baghouse are a mixture of limestone and lime, and are useful for the first lime neutralisation tank in the HDS circuit.

Cold air is blown to the bottom of the kiln to cool the burnt lime product. The cooled lime product, which is discharged intermittently from the bottom of the kiln, is conveyed to a roll crusher and then transferred to a crushed burnt lime silo by a conveyor. The burnt lime is fed from the silo via a screw feeder directly to a 2,438 x 4,572 mm ball mill where slaking will be carried out. The minimum slaking time will be 6 minutes. After slaking, the lime slurry is pumped from a pumpbox to six 200 mm cyclones with the underflow being recycled to the ball mill and the overflow flowing by gravity to a lime slurry surge tank. From this surge tank, the lime slurry is pumped to two lime slurry storage tanks in the CIL circuit and HDS circuit.

Recoverability

The extent of testing and pilot plant operation that supports the Feasibility Study is appropriate for this level of project evaluation. The extent of investigation into metallurgical variability and the level of testing undertaken give AMEC a high level of confidence that the predicted recoveries for the various ore types will be very close to what would be experienced if the selected flowsheet was constructed.

19.3		Markets

		Gold is traded on world markets, with benchmark prices generally based on the London gold market quotations. Gold bullion is held principally as a safeguard against the collapse of paper assets such as stocks, bonds, and other financial instruments that are receivable in fiat currencies. Central banks and other official institutions hold approximately 31,000 tonnes of gold, representing almost 22% of the above-ground stocks. Due to the size of the international bullion market and above-ground stocks, individual gold producers or other market participants generally do not influence pricing or total quantities offered and sold. Gold is primarily used for jewellery and in industrial applications, which account for approximately 76% of the annual demand. Other uses include coin and bar fabrication, dentistry, and decorative applications.

19.4		Contracts

		Commencement of construction is subject to Placer Dome issuing a Project Notice to the Dominican Republic and this is conditional on receipt of permits for the mine and power plant and the Dominican Republic fulfilling all of its contractual obligations to the project.


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After giving the Dominican State a Project Notice, PDDC shall assume the responsibilities for all historical environmental matters within the boundaries of the Development Areas designated in the feasibility study; provided that the Dominican state will remain liable for all offsite contamination and public health claims resulting from Historical Environmental Matters.

The Dominican State will relocate at its cost those persons living in the Los Cacaos section of the Sites and pay all indemnification and other compensation to such persons in connection with the relocation.

The Dominican State shall also acquire and lease to PDDC at no additional cost all land and mineral rights deemed necessary by PDDC for the permanent disposal of tailings and waste and shall, in compliance with the applicable Environmental and Social Guidelines and Policies, relocate those persons living on land required for the disposal of tailings. The Dominican State is also responsible for termination of employment and satisfaction of any labour claims of ex-Rosario employees.

The Dominican State shall expropriate lands deemed necessary by PDDC for operation of the project and PDDC shall not be obligated to pay additional consideration to the Dominican State with respect to any such lands within the boundaries of the Fiscal Reserve or within the Los Cacaos drainage. The Dominican State shall provide from a permanent reliable source all water rights necessary to conduct the operations in accordance with the Feasibility Study, subject to all laws and required permits.

19.5		Environmental

		The Dominican Republic Environmental Law No. 64/2000 establishes the process of obtaining environmental permits. There are a number of other technical permits that will have to be assessed by corresponding agencies.

		PDDC’s legal review indicates that there are 29 permits required to build and operate the Pueblo Viejo project. These are of three types: those needed before the environmental license, those included in the environmental license and those needed after receipt of the environmental license.

		Permits obtained to date include:

	•		Presentation of “Analisis Previo” Form.

	•		Definition and acceptance of Terms of Reference for the Environmental Impact Assessment, which was developed in conformance with local, PDI, and World Bank Guidelines.


Project No.: 147107		Page 19-23

	•		Soil use permit for mining reserve and waterline corridor.

	•		Environmental authorization for soil use, mining reserve.

The EIA is anticipated to be submitted to the Secretaria de Estado de Medio Ambiente y Recursos Naturales (SEMARN) during the fourth quarter of 2005.

Post-closure costs have been planned for, and additional post-closure costs will be covered by funds established during mine operation

19.6		Taxes

		PDDC has the following tax obligations under the Special Lease Agreement (SLA) with the Dominican state:

	•		25% Income tax

	•		5% Withholding tax of loan interest

	•		Immunity of taxes during construction of the mine.

Imports into the Dominican Republic by PDDC will be immune from all taxes and duties over the term of the SLA.

19.7		Capital Cost Estimates

		The capital cost estimate of US$1.348 billion for the Pueblo Viejo project includes direct costs for general site preparation, the open pit mine, the processing plant, ancillary facilities, tailings disposal, site access road and offsite power generation and high voltage transmission line. Also included are indirect costs for design and engineering, project management, construction overheads, PDDC operations overheads, warehouse inventory and freight. A 10% contingency has been added to the above costs and an allowance for initial operating expenses has been added “below the line.” All costs are stated in 2^nd quarter 2005 U.S. dollars.

		A summary of capital costs by major items is provided in Table 19-4.


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Table 19-4: Capital Costs by Major Items


	Cost US$
Area	(M)

Site preparation and access road		30.6
Process plant incl oxygen and lime plants		426.5
Open pit mine		54.3
Tailings disposal		30.6
Ancillary facilities and utilities		75.0
Engineering contingency on above		25.6

Subtotal direct costs		642.6

Construction overheads		57.7
Operations overheads		29.4
Project management		68.4
Design and engineering		44.9
Warehouse inventory		20.6
Freight		29.5

Subtotal indirect costs		250.5

Power supply and powerline (incl both direct and indirect costs)		329.1

Total construction cost		1,222.2

Contingency		121.8
Cost including contingency		1,344
Initial operating expense		3.8
Total estimated cost		1,347.8

AMEC did not review power plant and transmission line costs prepared for PDTS by SNC Lavalin because this is outside the expertise of AMEC cost estimators. AMEC assumes that the estimates were prepared by a Qualified Person. Otherwise, AMEC believes that the estimated capital costs are reasonable and complete and that they have been prepared in accordance with industry standards at levels of detail commensurate with a feasibility study.

19.7.1		Mill Operating Cost

		The mill operating cost is presented as a formula that permits the calculation of the operating cost per tonne as a function of power cost and ore sulphur content. It has been developed for a milling operation treating 1,200 t/d of sulphur. It is important to note that costs related to cyanide destruction, tailings water treatment, fresh water pumping to the mill, and ARD neutralization reagents are excluded. Under the assumed unit power cost of 0.03 USS/kW-h in the base case, with an 8% sulphur ore feed, power accounts for 30% of the mill operating cost, labour for 12%, and consumables, including mill liners, mill media, and chemicals, for 41%. Consumables therefore represent a large component of the costs.


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In developing the formula indicated above, a similar algorithm was created for calculation of the operating cost to produce ground limestone slurry (with no dependence on the ore sulphur content), and for the production of burned lime.

The build-up of these formulas is described in the feasibility study. AMEC has reviewed these calculations and is in agreement that the derivation methodology is appropriate for a feasibility level of evaluation.

It is important to note that the capital and operating costs are based on an aggressive oxygen utilization assumption (2 tonnes of oxygen per tonne of sulphur). For pyrite oxidation, stoichiometry requires 1.875 tonnes oxygen per tonne of sulphur, and so the assumed utilization would be 94%. AMEC would expect oxygen utilization to be 80% to 85% based on knowledge of existing pressure oxidation operations. Placer’s assumption is believed to be technically feasible, and Placer cites evidence of operating at this oxygen consumption level at one of its other commercial circuits, when conditions are carefully controlled. Importantly, the absence of carbonate in Pueblo Viejo ore means that there will be only a small amount of carbon dioxide gas formation in the autoclave, which will result in less need to vent from the autoclave gas phase with a commensurate loss in oxygen. This results in the availability of oxygen for sulphur oxidation being better than what would be expected from carbonate rich ores. The ability to achieve high utilization depends on selection of a source of high purity (cryogenic) oxygen. However, the departure from industry experience with this assumption needs to be highlighted, albeit the lack of carbonate in the autoclave feed as noted does enhance oxygen utilization with a commensurate impact on capital and operating costs.

19.8		Economic Analysis

		The Economic analysis of the Pueblo Viejo project is based upon discounted cash flow analysis on a post-local tax basis. Projections for annual revenues and costs are based on data developed in the July 2005 Feasibility Study, covering the mine, process plant, power plant and transmission line, capital expenditures and operating costs.

		Metal prices for the base case (US$400/oz Au and US$6.00/oz Ag) provided by Placer Dome conform to the prices used to compute reserves. Based on the assumptions used, discounted cash flow analysis indicates that the project offers a positive return at a discount rate of 5%, with a payback period of less than half the mine life.

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


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As with many projects of this type, the Pueblo Viejo project is most sensitive to changes in metal price and rather less so to changes in operating cost and capital expenditure. The project assumptions used for the sale of excess power above that required for the project and the sale of the power station at the end of the mine life significantly impact the NPV and IRR of the Pueblo Viejo project. PDTS used long-term prices for energy and short-term prices for grinding media, cyanide, reagents, and explosives.

The mine schedule shows a declining grade of material being processed over the mine life. While this may allow the discounted cash flow to be maximized it also causes the operation to become less profitable as time passes. This condition, if unexpectedly combined with any significant drop in gold price, could result in the profitability being impaired.

In AMEC’s opinion, the level of detail used in the economic analysis is appropriate for a feasibility study.


Project No.: 147107		Page 19-27

20.0

CONCLUSIONS

20.1		Introduction

		Mineral Resources and Mineral Reserves for the Pueblo Viejo project are supported with the results of a Feasibility Study completed in July 2005, which followed prefeasibility work in 2003 and 2004. The Feasibility Study provides substantially comprehensive details regarding the geology, exploration data, environmental studies, metallurgy, mine designs, production plans, process and infrastructure designs, power supply, capital and operating cost estimates and economic analyses.

2.2		Conclusions

		AMEC reviewed pertinent geological, resource estimation, mining, metallurgical, and capital cost information for the Pueblo Viejo project to assess the mineral resource and mineral reserves announced by Placer Dome on 27 September 2005. The following is a list of general conclusions reached by AMEC as a result of its reviews.

	1.		The geology of the Moore and Monte Negro deposits is well understood. Lithological, alteration and structural controls of mineralization are well defined and are used in domaining for grade estimation, metallurgy and mine designs.

	2.		Resource estimates are supported with results from 573 RC and core drill holes (349 drill holes for Moore deposit, 224 drill holes for Monte Negro deposit) totalling 80,261 meters. Drilling was carried out in multiple campaigns by Rosario, GENEL, MIM and PDDC.

	3.		Verification of historical data regarding drill hole surveys, assays, and drill hole logs is problematic in some cases due to missing records. AMEC was able to directly verify the assay and geology databases relative to original records for many of the Rosario drill holes and all of the GENEL and PDDC drill holes. Validation of results for MIM and other Rosario drill holes had to be achieved indirectly through comparison of PDDC and GENEL holes to Rosario and MIM holes by nearest-pair analysis and inspection of cross sections.

	4.		Sampling methods for each drilling campaign were generally acceptable to the extent that they could be documented. Sample preparation and assay protocols are documented only for GENEL and PDDC drilling campaigns and these were generally acceptable. Sample preparation protocols used by PDDC could be improved in future programs.

	5.		Assay quality assurance and quality control protocols and results vary significantly depending on the drilling campaign. Some of Rosario assays are supported by check assays. MIM drilling had not QA/QC. GENEL and PDDC assay programs used


Project No.: 147107		Page 20-1

			QA/QC to assess assay precision and accuracy. PDDC QA/QC protocols should be improved in the future. Results show generally acceptable precision and accuracy in the Rosario drilling with check assays, with all GENEL drill holes and with all PDDC drill holes. PDDC assay precision could be improved with better sample preparation protocols. Drill holes without QA/QC support were indirectly validated with nearest pair twin analyses and inspection of cross sections using PDDC and GENEL drill holes that were assayed with QA/QC.

	6.		Resource models were developed using standard industry methods. Block dimensions are relatively small at 10 m x 10 m x 10 m. Grade interpolation was controlled by use of multiple domains based on lithology, alteration, and structural setting, and these were well designed to separate different grade populations. Estimates use an appropriate compositing strategy and high-grades are properly cut. AMEC validated resource estimates and found that the estimates are globally unbiased.

	7.		Mine designs are based on appropriate criteria including cutoff grades, operating costs, gold price, metallurgical recovery, and geotechnical data.

	8.		The cutoff grade strategy is aggressive and uses higher cutoffs in the early years of mining to increase cash flows. Reasonable justification for the cutoff strategy is provided. Production rates are designed for 40% to 50% of the ore (higher grade) to be processed at the time of mining and 50% to 60% of the remaining ore (lower grade) to be stockpiled for later processing. AMEC has reviewed this schedule and found it to be reasonable.

	9.		Maintenance of cash flows through the project life requires diligence in ore blending and maintenance of the projected cutoff grades. PDDC has demonstrated that this is possible, but the requirements add risk to the project.

	10.		The economic analysis of the project is based upon discounted cash flow analysis on a post-local tax basis. The project has a positive rate of return at a 5% discount rate, producing a payback of less than half the mine life.

	11.		Process designs of whole ore pressure oxidation followed by CIL cyanide leach are supported with appropriate metallurgical testwork and plant designs.

	12.		AMEC reviewed capital and operating cost estimates and found them to include all appropriate cost items, and to be reasonable and based on acceptable estimation methods.

	13.		PDDC’s assessment of environmental risks, liabilities, and costs associated with construction and operation of the Pueblo Viejo mine has been conducted thoroughly. Post-closure costs have been planned for and additional post-closure costs will be covered by funds established during mining operations.

	14.		Mineral Resources are classified as Measured, Indicated, and Inferred in accordance with CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines


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(2000) and therefore comply with NI 43-101. Proven and Probable Mineral Reserves are developed using Measured and Indicated Resources and appropriate mine designs. AMEC accepts the Mineral Resource and Mineral Reserve statements.


Project No.: 147107		Page 20-3

21.0		REFERENCES

		Canadian Institute of Mining, Metallurgy and Petroleum, 2000, CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines, August 20, 2000.

		Keech, C.G., 2004, Report on Check Assay Data for the Pueblo Viejo Project, Dominican Republic; 21 April 2004 Internal Placer Dome Report, 86 p.

		Keech, C.G. and O’Brien, D. L., 2004, Pueblo Viejo Interim Status Report Volume 1 – Geology; September 2004, Placer Dome Internal Report prepared by Placer Dome Technical Services Group, 469 p.

		Lockhart, E. and Bowen, B., 1997, Report on Geological Modelling and Resource Estimation of the Moore and Monte Negro Deposits of Rosario Dominicana Pueblo Viejo Mine, Dominican Republic; 22 July 1997 Consultant’s report to GENEL Dominicana, 37 p. plus appendices.

		Pitard, F.F., 1984, the Sampling Characteristics of Gold and Silver in Rosario Dominicana Sulphide Ore; 14 September 1984, Consultant’s Report to AMAX Extractive Research & Development, Inc 27 p.

		BGC Engineering Inc. 2004. Pueblo Viejo Project Memorandum. Pre-feasibility Earthworks Cost Assumptions Update, December 2004.

		Environmental Geochemical International. 2003. Peer Review of ARD Management Program. Report prepared for Placer Dome Inc., July 2003.

		Kesler, S.E., Campbell, I.H., Smith, C.N., Hall, C.M., and Allen, C.M., 2005, Age of the Pueblo Viejo Gold-Silver Deposit and Its Significance to Models for High-Sulphidation Epithermal Mineralization: Economic Geology, v. 100, p. 253-272.

		Morjwijk Enterprises Limited. 2003. Pueblo Viejo Project: Baseline Report on Acid Rock Drainage and Metal Leaching. Report prepared for Placer Dome Inc., November 2003.

		Morjwijk Enterprises Limited. 2004. Supplementary Geochemical Information and Predictions for Pre-feasibility Studies. Report prepared for Placer Dome Inc., September 2004.

		Muntean, J.L., Kesler, S.E., Russell, N., and Polanco, J., 1990, Evolution of the Monte Negro Acid Sulphate Au-Ag Deposit, Pueblo Viejo, Dominican Republic: Important Factors in Grade Development: Economic Geology, v. 85, p. 1738-1785.


Project No.: 147107		Page 21-1

Nelson, C.E., 2000, Volcanic Domes and Gold Mineralization at the Pueblo Viejo District, Dominican Republic: Mineralium Deposita, v. 35, p. 511-525.

Norman, Tony, 2002, “Pueblo Viejo Pit Mapping,” Memo Prepared for Placer Dome Inc, December 16, 2002.

Panteleyev, A., 1996, Epithermal Au-Ag-Cu: High Sulphidation; in, Selected British Columbia Mineral Deposit Profiles, Volume 2, Lefebure, D.V., and Höy, T., editors, British Columbia Ministry of Energy, Mines and Petroleum Resources, p. 37-39.

Placer Dome Technical Services Limited. 2004. Pueblo Viejo Project Pre-Feasibility Report. Volume 4 – General Matters. Report produced for Placer Dome Dominicana Corporation, December 2004.

Sillitoe, R.H., and Bonham, H.F. Jr., 1984, Volcanic Landforms and Ore Deposits: Economic Geology, v. 79, p. 1286-1298.


Project No.: 147107		Page 21-2

The undersigned prepared this Technical Report, titled Technical Report and Qualified Persons Review, Pueblo Viejo Project, Dominican Republic, dated October 26, 2005, in support of the public disclosure of Mineral Reserve and Mineral Resource estimates for the Pueblo Viejo Project as of 27 September 2005. The format and content of the report are intended to conform to Form 43-101F1 of National Instrument 43-101 (NI 43-101) of the Canadian Securities Administrators.

Signed


/s/ Susan Lomas, P.Geo.
Susan Lomas, P.Geo.		October 26, 2005

/s/ Tracy E. Barnes
Tracy E. Barnes		October 26, 2005

/s/ Lynton S. Gormely
Lynton S. Gormely		October 26, 2005

/s/ Ted Eggleston
Ted Eggleston		October 26, 2005

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