Pueblo Viejo Gold Technical Report

GOLDCORP INC
Pueblo Viejo Technical Report

EXECUTIVE SUMMARY

This Technical Report on the Pueblo Viejo Gold Project (PVGP) in the Dominican Republic has been prepared in accordance with the requirements of National Instrument 43-101 (NI 43-101), “Standards of Disclosure for Mineral Project”, by Qualified Persons Mr H A Smith and Mr P R Stephenson of AMC Mining Consultants (Canada) Ltd (AMC) of Vancouver, Canada, Mr M G Butcher of Goldcorp Inc. (Goldcorp) of Vancouver, Canada, and Mr C A Carr of Rescan Environmental Services Ltd. (Rescan) of Victoria, Canada, on behalf of Goldcorp. Mr Smith, Mr Stephenson and Mr Carr visited the PVGP site in March 2008 where they examined all relevant surface features, infrastructure and core / sample facilities, reviewed representative plans and sections and held discussions with key personnel.

Unless otherwise stated, costs are in US dollars and measurements are metric.

Goldcorp owns 40% of the PVGP, the other 60% being owned by Barrick Gold Corporation (Barrick) which is also the operator. On February 21, 2008, Goldcorp received the results of an update to an existing 2005 Placer Dome Feasibility Study (PDFS) on the PVGP prepared by Barrick (the Barrick 2007 Feasibility Study or Barrick FS). On or before March 31, 2008, Goldcorp was required to file its Annual Information Form (AIF) and Goldcorp believed that in order to provide up-to-date, full, true and plain discourse, it was necessary that the information contained in the Barrick FS form the basis of the scientific and technical information on the PVGP contained in the AIF. As the Barrick FS information was new material scientific or technical information, filing the AIF containing this information triggered a requirement to file a technical report to support such information not later than the time the AIF is filed. Since the time frame between the receipt of the Barrick FS and the deadline for filing the AIF was short, Goldcorp applied for, and was granted, exemptive relief from the appropriate securities regulatory authorities from the requirement that it file a technical report for the Pueblo Viejo Project not later than filing its AIF, provided that:

	1.		This annual information form includes the following cautionary language:

	“The technical disclosure, including the Mineral Reserve and Mineral Resource estimates, in this annual information form with respect to the Pueblo Viejo Project has not been supported by a technical report prepared in accordance with NI 43-101. A technical report is being prepared by qualified persons under NI 43-101 and it will be available for review on the SEDAR website located at www.sedar.com under the Corporation’s profile on or before May 15, 2008. Readers are advised to refer to that technical report when it is filed.” and

	2.		Goldcorp files the technical report as soon as practicable but, in any event, not later than May 15, 2008.

The mineral resource and mineral reserve estimates contained in the Barrick FS were as at end-June 2007 and the Barrick FS report was effective as at third quarter 2007. Further exploration and technical / economic studies have been undertaken since June 2007 and updated mineral resource and mineral reserve estimates, effective end-December 2007, were released by Barrick in early 2008. These later mineral resource and mineral reserve estimates are not the subject of this report.



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Pueblo Viejo is located in the central part of the Dominican Republic on the Caribbean island of Hispaniola, approximately 100 km northwest of the national capital of Santo Domingo. The Pueblo Viejo property, referred to as the Montenegro Fiscal Reserve (MFR), covers an area of 4,880 hectares. It encompasses areas previously held by Rosario Dominicana S.A. (Rosario), which mined oxide and limited transitional material in the Moore, Monte Negro, Mejita, and Cumba deposits by open cut methods between 1972 and 1998.

During 2000, the Dominican State invited international bids for the leasing and mineral exploitation of the Pueblo Viejo sulphide deposits. Placer Dome Dominicana Corporation (PDDC), a subsidiary of Placer Dome Inc. (Placer) won the bid, and a Letter of Intent was signed in August 2001, pursuant to which the parties negotiated a Special Lease Agreement (SLA) for the MFR. The SLA was subsequently ratified by the Dominican National Congress, was published in the Official Gazette of the Dominican Republic in May 2003 and became effective on July 29, 2003. Through the SLA, the Dominican State granted to PDDC an option to lease the MFR free and clear of all defects, claims or encumbrances, for the term of the lease, for exploitation of the minerals contained in the MFR under the terms, conditions, stipulations, and agreements set forth in the SLA.

In February 2006, Barrick acquired control of Placer and at the same time, sold a 40% stake in the Placer subsidiary that owned PDDC to Goldcorp. In December 2006, PDDC was renamed Pueblo Viejo Dominicana Corporation (PVDC) and the change of name was officially registered with the Government of the Dominican Republic. For convenience, in this report Barrick is identified as the PVGP operating company, although under the terms of the agreement between Barrick and Goldcorp, Barrick has designated a separate Barrick subsidiary as the “Operator” of the PVGP.

In 24 years of production, the Pueblo Viejo mine produced a total of 5.5 M oz of gold and 25.2 M oz of silver.

The Pueblo Viejo property comprises several high sulphidation epithermal deposits of which Moore and Monte Negro are the largest. The deposits form funnel shaped envelopes of advanced argillic alteration where hydrothermal fluids migrated upwards and laterally along permeable horizons. Mineralization is predominantly pyrite as disseminations, layers, replacements, and veins with lesser amounts of sphalerite and enargite. Past mining operations have stripped the deposit areas of any surface oxidation and the oxide mineralization is now depleted.

Gold is intimately associated with pyrite veins, disseminations, replacements, and layers within the zones of advanced argillic alteration, occurring as sub-microscopic (less than 0.5 µm) inclusions and in solid solution within the crystal structure of the pyrite. It is present as native gold, sylvanite (AuAgTe₄), and aurostibnite (AuSb₂). Gold values are generally 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.

Subsequent to the main phase of the Rosario mining operation in the late 1990s, several companies conducted exploration over the deposits prior to Placer winning the Dominican State bid in 2000. Placer conducted extensive drilling and sampling programs in 2002 and 2004, including technical / economic studies and mineral resource / reserve estimation. This



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work was compiled into a feasibility study completed in July 2005. After gaining control of Placer, Barrick conducted its own extensive exploration, technical and economic programs, culminating in the completion of an updated feasibility study (the Barrick FS), which information, as indicated above, was made available to Goldcorp on February 21, 2008.

The mineral resource and mineral reserve estimates published as part of the Barrick FS are as follows:

Mineral Resource Estimates, 1.4 g/t Au Cut-off Grade
(Effective Date of Mineral Resource Estimate June 30, 2007)


		Tonnes		Au		Au		Ag		Ag		Cu		Cu		S		Zn		Zn
		(M)		(g/t)		(Moz)		(g/t)		(Moz)		(%)		(Mlb)		(%)		(%)		(Mlb)
	Measured		4.6		3.3		0.5		16.9		2.5		0.07		6.7		7.5		0.63		64.5
Monte Negro	Indicated		80.9		2.9		7.5		13.8		35.9		0.06		99.9		7.5		0.50		888.2
	Total		85.5		2.9		8.0		14.0		38.4		0.06		106.6		7.5		0.51		952.7
	Measured		7.9		3.3		0.8		18.3		4.6		0.11		19.5		8.2		0.86		150.1
Moore	Indicated		155.2		2.8		13.9		12.9		64.2		0.09		301.1		7.8		0.58		1,984.3
	Total		163.1		2.8		14.8		13.1		68.9		0.09		320.5		7.8		0.59		2,134.4
	Measured		12.5		3.3		1.3		17.7		7.1		0.10		26.2		7.9		0.78		214.6
Combined	Indicated		236.1		2.8		21.4		13.2		100.1		0.08		400.9		7.7		0.55		2,872.5
	Total		248.6		2.8		22.7		13.4		107.2		0.08		427.1		7.7		0.56		3,087.1
Goldcorp Share (40%)			99.4		2.8		9.1		13.4		42.9		0.08		170.8		7.7		0.56		1,234.8

	Inferred		81.4		2.5		6.5		3.4		9.0		0.02		40.0		7.7		0.02		33.4
Goldcorp Share (40%)			32.6		2.5		2.6		3.4		3.6		0.02		16.0		7.7		0.02		13.4

Mineral Reserve Estimates, Variable Cut-off Value
(Effective Date of Mineral Reserve Estimate June 30, 2007)


		Tonnes		Au		Au		Ag		Ag		Cu		Cu		S
		(M)		(g/t)		(Moz)		(g/t)		(Moz)		(%)		(Mlb)		(%)
	Proven		4.3		3.3		0.5		17.9		2.5		0.07		7.0		7.0
Monte Negro	Probable		66.3		2.9		6.3		15.6		33.3		0.06		86.0		6.8
	Total		70.6		3.0		6.7		15.8		35.8		0.06		93.0		6.8
	Proven		6.9		3.4		0.8		19.6		4.4		0.12		18.0		7.7
Moore	Probable		128.2		2.9		11.9		14.0		57.8		0.10		277.0		7.4
	Total		135.1		2.9		12.7		14.3		62.1		0.10		295.0		7.4
Rosario Stockpiles	Probable		8.4		2.3		0.6										5.4
	Proven		11.2		3.4		1.2		18.9		6.8		0.10		25.0		7.5
Combined	Probable		202.9		2.9		18.8		14.6		91.0		0.09		363.0		7.1
	Total		214.1		2.9		20.0		14.8		97.9		0.09		388.0		7.1
Goldcorp Share (40%)			85.6		2.9		8.0		14.8		39.2		0.09		155.2		7.1

Resources inclusive of resources converted to reserves

Metal prices used for resources: gold $650.00/oz, silver $11.50/oz, copper $2.25/lb

Metal prices used for reserves: gold $575.00/oz, silver $10.75/oz, copper $2.00/lb

AMC reviewed all inputs into the mineral resource and mineral reserve estimates, including the geological model, nature, quality (QA / QC) and distribution of exploration data, statistical and geostatistical studies, resource modelling and classification procedures, mine plan, processing parameters, environmental and social aspects, legal and governmental aspects, capital and operating costs, mineral reserve estimation and classification procedures and project financial analysis. A brief description of each major item and AMC’s conclusions and recommendations are presented below.



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The Barrick FS resource estimate is based on 138,349m of drilling in 1,814 holes (diamond, rotary and reverse circulation), on a spacing averaging 60-80m in the main parts of the deposits. The model was prepared in mid-2007, although the cut-off date for gold assays was January 2007. Extensive validation and verification programs were implemented by Placer, Barrick and other companies and, in AMC’s opinion, the resource database is free from major defects and is of an acceptable quality to support a feasibility study. Any remaining deficiencies are unlikely to materially affect global resource estimates, but may impact in places on local estimates. AMC recommends that attention continue to be paid to the quality of historic drilling information, with targeted replacement drilling being undertaken where necessary.

The geological model underpinning the resource estimate assumes that higher grade mineralization is controlled by north to north-west striking, steep-dipping feeder zones which flatten out near surface. In the absence of geological solids to define the feeders, probability indicators at cut-off grades of 5 g/t gold and 1 g/t gold were used to delineate higher and lower grade mineralization respectively. After top-cutting, gold assays were composited to 10m, statistical and geostatistical studies undertaken and variography derived, and gold grades were interpolated into 10m by 10m by 10m sized blocks using inverse distance cubed (ID³). Sulphur grades were interpolated using inverse distance squared (ID²). Resources were classified based mainly on distance between blocks and composites, with Measured Resources being restricted to blocks intersected by an assayed drill hole.

In AMC’s opinion, most of the components of the resource estimate are of a good standard, but two give cause for concern.

	1.		The use of ID³ for grade interpolation is relatively unusual in feasibility study resource estimates for gold deposits. It minimises the degree of grade smoothing, thus tending to maintain the variability of grades as reflected by composite samples, but potentially results in conditional bias — a bias that depends on the cut-off grade applied. The tendency is to over-estimate high grades and under-estimate low grades. This may not be a material issue when the deposit is planned to be mined at around its average grade, as the conditional biases may approximately balance out. However, Pueblo Viejo will be mined at a higher than average grade for the early years, with lower grade material being stockpiled for treatment in the later years. In this situation, a conditional bias can be a material matter.

			In order to investigate this possibility, AMC re-estimated gold grades for the deposit using ordinary kriging (OK), a technique that imposes a degree of grade smoothing and that should, ideally, result in an unbiased estimate. AMC then compared the resource estimates for the ID³ model and the OK model for the first two years of the direct mill feed component of planned mine production, the years of highest gold production. The OK estimate resulted in an average gold grade and contained gold around 10% lower than for the ID³ estimate. In AMC’s opinion, this confirms that the use of ID³ may be a material issue in the early mining years when revenue from gold production has a significant effect on the Net Present Value (NPV) of the project.

			Since AMC’s estimate was a check review undertaken with limited time, the results should be confirmed before acting on the findings. AMC recommends that more detailed investigations be undertaken to assess the validity of AMC’s conclusions. If they are shown to be valid, it may be advisable to undertake infill drilling in selected



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			parts of the Moore and Monte Negro deposits so that the issue of the gold grade interpolation method can be more thoroughly evaluated.

	2.		In AMC’s view, the classification of Measured Resources applied in the Barrick FS, which takes no account of continuity of mineralisation between drill holes, is not logical, is inconsistent with the geology of the deposit and has resulted in a substantial under-statement of Measured Resources (and therefore of Proved Reserves). AMC recommends that the approach to Measured Resource classification be reviewed.

The mine plan at Pueblo Viejo is based on a mining rate (approximately 40,000 tonnes per day) substantially exceeding the processing rate (approximately 24,000 tonnes per day), resulting in a 16 year life for the Moore and Monte Negro open pits compared with 26 years for the processing plant. Ore from several different areas of the mine will be mined concurrently and stockpiled according to gold content, metallurgical characteristics and sulphur grade. Ore with a higher gold grade will be mined and processed in the earlier years to benefit project economics. Ore with varying sulphur grades will be blended for processing at around an average sulphur content of 6.75% in order to maximize utilization of the autoclaves. The maximum size of the medium to long term stockpile is 82 Mt in Year 15, and material passing through the stockpile totals 127 Mt, or around 60% of all ore material mined.

AMC believes the operational strategy to be sound, but notes that the average sulphur content of the reserve is around 6.75%, meaning that stockpile control of sulphur grades will be critical. AMC also notes that there is an anticipated decay of stockpile sulphur content according to a formula calibrated with some limited information, including drilling, from sulphide material stockpiled between 1994 and 1998. AMC recommends that, if more detailed information is available concerning the original sulphur grades and specific time of deposition of these stockpiles, then additional drilling be undertaken to corroborate or better define the sulphur decay curve.

The open pit mine design, planning and optimization process is in line with common industry practice. Planned bench height is 10m with pit slope angles based on studies by a recognized consultant. Further geotechnical work is recognized by Barrick as being necessary in order to bring the geotechnical design aspects to a true feasibility level. Appropriate operating and, where necessary, sustaining capital costs were applied (these are commented on later) and a Ranking Index applied to each block to allow blocks with better gold, silver and copper grades and lower sulphur grades, to be selected for earlier mining and processing. Sensitivity analyses showed the pit sizes and recovered ounces to be moderately sensitive to gold prices and insensitive to pit slopes.

Limestone quarry scheduling is required to meet the needs for processing, tailings dams and other construction. Three limestone deposits (Quemados, Plant, and Las Lagunas) have been scheduled according to quality requirements. Although insufficient limestone tonnages have been delineated at this stage for total project requirements, AMC is satisfied that additional potential sources will yield sufficient tonnages of appropriate quality. AMC understands that it is the intention of Barrick to do additional definition drilling for both supply and quality purposes.



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AMC is satisfied that all aspects of the mine design, planning and optimization process have been undertaken to normal industry standards and all areas of material risk identified. There is a particular risk in years 1 and 2 when autoclave capability is still building, but high gold production is projected. A good understanding of the location and extent of high grade areas will be required, together with very selective mining and a disciplined stockpiling process. This also speaks to the previously identified issue about the use of ID³ for gold grade interpolation.

Metallurgical testwork conducted by Placer and Barrick showed that approximately 55% to 70% of the gold is encapsulated in sulphide minerals and is not recoverable by cyanide leaching without prior destruction of the sulphide matrix. There is also significant preg-robbing of gold by organic carbon in certain rock types. Extensive bench-scale and pilot plant testwork showed that pressure oxidation of the whole ore followed by carbon-in-leach (CIL) cyanidation of the autoclave product would recover 88% to 95% of the gold and 86% to 89% of the silver. The decision was made to use pressure oxidation in autoclaves at 230^oC to liberate the gold.

The proposed process design for Pueblo Viejo would make it one of the most complex, large precious metals projects in operation. However, considerable effort has gone into process development and AMC is satisfied that it represents a technically and economically viable treatment route that is not dependent on unproven technology. As previously noted, given the very strong dependency of process plant throughput and operating cost on sulphur grade, careful ore scheduling to maintain optimal sulphur levels is very important. There is some risk that excessive scale formation will occur at the slightly higher than typical autoclave operating temperature. This may lead to a need to reduce the operating temperature and result in slightly lower precious metals recoveries for some ore types.

Construction and operation of the Project will result in the physical and economic displacement of 369 households from three geographic areas, called Displacement Zones, within the overall Project Development Area. At the request of the government, Barrick assisted in the preparation of a Resettlement Action Plan (RAP), spending approximately $1.5 million in funding the assistance of expert technical personnel, local consultants, and local personnel and giving over a year of support to the government. The RAP was approved in September 2007 and signed by the representatives of the three communities, the Dominican State, Barrick, and the Catholic Church. The last two parties participated as observers and ensured that the process followed World Bank guidelines

There are a number of environmental issues at Pueblo Viejo. Past mining has resulted in uncontrolled release of acid rock drainage (ARD) and elevated metals originating from waste rock dumps, ore stockpiles and open pit rock walls. Two acid drainage treatment plants were constructed to treat the contaminated water; however, as of December 2007 only one treatment plant was operational. Barrick plans for mine development are designed to remediate or mitigate the majority of the existing problems within the project development area and to also improve environmental conditions outside the project area. The collection and treatment of contaminated surface water should effectively improve water quality and reduce both short-term and long-term environmental risks. However, Barrick has recognised that treatment strategies for the long-term post closure period will have to be developed.

It appears that the extent of existing ARD groundwater contamination and the potential for treatment, if necessary, is not sufficiently defined and requires further work.



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A tailings and waste rock storage facility (El Llagal) has been designed to store potentially acid generating materials including waste rock, tailings and water treatment sludge in a permanently flooded condition to minimize the potential for development of ARD. Two main earth embankment dams have been designed to meet Canadian dam safety standards. Environmental risks should be low if good construction, operation and monitoring practices are followed. A suitable long-term dam monitoring and maintenance program will be required post closure as part of risk management procedures for the mine. Should the mine close prematurely, there is a risk that some of the potentially acid generating waste rock deposited in the tailings and waste rock storage facility would be exposed above water, and also that the low grade ore stockpile would not be fully treated and present an ARD issue. There is no reason at this stage to anticipate such a scenario.

There is a risk that water quality may not comply with criteria for release to the receiving environment after the 75 year post closure period. It may therefore be necessary to consider longer term water treatment options and ensure adequate financial resources are available for continued treatment until such time as passive systems can be implemented to control the water quality.

In the feasibility study, power supply for the mine and process plant is indicated as being provided by a third party from a new, coal-fired power plant to be built on the south coast of the Dominican Republic. Distribution would be via a dual circuit 230 kV transmission line (111 km) along a corridor from the source to the mine site. The cost of building the transmission line is included in the construction cost of the project. Power cost has been assumed at US$0.10/kWh during the first year of operation, and US$0.08./kWh thereafter in the operating cost model.

Information from Barrick subsequent to the feasibility study indicates that power may not now come from a coal-fired plant but, possibly, from a combination of a new and existing heavy fuel oil generation facilities. The Owners and AMC recognize this item as one of the major project uncertainties, both in terms of the source of supply and its operating (and, possibly, capital) cost implications.

An Australian company, Las Lagunas Ltd, was granted a limited project approval in December, 2006 for exploitation of the Las Lagunas development area. Various environmental issues regarding the construction, operation, and closure of Las Lagunas and partial overlap with PVGP development areas have led to Barrick being in a position of conflict with the potential operation of Las Lagunas. These issues, as far as AMC is aware, remain unresolved.

The capital cost estimate in the Barrick FS is $2.59 billion, including a contingency of $291.7 million. It includes all engineering, procurement and construction costs for the mine development, process facilities and support infrastructure. It covers initial capital costs to bring the project into production and expenditures from the start of detailed engineering to the point of loading ore into the crusher. AMC is satisfied that the estimate is a reasonable, feasibility level reflection of project costs that would be incurred if the project were executed in accordance with the construction plan, schedule, and implementation plan as described in the Barrick FS. The project schedule is ambitious but realistic; however issues such as permitting and power supply could certainly affect the rates of project execution and capital expenditure.



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Operating cost estimates have been largely generated from first principles. AMC accepts that these estimates are done to industry standard and represent a reasonable projection of expenditures for operation of the Pueblo Viejo site as envisaged in the Barrick FS. Processing costs are seen to be the major component of operating cost at about 75% of the total. Power alone is at 36% of the total. Since the generation of the Barrick FS estimates, the price of oil has risen quite significantly. This obviously affects equipment fuel costs, but could also have a major effect on the price of power should heavy fuel oil be the energy source for the eventual generation facility or facilities. AMC also notes increased vulnerability to lower metal prices in the final 10 years of the project when mining is completed and the remaining ore to be processed comes entirely from the low grade stockpile. Unit operating cost declines from around $300/t to about $240/t, but cash cost/oz increases from around $370 to the range of $450 to $550.

The project is generally seen to be very sensitive to metal prices and operating costs, but also to be vulnerable to a reduction in gold production of less than 10% with the gold price at or below the Base Case level of $700/oz. The risk of a less than anticipated gold production rate, which is probably highest in the early years of operation, has specific relevance for both delivery of grade to the mill, and to the capability of the processing plant to operate as projected. Provision of power, the means of which is still uncertain, could have decidedly negative effects for both capital and operating costs, particularly at a time of high oil prices. On the positive side, sustained metal prices around $800/oz., or greater, show high economic returns and resilience against a significant increase in both capital and operating costs; and a decidedly positive net cash flow is the result for all considered scenarios other than where there is a 20% or more reduction in all metal prices or a 20% or more increase in total operating cost (all other parameters as per the Base Case).



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CONTENTS

EXECUTIVE SUMMARY


1	INTRODUCTION AND TERMS OF REFERENCE			1
	1.1	Units of Measurement and Conversion Factors		2

2	RELIANCE ON OTHER EXPERTS			4

3	PROPERTY DESCRIPTION AND LOCATION			5
	3.1	Location		5
	3.2	Land Status, Ownership and Special Lease Agreement		5

4	ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY			10
	4.1	Access		10
	4.2	Climate and Physiography		10
	4.3	Infrastructure and Local Resources		10

5	HISTORY			11
	5.1	Pre-1969		11
	5.2	Rosario / AMAX (1969-1992)		11
	5.3	Privatization (1996)		12
		5.3.1	GENEL JV	13
		5.3.2	Mount Isa Mines	13
		5.3.3	Newmont	13
		5.3.4	Placer	13
		5.3.5	Other Information	13

6	GEOLOGICAL SETTING			14
	6.1	Regional Geology		14
	6.2	Local Geology		15
		6.2.1	Introduction	15
		6.2.2	Hydrothermal Alteration	15
		6.2.3	Weathering	17
		6.2.4	Moore Deposit	17
		6.2.5	Monte Negro Deposit	18

7	DEPOSIT TYPES			20

8	MINERALIZATION			21
	8.1	General Description		21
	8.2	Metal Occurrence and Distribution		21
		8.2.1	Gold	21
		8.2.2	Silver	23
		8.2.3	Zinc	23
		8.2.4	Copper	24
		8.2.5	Lead	24



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		8.2.6	Moore Deposit	24
		8.2.7	Monte Negro	26

9	EXPLORATION PROGRAMS			30
	9.1	Barrick 2006 Work Program		30
		9.1.1	2006 Phase 1 Drilling Program	30
		9.1.2	2006 Phase 2 Drilling Program	31
	9.2	AMC Opinion		34

10	DRILLING			35
	10.1	Introduction		35
	10.2	Pre-Barrick Drilling Campaigns		38
		10.2.1	Rosario Drilling	38
		10.2.2	GENEL JV Drilling	39
		10.2.3	MIM Drilling	39
		10.2.4	Historical Drill Hole Surveying	40
		10.2.5	Placer Drilling	40
	10.3	2006 Barrick Drilling Program		41
	10.4	AMC Opinion		41

11	SAMPLING METHOD AND APPROACH			42
	11.1	Pre-Placer Drilling Programs		42
	11.2	Placer Diamond Drilling		42
	11.3	Barrick Diamond Drilling		42
	11.4	Sample Quality, Sample Recovery and Related Issues		42

12	SAMPLE PREPARATION, ANALYSES AND SECURITY			43
	12.1	Sample Preparation and Assaying Procedures		43
		12.1.1	Rosario	43
		12.1.2	GENEL JV	43
		12.1.3	MIM	43
		12.1.4	Placer	44
		12.1.5	Barrick	44
	12.2	QA/QC Procedures		45
		12.2.1	Rosario Check Assays, 1978	45
		12.2.2	Rosario Check Assays, 1985	45
		12.2.3	GENEL JV Checks	45
		12.2.4	Placer Checks, 2002	46
		12.2.5	Placer Checks, 2004	46
		12.2.6	ALS Chemex Quality Control	46
		12.2.7	Acme Check Assay Program	46
		12.2.8	Barrick Checks, 2006	47
	12.3	Summary		51
	12.4	AMC Opinion		52

13	DATA VERIFICATION			53
	13.1	Verification of Pre-Placer Data		53
		13.1.1	Database Development	53
	13.2	Rosario Pseudo-Twin Assay Pairing		53
		13.2.1	Historical Twinned Hole Comparisons	54



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	13.3	Verification of Pre-Barrick Data		55
		13.3.1	Verification of Placer Data	55
		13.3.2	Down-Hole Contamination of RC and Rotary Holes	55
		13.3.3	Cross Sectional Review of MIM, Rosario, and Placer Drilling	55
		13.3.4	Gold-Grade Distribution Comparisons	55
		13.3.5	Summary	58
		13.3.6	AMC Opinion	58

14	MINERAL RESOURCE ESTIMATES			59
	14.1	Introduction		59
	14.2	2005 Placer Mineral Resource Estimate		59
		14.2.1	Introduction	59
		14.2.2	Drill Hole Database	59
		14.2.3	Geological Modelling	60
		14.2.4	Topography	63
		14.2.5	Bulk Density	64
		14.2.6	Data Compositing	64
		14.2.7	Top Cutting	64
		14.2.8	Variogram Analysis	65
		14.2.9	Interpolation Plans	65
		14.2.10	Model Validation	65
		14.2.11	Resource Classification	66
		14.2.12	2005 Placer Mineral Resource Estimation Results	67
	14.3	2007 Barrick Feasibility Study Mineral Resource Estimate		67
		14.3.1	Introduction	67
		14.3.2	Geological Model	67
		14.3.3	Drill Hole Database	67
		14.3.4	AMC Opinion	68
		14.3.5	Topography	69
		14.3.6	Coordinate Units	70
		14.3.7	Raw Assay Statistics	70
		14.3.8	AMC Comment	70
		14.3.9	Top Cutting	72
		14.3.10	AMC Comment	73
		14.3.11	Assay Compositing	73
		14.3.12	Geological Solids and Model	75
		14.3.13	AMC Opinion	76
		14.3.14	Block Model	76
		14.3.15	Bulk Density	76
		14.3.16	Variography	77
		14.3.17	Gold Grade Estimation	78
		14.3.18	AMC Opinion	81
		14.3.19	Sulphur Grade Estimation	81
		14.3.20	AMC Opinion	82
		14.3.21	Resource Classification	82
		14.3.22	AMC Opinion	83
		14.3.23	Block Model Validation	83



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		14.3.24	AMC Comment	84
		14.3.25	Mineral Resource Summary	85
		14.3.26	Comparison with Placer 2005 Estimate	85
	14.4	AMC Comment and Opinion on Barrick 2007 Feasibility Study Resource Estimate		86
	14.5	Barrick Early-2008 Resource Estimate		87

15	ADJACENT PROPERTIES			88

16	MINERAL PROCESSING AND METALLURGICAL TESTING			89

17	ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES			90
	17.1	Mining Operations		90
		17.1.1	Site Conditions & Choice of Mining Method	90
		17.1.2	Mine Design Factors	92
		17.1.3	Mine Design and Planning Process	97
		17.1.4	Application of Variables	100
		17.1.5	Open Pit Optimization and Sensitivity Analysis	102
		17.1.6	Open Pit Design and Sequencing Method	105
		17.1.7	Mine Production Schedule & Forecast	107
		17.1.8	Mine Equipment	115
		17.1.9	Workforce Requirements	117
		17.1.10	Mineral Reserve Estimate	118
		17.1.11	Mineral Resource Estimate	118
		17.1.12	AMC Assessment and Opinion	118
	17.2	Mineral Processing and Metallurgical Testing		121
		17.2.1	Introduction	121
		17.2.2	Ore Mineralogy	121
		17.2.3	Metallurgical Investigation of Process Options	123
		17.2.4	Recoverability	131
		17.2.5	Limestone and Lime Plant Description	141
		17.2.6	Process Risk Summary	142
	17.3	Infrastructure		144
		17.3.1	General Infrastructure	144
		17.3.2	Power Plant	144
		17.3.3	Site Electrical System	144
		17.3.4	Process Control Facilities	145
		17.3.5	Communication Facilities	145
		17.3.6	Fuel	145
		17.3.7	Water Supply	145
		17.3.8	Storm Water	146
		17.3.9	Waste Management	146
		17.3.10	Sewage Treatment	146
		17.3.11	Fire Protection	146
		17.3.12	Dust Control	146
		17.3.13	Landfill	146
		17.3.14	Las Lagunas Project	147



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		17.3.15	Resettlement Action Plan (RAP)	147
		17.3.16	AMC Comments	147
	17.4	Markets		148
		17.4.1	Metal Prices	148
		17.4.2	Doré Shipping and Refining	148
		17.4.3	Copper Concentrate Shipping and Refining	148
	17.5	Contracts		148
	17.6	Environmental Considerations		149
		17.6.1	Scope	149
		17.6.2	Authorizations and Responsibilities	149
		17.6.3	Environmental Standards	152
		17.6.4	Existing Environmental Conditions	152
		17.6.5	Environmental Baseline Studies	153
		17.6.6	Environmental Issues for Mine Operation	156
		17.6.7	Mine Closure and Post Closure Impacts	159
		17.6.8	Reclamation and Bond	160
		17.6.9	Risks and Liabilities	160
	17.7	Taxes		162
		17.7.1	Taxes and Payments	162
		17.7.2	Special Lease Agreement Summary	162
	17.8	Capital Costs		162
		17.8.1	Basis of Estimate	162
		17.8.2	Capital Cost Summary	163
		17.8.3	Estimate Base Date and Exchange Rates	164
		17.8.4	Contingency	164
		17.8.5	Exclusions	164
		17.8.6	Sustaining Capital Costs	165
		17.8.7	AMC Comment	165
	17.9	Operating Costs		166
		17.9.1	Operating Cost Summary	166
		17.9.2	Operating Cost Areas	167
		17.9.3	AMC Comment	169
	17.10	Economic Analysis		170
		17.10.1	Revenue	170
		17.10.2	Capital Expenditure	170
		17.10.3	Net Cash Flow, NPV, IRR	171
		17.10.4	Sensitivity Analysis	172
		17.10.5	Payback Period	173
		17.10.6	AMC Analysis and Comment	173

18	OTHER RELEVANT DATA AND INFORMATION			178

19	INTERPRETATION AND CONCLUSIONS			179
	19.1	Introduction		179
	19.2	Conclusions		179

20	RECOMMENDATIONS			183

21	REFERENCES			184



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22	DATE AND SIGNATURE PAGE	185

23	QUALIFIED PERSON’S CERTIFICATES	186

TABLES


Table 3.1	PVDP Permit Status at December 2007	7
Table 8.1	Mineralogically Determined Deportment of Gold	23
Table 10.1	Summary of Drilling Campaigns	35
Table 10.2	Rosario Drill Hole Summary	38
Table 11.1	Sample Interval Data for Rosario, GENEL JV and MIM Drill Holes	42
Table 12.1	Summary of Placer / ALS Assaying Procedures	44
Table 14.1	Drill Holes and Metres used for 2005 Placer Resource Estimate	59
Table 14.2	Moore Deposit Zone Names, Placer 2005 Estimate	61
Table 14.3	Monte Negro Deposit Zone Names, Placer 2005 Estimate	62
Table 14.4	Resource Classification and Estimation Statistics, 2005
	Placer Estimate	66
Table 14.5	2005 Placer Resource Summary at 1.7 g/t Au Cut-off
	Grade (100% Basis)	67
Table 14.6	Drill Holes and Metres used for 2007 Barrick Resource Estimate	68
Table 14.7	Gold Assay Statistics	71
Table 14.8	Block Model Geometry	76
Table 14.9	Search and Sample Selection Parameters for Gold Indicator
	Estimates (High and Low Grade)	79
Table 14.10	Search and Sample Selection Parameters for Gold Grade Estimates	80
Table 14.11	Classification Criteria	83
Table 14.12	Total Mineral Resources at a 1.4 g/t Au Cut-off Grade	85
Table 14.13	Comparison of 2005 Placer and 2007 Barrick FS Resource
	Estimates (100%)	86
Table 17.1	Main Statistics for Densities	94
Table 17.2	Ore Domains and Metallurgical Recoveries	94
Table 17.3	Block Model Basic Parameters	98
Table 17.4	Metal Fields	99
Table 17.5	Category Field	99



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Table 17.6	Metallurgical Field	99
Table 17.7	Payable Metal Transport and Refining Charges	100
Table 17.8	El Llagal Sustaining Capital Costs	102
Table 17.9	Pit Optimization Slope Angles	103
Table 17.10	Pueblo Viejo Pit Optimization Tonnages	104
Table 17.11	Pit Optimization Sensitivity to Gold Price	104
Table 17.12	Pit Optimization Sensitivity to Pit Wall Slope	105
Table 17.13	Slope Design Parameters based on Piteau Recommendations	106
Table 17.14	Phase Mining Sequence	107
Table 17.15	Autoclave Ramp-Up	109
Table 17.16	Summary of Long-Term Mine Plans	110
Table 17.17	High Sulphur Ore Cut-off Grades	111
Table 17.18	Phase Ore Mining by Period	113
Table 17.19	Project Limestone Requirements	114
Table 17.20	Limestone Tonnages and Uses	114
Table 17.21	Total Mine Labour per Period	117
Table 17.22	Mineral Reserves	118
Table 17.23	ID³ and OK Estimates of Mill Feed Y01 and Y02	121
Table 17.24	Summary of Metallurgical Test Programs (from PAH)	124
Table 17.25	Comminution Testwork	126
Table 17.26	Limestone and Lime Plant Design Basis (Expansion 24,000 t/d
	Ore Processing Rate)	141
Table 17.27	Project Capital Cost Estimate (as at Q3 2007)	163
Table 17.28	Project Capital Cost — 18kt/d and 24kt/d — by Responsibility	164
Table 17.29	Sustaining Capital Cost Summary*	165
Table 17.30	Operating Cost Summary	170
Table 17.31	Total Projected Revenues by Product	170
Table 17.32	Cash Flow Summaries Base Case	173
Table 17.33	Cash Flow Summaries Downside Case	173
Table 17.34	Cash Flow Summaries Optimistic Case	173
Table 17.35	AMC Economic Sensitivity Analysis	177



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FIGURES


Figure 3.1	Location Diagram	5
Figure 3.2	Montenegro Fiscal Reserve	6
Figure 5.1	Rosario Mine Workings and Plant	12
Figure 6.1	Regional Geology	14
Figure 6.2	Geological Cross Sections, Moore and Monte Negro Deposits	16
Figure 6.3	Block Model Gold Grades relative to Propylitic Boundary	17
Figure 8.1	Alteration and Mineralization, Moore Deposit Section 94600 N, West Flank and Vein Zone (Looking North)	25
Figure 8.2	Alteration and Mineralization, Monte Negro Deposit Central Zone, Section 95650N (Looking North)	27
Figure 9.1	Main Results of 2006 Phase 1 Drilling Program	31
Figure 9.2	Main Results of 2006 Phase 2 Drilling Program	32
Figure 9.3	Main Results of 2007 Drilling Program	33
Figure 9.4	Monte Oculto Discovery	34
Figure 10.1	Location of all Drill Holes, Moore Deposit	36
Figure 10.2	Location of all Drill Holes, Monte Negro Deposit	37
Figure 12.1	Barrick 2006 / 07 QA/QC Results — Blanks	47
Figure 12.2	Barrick 2006 / 07 QA/QC Results — Standards PV2, PV4, PV5	48
Figure 12.3	Barrick 2006 / 07 QA/QC Results — Standards PV1, PV7	49
Figure 12.4	Barrick 2006 / 07 QA/QC Results — Core Duplicates	50
Figure 12.5	Barrick 2006 / 07 QA/QC Results — Sample Grain size	51
Figure 13.1	AMEC Comparison of Placer and Rosario Drill Hole Assays within 10 m	54
Figure 13.2	Frequency Distribution of Gold by Drilling Campaign: All Drill Holes vs. Placer Rotary	56
Figure 13.3	Frequency Distribution of Gold by Drilling Campaign: All Drill Holes vs. Barrick DDH	57
Figure 14.1	Moore Section 94600 N, Lithology and Structural Domains	62
Figure 14.2	Monte Negro Section 95800N Lithology and Structural Domains	63
Figure 14.3	Frequency Distribution of Gold by Drilling Campaign: All Drill Holes vs. Rosario RC	69



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Figure 14.4	Frequency Distribution of Raw Gold Assays for All Rock Types	72
Figure 14.5	Raw Gold Assay Cutting	73
Figure 14.6	Gold Assay Statistics at Varying Composite Lengths	74
Figure 14.7	Frequency Distribution of Gold Grades in 10m Composites vs. Raw Assays	75
Figure 14.8	Down-Hole Correlogram Gold	77
Figure 14.9	Omni-Directional Correlogram Gold	78
Figure 14.10	Frequency Distribution of Sulphur Grades in 10m Composites	81
Figure 14.11	Omni-Directional Correlogram — Sulphur	82
Figure 14.12	Composite — Model Block Gold Grade Comparison	84
Figure 17.1	Moore Pit from Monte Negro	91
Figure 17.2	Old Processing Plant at Pueblo Viejo	92
Figure 17.3	Ore Treatment Rate	93
Figure 17.4	Workings in Monte Negro Pit	96
Figure 17.5	Pit Optimization Illustration	103
Figure 17.6	Section at 95,600 Monte Negro	107
Figure 17.7	Ore to Crusher	112
Figure 17.8	Mine Daily Movement (excluding Quarries)	112
Figure 17.9	Equipment Hours Model	116
Figure 17.10	Effect of Gold Head Grade on Gold Recovery	127
Figure 17.11	Effect of Temperature on CIL Silver Extraction from Lime Boil Plant Operation	128
Figure 17.12	Relationship between Gold Recovery and Organic Carbon Content	129
Figure 17.13	Pueblo Viejo — Simplified Flowsheet	132
Figure 17.14	Development Areas of Las Lagunas	151
Figure 17.15	Catchment Areas	157
Figure 17.16	Project Capital Cash Flow	171
Figure 17.17	Project Net Cash Flow	172

Distribution list:


3 copies to:		Mr R Bryson, Goldcorp Inc, Vancouver
1 copy to:		AMC Vancouver office
1 copy to:		AMC Melbourne office



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1		INTRODUCTION AND TERMS OF REFERENCE

This Technical Report on the Pueblo Viejo Gold Project (PVGP) in the Dominican Republic has been prepared for Goldcorp Inc. (Goldcorp) of Vancouver, Canada by Mr H A Smith and Mr P R Stephenson of AMC Mining Consultants (Canada) Ltd (AMC) of Vancouver, Canada, Mr M G Butcher of Goldcorp, and Mr C A Carr of Rescan Environmental Services Ltd. (Rescan) of Victoria, Canada. It has been prepared in accordance with the requirements of National Instrument 43-101 (NI 43-101), “Standards of Disclosure for Mineral Project”, of the Canadian Securities Administrators (CSA) for lodgement on CSA’s “System for Electronic Document Analysis and Retrieval” (SEDAR).

Goldcorp owns 40% of the PVGP, the other 60% being owned by Barrick Gold Corporation (Barrick) which is also the operator. . On February 21, 2008, Goldcorp received the results of an update to an existing 2005 Placer Dome Feasibility Study (PDFS) on the PVGP prepared by or on behalf of Barrick (the Barrick 2007 Feasibility Study or Barrick FS). This Barrick update includes, among other elements, an updated production schedule, revised process circuit and an updated capital cost estimate. On or before March 31, 2008, Goldcorp was required to file its Annual Information Form (AIF) and Goldcorp believed that in order to provide up-to-date, full, true and plain discourse, it was necessary that the information contained in the Barrick FS form the basis of the scientific and technical information on the PVGP contained in the AIF. As the Barrick FS information is new material scientific or technical information, filing the AIF containing this information triggered a requirement to file a technical report to support such information not later than the time the AIF was filed. Since the time frame between the receipt of the Barrick FS and the deadline for filing the AIF was short, Goldcorp applied for, and was granted, exemptive relief from the appropriate securities regulatory authorities from the requirement in NI 43-101 that it file a technical report for the Pueblo Viejo Project not later than filing its AIF provided that:

This annual information form includes the following cautionary language:

“The technical disclosure, including the Mineral Reserve and Mineral Resource estimates, in this annual information form with respect to the Pueblo Viejo Project has not been supported by a technical report prepared in accordance with NI 43-101. A technical report is being prepared by qualified persons under NI 43-101 and it will be available for review on the SEDAR website located at www.sedar.com under the Corporation’s profile on or before May 15, 2008. Readers are advised to refer to that technical report when it is filed.” and

Goldcorp files the technical report for the Pueblo Viejo Project as soon as practicable but, in any event, not later than May 15, 2008.

The cut-off dates for exploration drilling data used for the Barrick FS were January 2007 for gold assays and June 2007 for sulphur assays. The mineral resource and mineral reserve estimates were as at end-June 2007 and the Barrick FS itself was effective as at third quarter 2007. Further exploration has taken place since June 2007 and updated mineral resource and mineral reserve estimates, effective end-December 2007, were released by Barrick in early 2008. Also further studies have been undertaken since completion of the Barrick FS. To the extent that these later estimates and studies are material to this Technical Report, commentary is included in the appropriate sections of the Report.

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The names and details of persons who prepared or contributed to this Technical Report are listed in Table 1.1.

Table 1.1 Persons who Prepared or Contributed to this Technical Report

Qualified

Ind of

Date of Site

Professional

Person

Position

Employer

Goldcorp

Visit

Designation

Sections of Report

Qualified Persons responsible for the preparation and signing of this Technical Report

Mr H A
Smith

Principal
Mining
Engineer

AMC Mining
Consultants
(Canada) Ltd

Yes

18-19 March,
2008

BSc, MSc, PEng
(Ont), PEng
(AB), MCIM

Section 15, Section 17 other than Processing and Environmental aspects, With Mr Stephenson, Sections 18-21 & Exec Summary

Mr P R
Stephenson

Principal
Geologist

AMC Mining
Consultants
(Canada) Ltd

Yes

18-19 March,
2008

BSc, MCIM,
FAIG, FAusIMM
(CP)

Sections 1-14. With Mr Smith, Sections 18-21 & Exec Summary

Mr M G
Butcher

Group
Metallurgist

Goldcorp Inc

No visit

BAppSc (App
Chem),
MAusIMM

Section 16 and Processing aspects of Section 17

Mr C A Carr

Senior
Geotechnical
Engineer

Rescan
Environmental
Services Ltd

Yes

18-19 March,
2008

BSc, PEng (BC),
MCGS, MCDA

Environmental aspects of Section 17

The scope of the personal inspection of the property undertaken by the Qualified Persons covered:

•		Interviews in Santiago with key Barrick personnel

•		Interviews on site with key Barrick and project personnel

•		Site tours of existing and planned site infrastructure

•		Examination of drill core, core processing and sample preparation facilities

•		On-site examination of plans, cross sections, photographs and other diagrams

The Technical Report is based on information provided by Goldcorp, a list of which is contained in Section 21 — References, on a site visit undertaken by the Qualified Persons in March 2008, and on discussions with Goldcorp and Barrick personnel.

Goldcorp was provided with a draft of this report to review for factual content and conformity with the brief.

This report is effective May 01 2008.

1.1 Units of Measurement and Conversion Factors

The Metric System or System International (SI) is the primary system of measure and length used in this report. Conversions from the Metric System to the Imperial System are provided below for general guidance.

Metals and minerals acronyms in this report conform to mineral industry accepted usage. Further information is available online from a number of sources, including web site: http://www.maden.hacettepe.edu.tr/dmmrt/index.html.



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The following conversion factors are used in this report:

1 hectare = 2.471 acres

1 hectare = 0.00386 square miles

1 square kilometre = 0.3861 square miles

1 metre = 3.28084 feet

1 kilometre = 0.62137 miles

1 gram = 0.03215 troy ounces

1 troy ounce = 31.1035 grams

1 kilogram = 2.205 pounds

1 tonne = 1.1023 short tons

1 gram/tonne = 0.0292 troy ounces/short ton

A more complete list of conversion factors can be found on the following web site: http://www.empr.gov.bc.ca/Mining/Geolsurv/MINFILE/manuals/coding/Hardcopy/appdvii.htm.

The term grams/tonne or g/t is equivalent to 1 ppm (part per million) = 1000 ppb (part per billion). Other abbreviations include: oz/t = ounce per short ton; Moz = million ounces; Mt = million tonnes; t = tonne (1000 kilograms); wt% = percent by weight; % = ppm/10,000; m = metre; km² = square kilometres; ha = hectare; BD = bulk density; SG = specific gravity; lb/t = pounds/tonne.

Dollars are expressed in Unites States currency (US$) unless otherwise stated.

Prices of gold and silver are stated in US$ per troy ounce (US$/oz). The price of copper is stated in US$ per pound (US$/lb).



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3		PROPERTY DESCRIPTION AND LOCATION

3.1 Location

Pueblo Viejo is located in the central part of the Dominican Republic on the Caribbean island of Hispaniola in the province of Sanchez Ramirez (Figure 3.1) The PVGP is 15 km west of the provincial capital of Cotui and approximately 100 km northwest of the national capital of Santo Domingo.

Figure 3.1 Location Diagram

3.2 Land Status, Ownership and Special Lease Agreement

Pueblo Viejo Dominicana Corporation (PVDC) is the holder of th e right to lease the Montenegro Fiscal Reserve (MFR) by virtue of a Special Lease Agreement of Mining Rights (Special Lease Agreement or SLA), which was ratified by the National Congress of the Dominican Republic by means of Resolution No. 125-02, dated as of August 26, 2002. Pursuant to the Special Lease Agreement, PVDC has the exclusive right to lease the Montenegro Fiscal Reserve (covering the Pueblo Viejo deposits and other related sites) free and clear of all defects, claims or encumbrances, for the term of the leases for exploitation of the minerals contained in the MFR under the terms, conditions, stipulations and agreements set forth in the Special Lease Agreement.



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Dominican State, with the consent of Rosario, terminated the Pueblo Viejo Concession and the Pueblo Viejo II Concession. On March 7, 2002, the Dominican State, by virtue of Presidential Decree No. 169-02, created the Montenegro Fiscal Reserve with an area of 3,200 ha. On August 3, 2004, The Dominican State, by virtue of Presidential Decree No. 722-04, modified the Montenegro Fiscal Reserve to include El Llagal.

The lease includes all surface rights and improvements (including the existing mill) owned by the Dominican State within the boundaries of the fiscal reserve. The initial term of the lease would be 25 years from date of the Project Notice. At PVDC’s election, the term of the lease could be further extended for another 25 years. If at the end of that 50 year period, both Parties agree, the term could be further extended for another 25 years for a total lease of 75 years.

The SLA is discussed further in Section 17.6.2.1, with Royalty and Tax payment requirements being detailed in Section 17.7.

Environmental liabilities are discussed in Section 17.6 and, particularly, Sections 17.6.4 and 17.6.9. Bonding issues for the project Environmental License are discussed in Section 17.6.8.2

Table 3.1 shows the status of PVDP Permits as of the date of the Barrick FS.

Table 3.1 PVDP Permit Status at December 2007


Permit #	Permit or Authorization	Status
1	Presentation of Análisis Previo Form	Completed
2	Terms of Reference	Obtained
3	Soil Use	Permit received for MN Reserve and Waterline Corridor July 12, 2005. It may need to be re-applied for.
4	Water Rights	Obtained, September 8, 2005 subject to the signature of an agreement (still in discussion) Additional rights are under review.
5	Environmental Authorization for Soil Use	Received for MN Reserve (Jul 5.05). Waterline Corridor defined, change soil use permit under review.
6	Hydraulic Works	Obtained with Environmental License; detailed engineering will be submitted. Document preparation for the UL-2 will start in January 2008.
7	Authorization for Deep Well Construction	Pending
8	Environmental Management Plan	Approval obtained with Environmental License (December 26, 2006). Update issued in September 07 to include Ag/Cu recovery.



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Permit #	Permit or Authorization	Status
9	Estudio de Factibilidad	Draft prepared.
10	Environmental License or Permit	Obtained December 26, 2006. ICA#1 submitted. ICA#2 in preparation — due March 2008.
11	Extraction of Non-Metallic Materials and Construction Material from Terrestrial Crust	Obtained December 26, 2006.
12	Authorization for Installation of Aqueduct and Drainage System	In progress — Early Works Package.
13	Authorization for Final Disposition of Non- Hazardous Solid Residues	Environmental License request to present the Plan — for notification only. Submission in preparation.
14	Authorization for Final Disposition of Hazardous Solid Residues	Environmental License request to present the Plan — notification only. Submission in preparation
15	Authorization for Final Disposition of Radioactive Residues	Environmental License request to present the Plan — notification only. Submission in preparation.
16	Authorization for Exploitation of Groundwater	Environmental License request to present Plan — for notification only — submission in preparation.
17	Hazardous Substances Transport	Environmental License request to present the Plan — notification only. Submission in preparation.
18	Mining Right of Way	Power line corridor pending.
19	Tree Cutting	In progress — Early Works Package.
20	Construction	In progress — Early Works Package.
21	Authorization for Processing Plants	Possibly obtained in SLA — under investigation.
22	Sanitary Authorization for Industrial Facilities	Under investigation.
23	Authorization for Food Installations	In progress — Early Works Package.
24	Authorization for Installation of Industries	In progress — Early Works Package for Construction Phase.
25	Announcement of Boiler Operation	(Operation Phase)
26	Authorization for Combustible Storage	In progress — Construction Phase.



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4		ACCESSIBILITY, CLIMATE, LOCAL RESOURCES, INFRASTRUCTURE AND PHYSIOGRAPHY

4.1 Access

Access from Santo Domingo is by a four lane, paved highway (Autopista Duarte) that is the main route between Santo Domingo and the second largest city, Santiago. This highway connects to a secondary Highway, #17, at the town of Piedra Blanca, approximately 78 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 PVGP is 22 km from Piedra Blanca.

The sufficiency of surface rights for PVGP mining operations is discussed in Section 3.2.

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 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. The highest point in the Cordillera Central is Pico Duarte at 3,175 m. Pueblo Viejo 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 Hatillo Reservoir. 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 1.8m, 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.

4.3 Infrastructure and Local Resources

Infrastructure and local resource issues and requirements are discussed in Section 17.2.

Personnel issues and requirements are addressed in Section 17.1.9.

Details of tailings and waste storage issues and requirements are discussed in Section 17.6.


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5		HISTORY

5.1 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 the second visit of Columbus in 1495 probably found the deposit being actively mined by the native population. The Spanish 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.

5.2 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 on to higher ground, the thickness of the oxide cap increased to a maximum of 80m, 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). 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 on the underlying refractory sulphide resource in an effort to continue the operation. Feasibility level studies were conducted by Fluor Engineers Inc. (Fluor) in 1986, and Stone & Webster Engineering / American Mine Services (SW/AMS) in 1992.

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.


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

Rosario continued to mine the 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 high cyanide consumption) and ore hardness. Mining ceased in the Monte Negro deposit 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 5.5 M oz of gold and 25.2 M oz of silver. Figure 5.1 shows a photographic overview of the Rosario mine workings and plant as at early 2008.

Figure 5.1 Rosario Mine Workings and Plant

5.3 Privatization (1996)

Lacking funds and technology to process the sulphide ore, Rosario Dominicana attempted two bidding processes to joint venture the property, one around 1992 and the other in 1996. In November 1996, Rosario selected Salomon Brothers (Salomon Smith Barney) to coordinate 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 (Joint Venture), Mount Isa


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Mines, and Newmont. This privatization was not achieved but each of the three companies conducted work on the property during their evaluations.

5.3.1 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 $6 million 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.

5.3.2 Mount Isa Mines

In 1997, Mount Isa Mines (MIM) conducted a due diligence program as part of its effort to win Pueblo Viejo in the privatization process. It conducted a 31 hole, 4,600m diamond drilling program, collected a metallurgical sample from drill core, carried out detailed pit mapping, completed induced polarisation (IP) geophysical surveys over the known deposits and Organizing aerial photography over the mining concessions to create a new (1997) surface topography. MIM also proposed to carry out a pilot plant and feasibility study using ultra-fine grinding/ferric sulphate leaching.

5.3.3 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.

5.3.4 Placer

Between 2003 and mid-2005, Placer completed extensive work on Pueblo Viejo including drilling, geological studies and mineral resource / reserve estimation. This work was compiled in a feasibility study completed in July 2005. The mineral resource and mineral reserve estimates are commented upon under Section 14 — Mineral Resource Estimation.

5.3.5 Other Information

Refer also to Sections 10 to 13 for additional information on historical exploration.


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6		GEOLOGICAL SETTING

6.1 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 (Figure 6.1).

Figure 6.1 Regional Geology

The Los Ranchos Formation consists of a lower complex of pillowed basalt, basaltic andesite flows, dacitic flows, tuffs, and intrusions, overlain by volcaniclastic sedimentary rocks, and interpreted to be a Lower Cretaceous intraoceanic island arc. 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 forming 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. The sedimentary basin and volcanic debris flows are considered to



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be of Neocomian age (121 Ma to 144 Ma). The Pueblo Viejo Member is bounded to the east by volcaniclastic rocks, and to the north and west by spilite flows and dacitic domes.

To the south, the Pueblo Viejo Member is overthrust by the Hatillo Limestone Formation, thought to be Cenomanian (93 Ma to 99 Ma), or possibly Albian (99 Ma to 112 Ma) in age.

Outside of the deposit areas, saprolite is as much as 25m 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.

6.2 Local Geology

6.2.1 Introduction

The Pueblo Viejo property comprises several high sulphidation (or acid-sulphate) epithermal deposits of which Moore and Monte Negro are the largest. (Figure 6.2) The deposits form funnel shaped envelopes of advanced argillic alteration where hydrothermal fluids migrated upwards and laterally along permeable horizons.

6.2.2 Hydrothermal Alteration

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. 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 mineralisation, siderite is completely replaced by pyrite.



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Figure 6.2 Geological Cross Sections, Moore and Monte Negro Deposits

In the Moore deposit, silica and kaolinite are more common in the upper parts of the system. In the now depleted oxide mineralization, silicification was closely associated with gold mineralization and caused mineralized zones to form hills with relief of about 200m. 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. 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.

The relationship of gold mineralization to advanced argillic alteration is shown in Figure 6.3, a cross section through the block model with the green zone representing the contact between advanced argillic alteration above and propylitic alteration below. The white zone represents the latest pit design. Blue, yellow and red lines outline blocks containing increasing gold mineralization.



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Figure 6.3 Block Model Gold Grades relative to Propylitic Boundary

6.2.3 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 80m at North Hill in the Moore deposit, to 50m in the South Hill and East Mejita deposits, to nothing in the stream valleys. The thickest oxide mineralization 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 (less than 100 µm) could be panned from only the highest grade zones. Silver was depleted in the near-surface parts of the oxide mineralization and enriched at the oxide-sulphide interface. Zinc and copper were leached from the oxide with the destruction of the sulphides.

6.2.4 Moore Deposit

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 (Pueblo Viejo sediments) overlying horizons of spilite, volcanic sandstone, and fragmental volcaniclastic rocks. The entire sequence has a shallow dip to the west (Figure 6.2).

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



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appearance, 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 north-north-east striking root zone to the FDP. The 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 Pueblo Viejo Sediments. The sub-basin is partially filled with a mixed sedimentary sequence consisting of inter-fingering Pueblo Viejo 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.

Bedding generally dips shallowly westwards (less than 25°) but locally, steep faults with north-north-east and north-north-west strikes have rotated bedding into steep orientations. The north-north-east faults preserve evidence for east-side-up and left-lateral sense of movement subsequent to mineralization. The north-north-east faults appear to link with a north-north-west 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. The 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.

6.2.5 Monte Negro Deposit

The Monte Negro deposit is located at the north-western margin of the sedimentary basin (Figure 6.2). Stratigraphy consists of inter-bedded carbonaceous sediments ranging from siltstone to conglomerate, 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 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 partly silicified, spilite-derived conglomerate. The horizon ranges in thickness from tens of metres to non-existent and is likely filling channels in the uneven spilite surface below.

Thin section work on the spilites suggest that an intrusive dome or near surface plug of unknown dimensions may exist under the west hill of Monte Negro. Dykes that intrude the Monte Negro stratigraphy include a steeply-dipping, north-north-west striking, propylitically altered, mafic dyke approximately 10m wide which is barren of gold mineralization. Similar but thinner dykes have been intersected in core in the west part of the deposit. Thin breccia dykes have also been mapped in the pit walls.

Inter-bedded 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, andesitic volcanic and volcaniclastic rocks generally dip shallowly (13°) towards the north-west. A steep north-north-west trending fault (Monte Negro Fault) with a west-side up sense of movement is interpreted to separate the sediments in the east from



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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 north-north-west sub-parallel to the strike of the north-north-west 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.

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



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7		DEPOSIT TYPES

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

The geological setting of the deposit is not certain at this time. Sillitoe and Bonham (1984), Muntean and others (1990), and 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.

The uncertainty as to origin has no practical impact on exploration of the deposit at the levels that may be mined by open pit methods. The areal 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.



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8		MINERALIZATION

8.1 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.

There were three stages of advanced argillic alteration associated with precious metal mineralization:

•		Stage I 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

•		Stage III occurred when hydro-fracturing of the silica cap produced pyrite-sphalerite-enargite veins with silicified haloes

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 north-north-west and can be 250m long, 100m wide and 100m 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 a number of local structures that crosscut the deposit. Those structures have a northerly trend at Monte Negro and Moore, with a northwest-southeast trend 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 south-west 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.

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.

8.2 Metal Occurrence and Distribution

8.2.1 Gold

Gold is intimately associated with pyrite veins, disseminations, replacements, and layers within the zones of advanced argillic alteration. Gold values are generally the highest in zones of silicification or strong quartz-pyrophyllite alteration. These gold-bearing alteration



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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 JV 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 north or northwest. 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 in the form of lead-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 that is intensely silicified and bounds high-grade mineralization at the surface. A second possibility is a deeper zone of mineralization that has been intersected by a vertical core hole testing an IP-chargeability anomaly approximately 100 m east of the main deposit.

AMTEL Laboratories of 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 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.

Gold occurs as native gold, sylvanite (AuAgTe4), and aurostibnite (AuSb2). The principal carrier of gold is pyrite where the sub-microscopic gold occurs in colloidal-size micro-inclusions (less than 0.5 µm) and as a solid solution within the crystal structure of the pyrite. The abundance of the gold minerals varies significantly between the different composites (see Table 8.1).



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Table 8.1 Mineralogically Determined Deportment of Gold


	MN-BSD		MN-VCL		MO-BSD		MO-VCL
Form and Carrier 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
Sub-Microscopic Gold
Micro-inclusions		51.7		33.4		28.0		19.8
Solid solution		13.6		21.5		26.4		10.3

There are four major forms of pyrite: Micro-crystalline, disseminated, porous, and coarse-grained. The micro-crystalline pyrite tends to have the highest gold concentration. This type of pyrite is also the most arsenic-rich, which renders it the most prone to oxidation, and the most difficult to liberate, as it forms complex intergrowths within the rock and with sphalerite. The coarse-grained form of pyrite has the lowest gold concentration and has a well-developed crystal habit making it less susceptible to oxidation.

Gold minerals are also found to a lesser extent as inclusions in enargite, quartz and lead-sulphosalts (primarily geocronite). Gold may also exist in the crystal structure of sulphosalts, such as enargite and geocronite but additional research is required.

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 sub-microscopic gold within the pyrite.

8.2.2 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 native silver and in pyrargyrite (antimony sulphide), hessite (silver telluride), sylvanite and petzite (gold tellurides) and tetrahedrite.

8.2.3 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 south-west 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 sub-microscopic gold mineralization.



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

8.2.4 Copper

Most of the copper occurs as enargite hosted in Stage III sulphide veins. Only trace amounts of chalcocite and chalcopyrite have been recorded. 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.

8.2.5 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 JV shows a strong correlation between gold and lead. Elevated lead values were found in the structural feeder zone in the Moore deposit and lead may provide clues on where to search for other feeder zones.

8.2.6 Moore Deposit

8.2.6.1 General

Pyrite-rich, gold-bearing veins at Moore have a mean width of 4 cm and are steeply-dipping with a trend commonly north-north-west. Subdominant pyrite vein-sets trend north-south and north-north-east. The orientation of pyrite veins and steep faults is similar, albeit with different dominant sets (north-north-west for veins and north-north-east for faults). This indicates a probable genetic link between steep faulting and vein development.

8.2.6.2 West Flank Zone

Thinly bedded carbonaceous siltstones and andesitic sandstones in the West Flank dip shallowly westwards. Dips increase towards the west where northerly-trending thrusts displace bedding (Figure 6.2).

Pyrite and limonite-rich veins with gold mineralization are sub-vertical and trend commonly north-north-west. The veins are oblique to the general north-north-east strike of bedding and do not appear to have been rotated. Quartz veins with gold trend north-west oblique to the pyrite veins 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 north-north-west.



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8.2.6.3 North and South Hills Zones

Bedding to the north of the Moore dacite porphyry dips shallowly westwards. Bedding has been rotated about both north-north-west and north-north-east axes. The change in bedding orientation reflects movement associated with north-north-west and north-north-east trending faults.

There are three steep-dipping, gold-bearing, pyrite-rich vein sets: north-west, north-east and north-south. North-west trending veins generally contain enargite and sphalerite, while north-east trending veins are more pyrite ± pyrophyllite rich. The average vein width is 3.5 cm.

The fault pattern is dominated by steep north-north-east trending faults that appear to link with north-north-west trending faults. A north-north-east 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 north-north-west 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.

8.2.7 Monte Negro

8.2.7.1 Monte Negro Central Zone

Pyrite-rich veins with gold mineralization are sub-vertical and have bimodal trends, which are interpreted to form conjugate sets. The mean width is 2 cm. The north-north-west trending set 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 north-north-east 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 north-north-west, and is 150m wide and up to 100m thick between the F5 Fault to the east and the Main Monte Negro Fault to the west.

The fault pattern is dominated by steep north-north-west 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 25m wide and 500m long. It is interpreted as a major fault that was active during and subsequent to mineralization.



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8.2.7.2 Monte Negro South Zone

Andesitic volcanic and volcaniclastic rocks with minor intercalations of carbonaceous sediments dip shallowly northwards. Close to the interpreted Monte Negro Fault, bedding dips more westerly and strikes north-north-west.

North-north-west trending steep faults displace bedding and dip towards the south-west. 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 cm to 6 cm). The veins are sub-vertical and trend north-west. The episodic vein fill demonstrates a clear paragenesis (massive pyrite-enargite-sphalerite-grey silica).

Shallow-dipping bedding and sub-vertical 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 150m along the West Thrust Fault (Figure 8.2).

8.2.7.3 Mineralisation 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 sub-vertical structures and stratigraphic zones. The stratigraphic shape of some zones may be controlled by sub-horizontal structures that contain pyrite veins. The veins are tens of centimetres wide 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 that served as conduits for hydrothermal fluids.

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 centre 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



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9		EXPLORATION PROGRAMS

Pre-Barrick exploration programs are covered in Sections 5, 10, 11 and 12.

9.1 Barrick 2006 Work Program

The main components of Barrick’s 2006 work program, which provided data for input to the Barrick FS, were:

•		Data compilation and integration

•		Rock sampling (300 samples) and pit mapping

•		Alteration studies on 1,427 soil samples, 3,591 rock samples and 5,249 core samples

•		Geophysical surveys. 41 km of IP Pole — Dipole. 132 km of ground magnetic readings on a 200m grid

•		Geochemical Survey. 1,482 samples collected for gold and ICP assaying

•		Two-phase diamond drilling program:

	•		Phase 1: 13 diamond drill holes, 3,772m

	•		Phase 2: 40 diamond drill holes, 6,334m

•

Updated mineral resource estimate

9.1.1 2006 Phase 1 Drilling Program

The main aims of the Phase 1 drilling program were to:

•		Identify new mineralization and better define known mineralization with the objective of developing new exploration targets and / or increasing mineral resources

•		Test priority targets for sterilization purposes

Both aims were successfully achieved. Five trends of mineralization were identified with potential to add significantly to mineral resources (Figure 9.1)



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10		DRILLING

10.1 Introduction

Drilling campaigns have been conducted by most of the participating companies in the PVGP over the years. Table 10.1 summarizes all drilling campaigns and the holes / metres used in the Feasibility Study mineral resource estimate.

Table 10.1 Summary of Drilling Campaigns


		Drill	Total Holes		Total Metres		Total Holes		Total Metres
DH Prefix	Company	Type	Included		Included		Excluded		Excluded
AH	Rosario	Rotary		0		0		534		14,368
CU	Rosario	Rotary		0		0		357		9,721
DDH	Rosario	DDH		181		22,966		0		0
DPV06	Barrick	DDH		60		14,710		0		0
GEN_MDD	Genel JV	DDH		11		2,098		0		0
GEN_MNDD	Genel JV	DDH		9		1,053		0		0
GT04	Rosario	DDH		13		1,939		0		0
HA	Rosario	Rotary		0		0		111		2,966
MIM_MN	MIM	DDH		16		2,065		0		0
MIM_MO	MIM	DDH		15		2,535		0		0
MN	Placer	Rotary		2		44		0		0
MO	Placer	Rotary		48		672		0		0
P	Rosario	RC		343		8,706		0		0
PD02	Placer	DDH		19		3,039		0		0
PD04	Placer	DDH		102		13,485		0		0
R	Rosario	Rotary		115		6,571		0		0
RC	Rosario	RC		64		10,002		0		0
RS	Rosario	Rotary		175		24,258		1		138
ST	Rosario	Rotary		551		22,951		79		1,833
SX	Rosario	Rotary		90		1,254		59		769
Totals		Rotary		981		55,750		1,141		29,795
		RC		407		18,708		0		0
		DDH		426		63,891		0		0
		Total		1,814		138,349		1,141		29,795

Figures 10.1 and 10.2 show the locations of drill holes on the property.



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10.2 Pre-Barrick Drilling Campaigns

10.2.1 Rosario Drilling

Rosario employed several drilling methods as summarized in Table 10.2. Geological information was recorded on paper log-forms or graphic logs for all core, reverse circulation (RC), and rotary percussion drill holes.

Table 10.2 Rosario Drill Hole Summary


Drill Hole	Drilling	Total in
Series	Period	Database	Description
P Pre-1975	Pre-1975	343	Pre-production. Shallow percussion holes tested oxide
			mineralization.

R Pre-1975	Pre-1975	57	Pre-production. Shallow rotary holes tested oxide
			mineralization.

MN	Late 1970s	2	Shallow percussion & rotary holes in exploration phase at
			Monte Negro tested oxide mineralization.

DDH < 159	Pre-1975	51	Pre-production. Shallow NQ diamond drill holes in Moore.
			Some angle holes. Mainly tested oxide mineralization.

DDH > 159	1980-91	103	Deeper NQ & PQ diamond drill holes. Vertical holes. Tested
			sulphide mineralization.

RS	1978-90	176	Deeper rotary holes tested sulphide mineralization. Some
			holes removed from resource.

RC	1984-85	64	Deeper reverse circulation holes tested sulphide
			mineralization.

ST	1987-93	552	Shallow rotary holes tested transitional oxide/sulphide
			mineralization.

CU	Early 1980s	252	Shallow rotary holes in Cumba deposit tested oxide
			mineralization.

SX	Post 1990	85	Shallow rotary holes.

MO	Post 1990	48	Shallow rotary holes in Moore deposit drilled to estimate
			recovery prior to excavating.

HA	Not known	64	Shallow rotary exploration holes.

AH	Not known	400	Shallow rotary holes in Arroyo Hondo tested oxide
			mineralization.


Total		2,166

Geology was recorded for deeper holes and for some of the shallow holes. Very few of the shallow holes are relevant to the 2007 mineral resource estimate. No photographs of the core were taken, a common practice in the 1970s and 1980s. The majority of holes were vertical with a drill hole spacing ranging 20m to 80m. No down-hole surveys were performed and 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. This was evaluated by previous operators and it was observed that:



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

•		Silver values were not affected by recovery.

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

Fluor concluded that poor core recovery affected gold grades but in both positive and negative ways. It also concluded that in the context of the whole deposit, statistical noise was apparent but the data was not biased. In AMC’s view, the latter conclusion would need to be underpinned by a study of the relationship between particle size and gold grade. In sulphide deposits, core losses can be small and still result in grade biases.

With respect to rotary and RC drill holes, the previous operator concluded that, with the exception of the P-series RC holes and the RS series of holes below the 250m elevation in the West Flank of the Moore deposit, there was no systematic high bias in RC gold values versus core gold values. Zinc values appeared to be affected by placering in overflowing RC sampling devices, resulting in a low bias in RC holes. As shown in Table 10.2, most of the shallow Rosario holes were drilled in oxide areas now mined out and have only limited, if any, influence on sulphide mineral resource estimates.

10.2.2 GENEL JV Drilling

In 1996, the GENEL JV drilled 20 holes at Pueblo Viejo, eleven in the Moore deposit and nine in the Monte Negro deposit (Table 10.1). Swiss-Boring was contracted to do the drilling using HQ core size. All holes were drilled at an angle. Down-hole surveys were performed but there is no record of the type of instruments used for the surveys. GENEL JV used a GPS system to locate drill holes and to survey the existing pits.

An audit in 2005 was able to verify 5% of the assay data from these holes and found no errors in the database.

10.2.3 MIM Drilling

In late 1996 and into 1997, MIM drilled 31 holes at Pueblo Viejo, 15 in the Moore deposit and 16 in the Monte Negro deposit (Table 10.1). 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 down-hole 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 and so the laboratory that analysed 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 presented as printouts.



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Placer personnel found some of the core, but because of its very poor condition, it was unable to be re-logged.

10.2.4 Historical Drill Hole Surveying

Surveying methods for GENEL JV holes are of accuracy suitable to support resource estimates. The accuracy of collar and down-hole surveys for Rosario and MIM drill holes cannot be confirmed, but from comparisons made between the results of these holes and results from more recent proximal holes of good quality, it has been taken to be sufficiently accurate to support resource estimates

10.2.5 Placer Drilling

Placer completed 3,039 m of core drilling in 18 holes during 2002 and 15,424 m of core drilling in 111 holes during 2004 (Table 10.1). The drilling was undertaken using thin-walled NQ rods that produce NTW (57 mm) core. All but one of the holes were angled. Core was oriented using a down-hole crayon-marking system

Drill pads were located using a GPS or surface plans where the GPS signal was weak. After completion, the drill hole locations were surveyed in UTM coordinates by a professional surveyor, translated into the mine coordinate system (truncated UTM), and entered into the drill hole database.

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 degrees.

Drill holes were logged on paper forms using codes, graphic logs, and geologists’ remarks. Geological information related to assay intervals was recorded on a geology log. A second log was used to record structural information and a third log used to record geotechnical information. Coded data and remarks were key-punched into Excel spreadsheets and edited on site by geology technicians. Coded data were later imported into Gemcom to generate sections for resource modelling.

The following data was recorded on the geological log:

•		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


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The following data was recorded on the structural log:

•		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

The following data was recorded on the geotechnical log (by technicians under the supervision of a geologist):

•		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 greater than 10 cm (RQD)

•		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.

10.3 2006 Barrick Drilling Program

Barrick completed 10,106 m of core drilling in 53 holes during 2006. The drilling was undertaken using thin-walled NQ rods that produce NTW (57 mm) core. Some holes were started on PQ and some holes were reduced to 42 mm. All holes were angled

Surveying and logging procedures were as for the Placer programs, the only difference being that Barrick’s logging was done electronically. AMC viewed Barrick’s core logging procedures during its site visit.

10.4 AMC Opinion

AMC reviewed the descriptions of drilling and related practices, visited drill hole sites and held discussions with site geologists. While the quality of drilling and related practices has varied over the history of the project (further commented on in Sections 11, 12 and 13 of this report), AMC is satisfied that they were undertaken in accordance with standards of the day. Subject to the qualifications expressed in Sections 12 and 13, AMC believes that the results may be relied upon for resource estimation purposes.


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11		SAMPLING METHOD AND APPROACH

11.1 Pre-Placer Drilling Programs

No information is available concerning the sampling strategies used by Rosario during its drilling programs. The record indicates that Rosario generally sampled core on 2m intervals with some samples based on lithology. RC holes were generally sampled on 2m intervals.

The GENEL JV sampled on 2m 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.

From the record, it appears that MIM samples were collected on 2m 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 11.1.

Table 11.1 Sample Interval Data for Rosario, GENEL JV and MIM Drill Holes


		Avg.
		Sample		Min Sample		Max Sample		No.		Avg. Au
Drill Hole		Interval		Interval		Interval		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 JV		2.00		1.40		2.30		520		2.51
MIM	MIM		1.97		0.20		8.00		2,309		2.21

11.2 Placer Diamond Drilling

Placer sample intervals were normally 2m, but were shortened at lithological, structural, or major alteration contacts. Prior to marking the sample intervals, geotechnicians photographed and geotechnically logged the core, then 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.

11.3 Barrick Diamond Drilling

Barrick’s core sampling procedures were (and continue to be) the same as Placer’s as described above, with the exception that 3m samples were used in non-mineralized zones. AMC viewed Barrick’s core sampling procedures during its site visit.

11.4 Sample Quality, Sample Recovery and Related Issues

See Sections 12 and 13 of this report.


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12		SAMPLE PREPARATION, ANALYSES AND SECURITY

12.1 Sample Preparation and Assaying Procedures

No aspect of sample preparation was conducted by an employee, officer, director or associate of Goldcorp.

12.1.1 Rosario

Samples were analysed by fire assay for gold and silver, by LECO combustion furnace for carbon and sulphur, and by atomic absorption (AAS) for copper and zinc. No details are available on crush sizes, sub-sample sizes, or final pulp sample weights used during sample preparation. It was reported in a feasibility study undertaken for Rosario by Stone & Webster International Projects Corporation in 1992 (Stone & Webster, 1992) that the analytical procedures used up to that time were of industry standard.

For the sulphide drilling program that started in 1984, two assay laboratories were present at site, a mainline laboratory responsible for gold, silver, copper, zinc, and iron analyses, and a sulphide laboratory responsible for carbon and sulphur analyses. Sample preparation methods are not documented for this period.

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

12.1.2 GENEL JV

It is inferred from discussions in GENEL JV documents, that samples were prepared on site by GENEL JV personnel. A one-third split of the core was crushed to minus 10 mesh, homogenized by passing through a Gilson splitter three times and sub-sampled to about 400g using a Gilson splitter. The sub-sample was packaged and sent to Chemex Laboratories Ltd. in Vancouver, BC, Canada (Chemex) where presumably the final pulverization was undertaken. In GENEL JV documents, the final pulp grain size is not stated.

Samples were assayed at Chemex for gold, silver, zinc, copper, sulphur and carbon. The procedures are not stated in GENEL JV documentation. A 32 element ICP analysis (G-32 ICP) was performed on each sample.

Security measures utilized by the GENEL JV are not documented.

12.1.3 MIM

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

Core from Rosario, MIM, and GENEL JV drilling was previously stored in inadequate storage facilities, which led to severe oxidation of the remaining core rendering it of limited value.


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12.1.4 Placer

During the 2002 and 2004 programs, drill core was sawn in half with a diamond blade saw at site. The entire second half of the 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 Placer employee. The core samples were sent to Vancouver using airfreight and were received by ALS. No record was kept of the state of the security tags when logged into ALS.

The samples received by ALS were prepared following ALS’s Prep-31 procedure. This included 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 g. The split was pulverized to better than 85% passing 75 µm (200 mesh) and was used for analysis. The remaining sample material (reject sample) was stored at WestCoast Mineral Storage in Aldergrove, BC, Canada.

Samples were assayed for gold, silver, copper, zinc, carbon, sulphur and iron using the analytical techniques listed in Table 12.1. 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.

Table 12.1 Summary of Placer / ALS 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%

All drill core samples from the Placer drilling programs 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 of varying intensities. The results showed that the total carbon analysis is representative of organic carbon in samples with advanced argillic alteration or silicification.

12.1.5 Barrick

The 2006 Barrick drill core was sawn in half with a diamond blade saw at site. The entire second half of core was kept for records and future metallurgical test work. The archived half of the 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


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and sealed with a numbered security tag. The core samples were sent to Vancouver by airfreight and were received by ALS.

Samples were assayed for gold and silver by fire assay and all other elements except carbon and sulphur by multi-element ICP. Sulphur and carbon were assayed by LECO.

12.2 QA/QC Procedures

12.2.1 Rosario Check Assays, 1978

Rosario sent 1,586 samples from ten drill holes to Union Assay Laboratory in Salt Lake City, Utah, USA, for check assays in 1978. The gold check assays exhibited substantial scatter, including several obvious outliers. Some of the scatter may have been due to sample swaps but most of it was unexplained. There was a small bias just outside a reasonable acceptance limit of 5%. Overall, excluding obvious outliers, the data corresponded reasonably well.

The silver data was similar to the gold data in the significant amount of scatter and the large number of outliers. There was a small (5%) bias between the laboratories. Copper exhibited a small amount of scatter and no appreciable bias between the laboratories. Zinc exhibited more scatter than copper but less than gold and silver, although some of the outliers appeared to be sample swaps. There was about a 7% bias between the laboratories (direction of bias not stated).

12.2.2 Rosario Check Assays, 1985

Rosario sent samples to three laboratories in 1985 to validate assays for gold, silver, carbon and sulphur. 392 samples were sent to the Colorado School of Mines Research Institute (CSMRI) for check assaying of the Au and Ag values in three batches. 236 samples were sent to Hazen Laboratories and 154 to AMAX Research and Development Laboratory for sulphur and carbon analysis. Results for these checks have not been located.

AMEC reviewed the CSMRI check data for its 2005 Technical Report for Placer. It reported that gold results generally corresponded well but that there were a number of outliers, possibly caused by sample swaps. AMEC also noted that there was a small bias between the two laboratories; and that the silver results generally agreed well, but, there were again a number of unexplained outliers, some of which were possibly due to sample swaps. The bias between the laboratories was reported by AMEC at about 7% (direction of bias not stated).

12.2.3 GENEL JV Checks

The GENEL JV used a combination of duplicate and standard reference samples (SRMs) to monitor the quality of its assays (Lockhart and Bowen 1997) and restricted its QA/QC to gold. 171 duplicate samples were submitted for analysis representing 11% of total samples. Results compared well with the relative error of the sample pairs at the 90th percentile being about 14%.


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Results of GENEL JV standards STD A, STD B, and STD C showed generally that assays of these standards were within acceptable limits, although it is not clear if sample batches with failures were re-assayed.

12.2.4 Placer Checks, 2002

The Placer 2002 QA/QC program consisted of submitting SRMs into the sample stream as every 20th sample. SRMs were purchased from CDN Resource Laboratories of Delta, BC, Canada, and corresponded to the approximate average gold grade and (the then) cut-off grade for the deposits. Placer did not routinely insert duplicate or blank samples for its 2002 drilling program.

Plots of gold versus batch number showed that the majority of the SRMs returned values within two standard deviations of their established means. 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. This provided a duplicate set of assays, albeit very small and possibly not representative of the whole set, which AMEC reviewed. It observed that the relative error at the 90th percentile was about 23%, somewhat higher than desirable. This may have been due to the sample preparation procedure (see Section 12.1.4), but the very limited data set precludes a firm conclusion.

12.2.5 Placer Checks, 2004

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).

For its 2004 drilling program, Placer inserted an SRM (GS-2, GS-4, and GS-9) and blank (barren limestone) with every batch of 20 samples. All the standards and the blank were assayed for gold, silver, carbon, sulphur, copper, iron and zinc. Gold was the only certified value and the ALS-Chemex gold assays were very close to the certified values indicating that ALS-Chemex generally performed well. Of 25 analyses for GS-2, only one gold value fell outside the pass-fail envelope, with the results for all other elements being within the expected range. Of 174 analyses for GS-4, there were three to six failures for most elements. Of 120 analyses for GS-9, there were only one or two failures per element. The blank (380 analyses) generally returned blank values, ten anomalous values apparently being due to sample mix-up with SRMs.

12.2.6 ALS Chemex Quality Control

ALS conducted analytical quality control in its laboratory by inserting blanks, standards, and duplicates into every sample run, results being reviewed by laboratory staff.

12.2.7 Acme Check Assay Program

As part of the Placer QA/QC program, sample pulps were sent to Acme Laboratories in Vancouver. 187 samples or 13% of the total samples were submitted for the 2002 drill program and were assayed for gold, silver, copper and zinc, while 247 samples were submitted for the 2004 drilling program and were assayed for gold only.


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Results for gold, copper and zinc indicated no significant difference between the two laboratories. However, Acme silver assays were on average about 12% higher than ALS assays. The reason for the difference was not determined as information was not available for SRMs, but it may have been due to differences in the analytical protocols.

12.2.8 Barrick Checks, 2006

The QA/QC procedure used for the 2006 drilling program consisted of the introduction of blanks, commercial SRMs for gold, and core duplicates into the sampling process.

•		Each batch was submitted with 75 samples, of which six were QC control (two blanks, two standards, and two core duplicates)

•		The blanks were a local limestone, the same used during the Placer campaign

•		The SRMs were purchased from Rocklab of New Zealand and correspond to the gold range of the deposit. These standards were used for most of the year

•		Five custom, reference materials were prepared by Barrick using the mineralization from Pueblo Viejo. The gold grade range corresponds to the range of the deposits

•		Core duplicates were also inserted approximately every 30^th sample

•		QA/QC results for 2006/07 drilling are summarized in Figures 12.1 to 12.5

Figure 12.1 Barrick 2006 / 07 QA/QC Results — Blanks


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Figure 12.5 Barrick 2006 / 07 QA/QC Results — Sample Grain size

•		Each sample is weighted

•		Granulometric control is done on 3% of the samples at the crusher stage and 3% on the pulverization stage.

For check samples that fell outside the control limits, Barrick examined the cause and, if it could not be attributed to sample mix-up, had the relevant batch re-assayed.

12.3 Summary

No information is available on sampling or sample preparation procedures for Rosario or MIM drilling programs. From limited information, sampling by the GENEL JV appears to have been consistent with common practice of the day. Sampling by Placer was performed to acceptable standards.

Sample preparation by both the GENEL JV and Placer involved sub-sampling at the crushing stage. Sub-sampling of samples prior to complete pulverisation is generally not recommended for gold deposits. The Pueblo Viejo deposits are characterized by very fine gold with no visible gold having been recorded, which somewhat reduces the risk associated with sub-sampling prior to complete pulverization. However, the presence of very fine gold does not preclude bias during sub-sampling as the gold can be preferentially located in fine (or coarse) particles.

QA/QC procedures have varied significantly during the history of work at Pueblo Viejo. In its 2005 Technical Report, AMEC noted inadequacies in the QA/QC data for the Rosario and MIM data, with which AMC agrees. In order to confirm the adequacy of the sample data for these programs, AMEC and Placer compared Rosario and MIM holes with nearby Placer



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13		DATA VERIFICATION

13.1 Verification of Pre-Placer Data

13.1.1 Database Development

American Mine Services (AMS), as part of the 1992 Stone & Webster (1992) study, 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 JV and MIM in 1995 and subsequently acquired by Placer. Placer compared the GENEL JV database with that provided by Rosario and confirmed that only minor changes had been made since AMS’s validation exercise. The changes were corrected based on original Rosario assay sheets and drill logs at the Pueblo Viejo site.

Placer 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 Placer report, “Report on the Comparison of the PDI02 and GENEL98 Drill Hole Databases for the Pueblo Viejo Project, Dominican Republic” (February 2003).

13.2 Rosario Pseudo-Twin Assay Pairing

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

AMEC constructed declustered QQ plots and confirmed the conclusions made by Placer. In grade ranges below 2 g/t, Rosario drilling appears to be biased slightly high compared to Placer drilling; but above 2 g/t and below 6 g/t the assays compare reasonably well. Above 6 g/t, Rosario drilling appears generally biased slightly high relative to Placer drilling.


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Figure 13.1 AMEC Comparison of Placer and Rosario Drill Hole Assays within 10 m

13.2.1 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 times thickness). Fluor concluded that analysis of gold, silver and sulphur results showed no significant overall bias but that “carbon assays were consistently lower by 7% and zinc assays lower on average by 36% than in the original hole.” One hole, RS-40, was removed from the resource estimation database because it appeared to have been drilled down a near-vertical mineralised structure.

Placer reviewed 20 twinned and closely spaced drill holes and compared the gold grades using a profile plot and a scatter plot. The results show good agreement between the different drilling methods, rotary, reverse circulation, and core, when the holes were closely spaced with the exception of some rotary holes that appeared to show down-hole contamination.

AMEC reviewed the database and identified 40 holes that were twinned in part or in whole by Rosario, GENEL JV, and MIM. None of the 2002 or 2004 Placer holes were twins of previously drilled holes. AMEC concluded that there was a wide divergence in comparisons between twinned holes that allowed no simple conclusions to be drawn. AMEC also observed that there is a tendency for RC holes to return somewhat higher grades and metal contents than core holes, due possibly in part to localized, down-hole contamination in the RC holes; and that 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. AMEC concluded that this probably explains many of the differences observed between twin holes.


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13.3 Verification of Pre-Barrick Data

13.3.1 Verification of Placer Data

AMEC compared one in 20 samples in the Placer 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.

13.3.2 Down-Hole Contamination of RC and Rotary Holes

AMEC investigated the possibility of down-hole contamination in RC and rotary drill holes. It concluded that 59 holes showed greater or lesser degrees of possible down-hole contamination:

•		9 Rosario RC holes (RC series)

•		16 Rosario rotary holes (RS series)

•		34 Rosario rotary holes (ST series)

13.3.3 Cross Sectional Review of MIM, Rosario, and Placer Drilling

Barrick visually reviewed assays for MIM, GENEL JV, Rosario, and Placer drilling in those parts of the Moore and Monte Negro deposits where the holes cross. In general, it found reasonable agreement of the orientation, tenor, and thickness of mineralization between drilling campaigns.

13.3.4 Gold-Grade Distribution Comparisons

Barrick used gold-grade histograms of the historical drilling campaigns, each compared to a histogram of gold assays from all drilling, to identify those campaigns with unacceptable gold grade biases. The comparisons were broken out by company and drilling type. Only the drill holes used for the resources estimate were considered.

The histograms show that the diamond core drilling from all campaigns compare well with the global distribution, with the exception of the Placer rotary holes (Figure 13.2) which are biased high and were possibly preferentially drilled in shallow high grade areas to better delineate early production. In AMC’s opinion, information from these holes should have been removed from the data base but this does not constitute a material issue to the project. The Barrick gold grades appear to be biased low (Figure 13.3), but this drilling was targeted at the periphery of the deposits so that lower overall grades would be expected.


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Figure 13.2		Frequency Distribution of Gold by Drilling Campaign: All Drill Holes vs. Placer Rotary


						Untransformed
						gold Statistics															Log Normal
	gold Cutoff					gold Cutoff					gold Cutoff					gold Cutoff					Approximation Model
	= 0.01 g/t					= 0.50 g/t					= 1.00 g/t					= 3.00 g/t							Standard		Third
	Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Mean		Deviation		Parameter

All Drill holes		100,895			2.392		84,646			2.798		66,387			3.376		24,935			6.025		0.40		1.00		0.00
incr. % and grade		16.1	%		0.274		18.1	%		0.697		41.1	%		1.783		24.7	%		6.025
Placer Rotary		546			3.546		487			3.946		454			4.177		234			6.098
incr. % and grade		10.8	%		0.244		6.0	%		0.773		40.3	%		2.134		42.9	%		6.098


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Figure 13.3 Frequency Distribution of Gold by Drilling Campaign: All Drill Holes vs. Barrick DDH


						Untransformed															Log Normal
	gold Cutoff					gold statistics					gold Cutoff					gold Cutoff					Approximation Model
	= 0.01 g/t					gold Cutoff = 0.50 g/t					= 1.00 g/t					= 3.00 g/t							Standard		Third
	Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Mean		Deviation		Parameter

All Drill holes		100,895			2.392		84,646			2.798		66,387			3.376		24,935			6.025		0.40		1.00		0.00
incr. % and grade		16.1	%		0.274		18.1	%		0.697		41.1	%		1.783		24.7	%		6.025
Barrick DDH		8,231			1.587		5,917			2.109		3,931			2.810		1,047			5.830
incr. % and grade		28.1	%		0.252		24.1	%		0.721		35.0	%		1.714		12.7	%		5.830


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13.3.5 Summary

Extensive evaluations of the possible bias introduced by various drilling procedures were undertaken by Fluor, Pincock, Allen & Holt (which reviewed a 1997 mineral resource estimate), Placer and AMEC and, more recently, by Barrick. AMC has also undertaken limited checks of database information against original data, has reviewed cross-sectional plots of drilling information and has reviewed checks and audits carried out by other parties. The following conclusions may be drawn:

•		Approximately 2.5% of the Rosario data (which comprises the largest proportion of the drilling database) have been verified against original documents. The Rosario core, RC and some rotary data are generally reliable and those that are considered to be of questionable validity have not been used for the 2007 mineral resource estimate. As noted earlier, most of the shallow Rosario drill holes were drilled in oxide areas now mined out and have limited, if any, influence on sulphide mineral resource estimates.

•		GENEL JV data has been verified against original documents and are believed to be reliable.

•		MIM data has not been verified against original documents and there is some risk involved with using that data. On the basis of comparisons between mineralized intersections in MIM holes and those in nearby Placer holes, the risk of using the MIM data is considered to be acceptable.

•		Placer data has been verified against original documents and is considered to be reliable.

•		Some Barrick data has been verified by AMC against original documents and is considered to be reliable.

13.3.6 AMC Opinion

In AMC’s view, appropriate efforts have been made to ensure that the Pueblo Viejo drilling database is free from major defects and of an acceptable quality to support feasibility study mineral resource estimates. It is believed that any remaining deficiencies will not materially affect global resource estimates, but may impact in places on local estimates. AMC recommends that attention continue to be paid to the quality of historic drilling information, with targeted replacement drilling being undertaken where necessary.



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14		MINERAL RESOURCE ESTIMATES

14.1 Introduction

The resource model created by Placer in 2005 provided the starting point for the 2007 Barrick FS mineral resource estimate and so details of this model are presented before describing the Barrick estimate.

14.2 2005 Placer Mineral Resource Estimate

14.2.1 Introduction

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 modelled: gold, silver and sulphur.

The rate at which ore can be processed in autoclaves is directly proportional to the sulphur content, which makes sulphur assays critical to the mine plan and ultimately the cash flow. At the time of the 2005 Placer resource estimate, a significant number of sulphur determinations were missing for samples in the Moore deposit. In order to fill in the missing data, a simple regression formula between gold and sulphur was developed for each geological domain. AMC has been unable to gain verification of the robustness of the regression formula but believes this issue not to be sufficiently material to affect project viability.. 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 Placer in-house software OP and GEOLOG for statistical analysis, variography, and kriging; Maptek’s Vulcan for 3-D geological interpretation and visualization; and SAGE2001, by Ed Isaaks for calculating variograms.

14.2.2 Drill Hole Database

A total of 80,261 m of drilling in 573 holes was used for resource estimation (Table 14.1).

Table 14.1 Drill Holes and Metres used for 2005 Placer Resource Estimate


	Moore						Monte Negro						Total
					Avg						Avg						Avg
	No. of				hole		No. of				hole		No. of				hole
Type	Holes		Metres		length		Holes		Metres		length		Holes		Metres		length
Core		185		27,271		147.4		129		18,231		141.3		314		45,502		144.9
RC		144		17,416		120.9		51		7,342		144.0		195		24,757		127.0
Rotary		20		2,821		141.4		44		7,181		163.2		64		10,002		156.3
Total		349		47,507				224		32,754				573		80,261



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Comparison of average hole lengths in Table 14.1 with data in Table 14.6 leads AMC to the conclusion that there are typographical errors in Table 14.1.

Those RC and rotary drill holes judged to be of questionable reliability, including those identified by AMEC as potentially having down-hole contamination, were removed from the resource estimation database.

14.2.3 Geological Modelling

Geological models for lithology, structure and alteration were produced in two- and three dimensions for both the Moore and the Monte Negro deposits. These were then superimposed upon each other to define the geological domains. The interpretations were wire-framed to create a three-dimensional model and the block model and drilling composites coded using the wire-framed model.

14.2.3.1 Moore

Placer defined seven lithological units for Moore (Table 14.2 and Figure 14.1). These units were then further subdivided into five structural domains, primarily defined by faults. The structural domains were defined by Placer using surface mapping and data from oriented core. To further define the mineralization controls, Placer interpreted the contact between advanced argillic 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 14.2). 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.

Placer investigated the trend of grades across boundaries between different units with three different methods: contact plots, cross-correlograms, and geo-bound plots. As a 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.



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Table 14.2 Moore Deposit Zone Names, Placer 2005 Estimate


										Block
Domain		Names	Litho		Struct		Alter		Description	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



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Figure 14.1 Moore Section 94600 N, Lithology and Structural Domains

14.2.3.2 Monte Negro

Placer 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, three of which contain significant mineralization. The procedures to build the Monte Negro alteration and vein models were 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 lithological domains produces 15 defined statistical zones. The Monte Negro zones are listed in Table 14.3. A typical Monte Negro cross section is shown in Figure 14.2.

Table 14.3 Monte Negro Deposit Zone Names, Placer 2005 Estimate


										Block
Domain		Names	Litho		Struct		Alter		Description	Count
	1	WBF	all			9		1	Western Boundary Fault		47,697
	2	WMF-MNS		1		1		2	West Monte Negro Fault- Monte Negro Sediments		47,578
	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 Sediments		20,314
	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



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										Block
Domain		Names	Litho		Struct		Alter		Description	Count
	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 14.2 Monte Negro Section 95800N Lithology and Structural Domains

14.2.4 Topography

The topographic surface for the model was based on a 1997 / 1998 set of aerial photographs taken by the GENEL JV. The available contours were at 4m intervals.



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14.2.5 Bulk Density

Bulk densities at Pueblo Viejo have been derived from a linear regression formula based on 152 pairs of density and sulphur samples calculated by AMAX Engineering and Mining Services from the GR series of diamond drill holes in 1985. The formula is:

Density = (0.0322 * sulphur %) + 2.617

Fluor checked the regression equation during its 1986 feasibility study. It used the equation to estimate the 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 density. Fluor noted 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 density 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 JV 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, BC, for sulphur analysis. 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. Placer accepted the bulk density formula.

14.2.6 Data Compositing

The assay data was combined into 2m composites. More than 95% of the samples are between 2m and 3m in length, and 82% are exactly 2m long.

14.2.7 Top Cutting

Since ordinary kriging was used for all grade estimates, it was necessary to examine the assay distributions to assess 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 that include an indicator correlation graph, a coefficient of variation graph, a contained metal graph, and decile analysis graphs.

The procedure used to select a cutting threshold was to examine each graph independently and, based on that graph alone, make an appropriate selection as to 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 gold composites grades were cut to the selected cutting limit while 112 of 21,853 silver composites were cut and 140 of 16,179 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. The number of total composites



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adjusted by cutting is low and did not significantly affect the global mean grades of gold, silver, and sulphur.

Within Monte Negro, a total of 92 out of 14,453 gold composites 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.

14.2.8 Variogram Analysis

Placer examined the spatial continuity of the gold, silver, and sulphur using correlogram maps and directional correlograms. Placer found it necessary to group 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. Placer then confirmed the fitted models with its directional variogram software.

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

14.2.9 Interpolation Plans

The basic interpolation plan was similar for gold, silver, and sulphur for both Moore and Monte Negro. Block size was set to 10m by 10m by 10m. A minimum of 8 and a maximum of 16 2m composites were used. 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.

14.2.10 Model Validation

Placer’s validation of the gold, silver, sulphur and calculated sulphur grade models consisted of five different checks:

•		Global mean check

•		Global variability check

•		Trend check/swath plots

•		Comparison with ST holes

•		Visual inspection

The global mean check compared the ordinary kriged estimate with a nearest neighbour estimate, which will provide a declustered estimate of the global mean. In all cases, the resource estimate appeared 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 with the variance of the estimated blocks. It is desirable that a degree of smoothing



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be imposed on the model such that the resource estimate will represent the distribution of tonnes and grade that would be expected at the cut-off 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 indicated that there had 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 modelling 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 appeared to fairly represent the drilling.

14.2.11 Resource Classification

Resource classification was based on the kriging estimation variance from the gold estimation. Placer chose the estimation variance threshold to try to classify interpolated material as Indicated. The classification statistics are shown in Table 14.4. 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.

Table 14.4 Resource Classification and Estimation Statistics, 2005 Placer Estimate


					Avg. krig		Avg		Avg no.		Avg no.		Avg no.
	Class		No. Blocks		variance		distance		composite		holes		octants
Measured
Moore		1		42,367		0.33		27.2		12.7		2.4		3.9
Monte Negro		1		25,361		0.24		25.0		13.4		2.4		4.2
Indicated
Moore		2		14,562		0.60		43.6		11.9		2.4		3.5
Monte Negro		2		5,911		0.51		39.7		10.7		2.0		3.3
Inferred
Moore		3		1,528		0.81		63.6		9.7		1.7		2.7
Monte Negro		3		938		0.78		55.1		8.2		1.2		2.3

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.



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14.2.12 2005 Placer Mineral Resource Estimation Results

The resource models were used to generate pit shells using a $450 /troy oz gold price, within which the resource estimates were contained. The 2005 Placer mineral resource estimate is summarized in Table 14.5.

Table 14.5 2005 Placer Resource Summary at 1.7 g/t Au Cut-off Grade (100% Basis)


Category	Tonnes (M)		Gold (g/t)		Gold (Moz)		Silver (g/t)		Silver (Moz)
Measured		118.6		3.2		12.3		18.1		69.2
Indicated		32.0		2.9		3.0		13.3		13.7
Measured + Indicated		150.6		3.2		15.3		17.1		82.8
Inferred		2.2		2.9		0.2		12.6		0.9

14.3 2007 Barrick Feasibility Study Mineral Resource Estimate

14.3.1 Introduction

The 2007 Barrick FS resource estimate was prepared by Barrick’s Tucson-based technical team following a site visit in January 2007 and using methods typical of all Barrick major projects. Cut-off dates of January 22, 2007 and 30 June 2007 were used for new gold and sulphur sample data respectively, at which time 48 additional diamond core holes had been completed by Barrick.

14.3.2 Geological Model

At the time of preparation of the 2007 estimate, site geologists were working on a new geological interpretation, although it was not available in a format that would have enabled the creation of a full 3-D model.

The new model uses brecciated feeders to explain the higher grade mineralization. At depth, those feeder zones are steeply-dipping and seem to be oriented similarly to the local structure, striking north-north-west for Monte Negro and almost due north for Moore. Nearing the surface, the breccias seem to sill out, almost flattening. In Moore, these flatter zones tend to follow lithology bedding, which dips west about 20°, while in Monte Negro, it seems to have a plunge of 10° to the south.

Since no geological solids were available to define these feeders, a probability indicator with a cut-off of 5 g/t was utilized to identify the extent of the higher gold grades. A second probability indicator at 1.0 g/t was used to separate the higher and lower grade mineralization. This cut-off was defined using a cumulative frequency plot as described below.

14.3.3 Drill Hole Database

A summary of drill holes used for the 2007 Barrick FS resource estimate is presented in Table 14.6, with more details, including excluded holes, provided in Table 10.1. A total of 138,349m of drilling in 1,814 holes was used for resource estimation.



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Table14.6 Drill Holes and Metres used for 2007 Barrick Resource Estimate

Moore

Monte Negro

Total

Avg

No. of

hole

No. of

hole

No. of

hole

Type

Holes

Metres

Length

Holes

Metres

length

Holes

Metres

length

Core

426

63,891

150.0

497

18,708

37.6

Rotary

981

55,750

56.8

Total

1,814

138,349

AMC has been unable to obtain a further breakdown of Table 14.6.

For the 2007 estimate, the excluded drill holes were consistent with those for the 2005 resource estimates, with the exception of the Rosario RC drilling. Many of the RC holes eliminated for the 2005 resource estimate on the basis of AMEC’s review of possible down-hole contamination were replaced by Barrick for the 2007 work. Barrick evaluated the effect of replacing them by comparing two resource models. One model used the same drilling selected in 2005 combined with the Barrick drilling. The other model added the RC holes eliminated in 2005. The gold grade difference between these two models was negligible, while eliminating the RC holes created gaps in the drill spacing resulting in a 3% reduction in tonnage.

14.3.4 AMC Opinion

In AMC’s view, evaluation of the impact of replacing the Rosario RC holes in the resource estimation database should have been undertaken on a local as well as global basis, as a global non-bias can mask significant local biases. However, on the basis of comparisons undertaken by Barrick using gold-grade histograms of historical drilling campaigns, each plotted against a histogram of gold assays from all drilling (Figure 14.3), any risk arising from re-inclusion of Rosario RC holes is not likely to be material to the project.



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14.3.6 Coordinate Units

Although the Project uses UTM coordinates, much of Placer’s previous work, as well as some of the consultants’ work, is in cropped or local units by trimming 300,000m from the easting and 2,000,000m from the northing.

14.3.7 Raw Assay Statistics

Drill hole assay data was loaded into a Vulcan database and gold-grade statistics compiled by rock type. The rock types were taken from the geologic log information provided in the database. The statistics are summarized in Table 14.7. The table shows that the gold grades at Pueblo Viejo are fairly well behaved, with coefficients of variation near 1 at the lowest cut-off grades. The table also shows that, while the bulk of the mineralization is hosted in volcanic rocks and breccias, a large proportion of the drilling and more than half of the mineralized intervals, have no logged rock type. Geologic logs were available mainly for Barrick, Placer and Rosario core only.

Gold-grade histograms and cumulative frequency plots were created for the important rock types. Barrick identified a strong break in the cumulative frequency distributions in the gold-grade range 0.2 g/t to 0.7 g/t (see Figure 14.4). The unlogged intervals have very few low-grade samples.

Most of the rock types contain significant mineralized components. One exception is the group of intrusive dykes that contain little or no mineralization. These are andesitic in composition and occur mainly in the Monte Negro area.

14.3.8 AMC Comment

AMC is unable to recognize the strong break in the cumulative frequency plot in the range 0.2 g/t to 0.7 g/t identified by Barrick. A break occurs around 0.3 g/t, but this likely reflects the precision with which very low grades can be assayed, rather than differences in geologic zones.


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Table 14.7 Gold Assay Statistics


								Statistics Above Cutoff										Capped Statistics Above Cutoff
	Au							Mean		Grade								Mean		Grade
	Cutoff		Total		Inc.			Au		Thickness		Inc.		Std.		Coeff. of		Au		Thickness		Percent		Std.		Coeff. of
	(g/t)		(m)		(%)			(g/t)		(g/t-m)		(%)		Dev.		Variation		(g/t)		(g/t-m)		of Total		Dev.		Variation
Total of all		0.0		110,543		23.4			2.2		241,470		1.9		3.2		1.5		2.1		236,498		97.9		2.7		1.3
Groups		0.5		84,646		16.5			2.8		236,853		5.3		3.5		1.2		2.7		231,881		97.9		2.8		1.0
below		1.0		66,387		50.1			3.4		224,134		52.4		3.7		1.1		3.3		219,162		97.8		2.9		0.9
		5.0		10,994		9.9			8.9		97,615		40.4		6.4		0.7		8.4		92,643		94.9		3.8		0.5

Spilite		0.0		9,359		23.5			2.0		18,942		2.2		2.9		1.4		2.0		18,655		98.5		2.6		1.3
		0.5		7,163		21.0			2.6		18,525		7.5		3.1		1.2		2.6		18,238		98.5		2.7		1.1
		1.0		5,197		46.4			3.3		17,106		50.9		3.3		1.0		3.2		16,819		98.3		2.9		0.9
		5.0		855		9.1			8.7		7,457		39.4		5.3		0.6		8.4		7,170		96.2		3.8		0.5

Volcanics		0.0		31,365		32.7			1.9		59,855		2.2		3.3		1.7		1.9		58,156		97.2		2.7		1.5
		0.5		21,106		16.6			2.8		58,565		6.2		3.8		1.4		2.7		56,867		97.1		2.9		1.1
		1.0		15,897		42.2			3.5		54,856		49.9		4.1		1.2		3.3		53,158		96.9		3.1		0.9
		5.0		2,649		8.4			9.4		25,017		41.8		7.3		0.8		8.8		23,318		93.2		4.1		0.5

Pyroclast-		0.0		2,219		66.0			0.6		1,371		14.2		1.2		1.9		0.6		1,371		100.0		1.2		1.9
ics		0.5		754		17.2			1.6		1,176		19.1		1.6		1.0		1.6		1,176		100.0		1.6		1.0
		1.0		372		15.3			2.5		914		49.1		1.9		0.8		2.5		914		100.0		1.9		0.8
		5.0		33		1.5			7.3		241		17.6		2.1		0.3		7.3		241		100.0		2.1		0.3
Intrusive		0.0		1,316		76.4			0.6		800		9.6		2.2		3.6		0.6		748		93.6		1.6		2.8
Dykes		0.5		310		9.1			2.3		723		10.2		4.0		1.7		2.2		672		92.9		2.6		1.2
		1.0		191		12.3			3.4		642		43.1		4.9		1.5		3.1		590		92.0		3.0		1.0
		5.0		29		2.2			10.3		297		37.1		9.8		1.0		8.5		245		82.7		4.5		0.5

Sediments		0.0		29		0.0			0.9		25		0.0		0.3		0.3		0.9		25		100.0		0.3		0.3
		0.5		29		58.4			0.9		25		46.3		0.3		0.3		0.9		25		100.0		0.3		0.3
		1.0		12		41.6			1.1		14		53.7		0.2		0.2		1.1		14		100.0		0.2		0.2
		5.0		0		0.0			0.0		0		0.0		0.3		0.0		0.0		0		0.0		0.3		0.0

Carbon-		0.0		897		50.5			1.3		1,160		2.5		2.7		2.1		1.3		1,134		97.7		2.4		1.9
Aceous		0.5		444		16.3			2.6		1,132		7.8		3.4		1.3		2.5		1,105		97.7		3.0		1.2
Sediments		1.0		298		26.6			3.5		1,041		44.4		3.8		1.1		3.4		1,014		97.5		3.3		1.0
		5.0		60		66	%		8.8		526		45.3		5.7		0.7		8.4		500		95.0		4.2		0.5

Breccias		0.0		1,009		26.1			2.1		2,073		2.6		2.8		1.4		2.1		2,070		99.9		2.8		1.4
		0.5		746		19.7			2.7		2,018		6.6		3.0		1.1		2.7		2,016		99.9		3.0		1.1
		1.0		547		44.3			3.4		1,880		47.4		3.2		0.9		3.4		1,878		99.9		3.2		0.9
		5.0		100		9.9			9.0		899		43.4		3.6		0.4		9.0		896		99.7		3.5		0.4

Hydro-		0.0		1,120		42.7			1.6		1,766		2.8		3.1		2.0		1.5		1,707		96.7		2.6		1.7
thermal		0.5		642		13.0			2.7		1,716		6.1		3.7		1.4		2.5		1,657		96.6		3.0		1.2
Breccias.		1.0		497		37.9			3.2		1,609		49.9		4.0		1.2		3.1		1,550		96.3		3.3		1.0
		5.0		72		6.4			10.1		727		41.2		7.2		0.7		9.3		668		91.9		4.8		0.5

Phreato-		0.0		1,606		30.0			1.3		2,083		5.7		1.7		1.3		1.3		2,083		100.0		1.7		1.3
Magmat		0.5		1,125		29.5			1.8		1,966		15.3		1.8		1.1		1.8		1,966		100.0		1.9		1.1
Breccias.		1.0		651		36.0			2.5		1,646		52.9		2.1		0.8		2.5		1,646		100.0		2.1		0.8
		5.0		72		4.5			7.6		545		26.1		2.4		0.3		7.6		545		100.0		2.4		0.3

Faults		0.0		36		33.2			2.8		104		0.7		3.8		1.3		2.8		104		100.0		3.8		1.3
		0.5		24		0.0			4.2		103		0.0		4.0		1.0		4.2		103		100.0		4.0		1.0
		1.0		24		50.3			4.2		103		44.2		4.0		1.0		4.2		103		100.0		4.0		1.0
		5.0		6		16.5			9.5		57		55.1		4.8		0.5		9.5		57		100.0		4.8		0.5

OVBD		0.0		179		46.6			1.6		277		3.7		1.9		1.2		1.6		277		100.0		1.9		1.2
		0.5		96		8.0			2.8		267		3.9		1.9		0.7		2.8		267		100.0		1.9		0.7
		1.0		81		38.7			3.2		256		64.5		1.8		0.6		3.2		256		100.0		1.8		0.6
		5.0		12		6.7			6.4		77		27.9		1.3		0.2		6.4		77		100.0		1.3		0.2

Unclassified		0.0		61,408		15.0			2.5		153,015		1.6		3.3		1.3		2.5		150,167		98.1		2.7		1.1
		0.5		52,209		15.6			2.9		150,637		4.3		3.4		1.2		2.8		147,790		98.1		2.8		1.0
		1.0		42,621		57.8			3.4		144,068		53.8		3.6		1.1		3.3		141,220		98.0		2.8		0.9
		5.0		7,108		11.6			8.7		61,772		40.4		6.3		0.7		8.3		58,925		95.4		3.7		0.4


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Figure 14.4 Frequency Distribution of Raw Gold Assays for All Rock Types


						Untransformed gold Statistics															Log Normal Approximation Model
	gold Cutoff = 0.01 g/t					gold Cutoff = 0.50 g/t					gold Cutoff = 5.00 g/t					gold Cutoff = 80.00 g/t							Standard		Third
	Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Mean		Deviation		Parameter

all rock types		110,543			2.184		84,646			2.798		10,994			8.879		9			118.107		0.15		1.05		0.00
incr. % and grade		23.4	%		0.178		66.6	%		1.890		9.9	%		8.789		0.0	%		118.107		0.13

14.3.9 Top Cutting

The cumulative frequency curves generated from the raw assays were used to determine cuts for the gold grades. A single cut was applied regardless of rock type. Some differences were seen in grade populations separated by logged lithology, but because of the large proportion of unlogged intervals, it was decided to use a global grade cut for all gold assays.

The global gold grade cumulative frequency plot and cutting statistics are shown in Figure 14.5. A gold-grade cut of 20 g/t was determined. A total sample length of 435m was capped, representing about 0.4% of the total assay population. While the cut removed only 2% of the total grade thickness, it reduced the coefficient of variation in the assay


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population from 1.48 to 1.22. The cut was applied to all raw assays prior to creating composites. The effect of the cut is summarized by rock type in Table 14.7.

Figure 14.5 Raw Gold Assay Cutting


						Untransformed gold Statistics															Log Normal Approximation Model
	gold Cutoff = 0.01 oz/t					gold Cutoff = 0.20 oz/t					gold Cutoff = 0.50 oz/t					gold Cutoff = 2.50 oz/t							Standard		Third
	Meters			Au (oz/t)		Meters Au			(oz/t)		Meters			Au (oz/t)		Meters			Au (oz/t)		Mean		Deviation		Parameter

raw assays		110,543			2.184		97,257			2.477		84,646			2.798		31,448			5.335		0.25		1.02		0.00
incr. % and grade		12.0	%		0.044		11.4	%		0.319		48.1	%		1.299		28.4	%		5.335

low cut		0.010				20.0 oz/t percentile					GT lost by capping					percent of GT >= 187.7 oz/t

						99.61%					2.06%					0.16%
Au cap (topcut)		20.00				percent of GT >= 20 g/mt					CV uncapped					CV capped

						5.66%					1.48					1.22

14.3.10 AMC Comment

AMC notes that around 50% of the samples are below 1 g/t Au. The inclusion of this large number of samples representing extensive, continuous zones of very low grades has somewhat masked the sample statistics and cumulative frequency plot. It would also have been preferable if Monte Negro and Moore had been studied separately. There is a slight risk in cutting high values on the basis of un-weighted (for length) raw assays rather than on weighted raw assays or composites. However, since over 95% of samples are between 2m and 3m in length, any impact should not be material.

14.3.11 Assay Compositing

Raw drill hole gold assays were capped and then grouped into composite assays measuring 10m in length down the hole. The composite length was chosen to match the planned bench height for mining. Only those drill holes selected for the resource estimate were used for composites. Composites were not broken at lithologic contacts since the geologic logging is not complete.


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Down-hole composites measuring 2m to 20m were created to determine the effect of composite length on the extent and grade of mineralization. Total composite length, grade and grade-thickness above cut-off were plotted for various composite lengths over a range of gold cut-off grades (Figure 14.6).

Figure 14.6 Gold Assay Statistics at Varying Composite Lengths

The plots show that in the operational cut-off grade range (0.5 g/t to 2.0 g/t), composite grade and total thickness above cut-off varies by 10% to 15% over the composite lengths plotted. The grade thickness does not change appreciably over the operational cut-off grade range.

The barren, intrusive andesitic dykes identified in the Monte Negro area were not separated from the composites but were allowed to dilute the overall composite grade. Final 10m


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composite gold grades were compared to raw assay grades to ensure that the compositing did not introduce any grade bias (Figure 14.7).

Figure 14.7 Frequency Distribution of Gold Grades in 10m Composites vs. Raw Assays


						Untransformed Gold Statistics															Log Normal Approximation Model
	Gold Cutoff = 0.01 g/t					Gold Cutoff = 0.50 g/t					Gold Cutoff = 5.00 g/t					Gold Cutoff = 80.00 g/t									Standard Third
	Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Mean		Deviation		Parameter

Raw Assays		110,543			2.184		84,646			2.798		10,994			8.879		9			118.107		0.15		1.05		0.00
incr. % and grade		23.4	%		0.178		66.6	%		1.890		9.9	%		8.789		0.0	%		118.107		0.13
10m Composites		118,719			2.034		89,212			2.650		9,947			7.801		0			0.000
incr. % and grade		24.9	%		0.172		66.8	%		2.003		8.4	%		7.801		0.0	%		0.000		0.11

14.3.12 Geological Solids and Model

Geological solids were created by site geologists by simple extrapolation from section centreline, but when checking the logged lithology codes against the solids, too many mismatches occurred for the solid shapes to be used.


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The only solids used to define geology were the interpolated, intrusive andesitic dykes, including several previously unidentified in the Monte Negro area. Although the dyke geometry was inaccurate due to extrapolating centreline dyke shapes halfway to the next section, the dykes were volumetrically accounted for and allowed to dilute the overall block grades.

At the time of resource estimate preparation, a structural model was being developed and only surface traces were available, with general strike and dip. Reviewing of assays in three dimensions permitted the definition of simple fault blocks for separating mineralized and un-mineralized areas.

14.3.13 AMC Opinion

AMC believes that the lack of any meaningful geologic zoning represents the greatest weakness in the model and that more time should have been spent in defining broad geologic zones before undertaking the geostatistical study. In AMC’s opinion, the indicator kriging method is a relatively poor substitute for geologic zoning

14.3.14 Block Model

A block model was generated, covering the extent of the 2005 model. A single model was defined, encompassing both Moore and Monte Negro areas. Blocks size was set at 10 m by 10 m by 10 m and the model was not rotated. Table 14.8 shows the block model geometry.

Table 14.8 Block Model Geometry


	X		Y		Z
	(m)		(m)		(m)
Minimum		374,300		2,093,900		-100
Maximum		376,900		2,096,700		500
Extent		2,600		2,800		600
Block Size		10.0		10.0		10.0

The andesitic dykes discussed above are often quite narrow. Flagging the centroid of 10m blocks would leave the dyke mass under-represented. A 1m block model was therefore created using the site defined shapes. The 1m model was then regularized into a 10m model, resulting in each block containing a percentage of dyke material. This dyke percentage was transferred to the main 10m model.

14.3.15 Bulk Density

All bulk density values were taken directly from the Placer 2005 model (see Section 14.2.5 of this report). AMC has been unable to determine whether the bulk density values took account of the additional sulphur determinations available by mid-2007, but does not believe this to be an issue of material significance to the project.


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Figure 14.9		Omni-Directional Correlogram Gold

14.3.17 Gold Grade Estimation

Three major estimation domains were defined for gold estimate: Moore, Monte Negro, and a Low Grade structural zone to the west of Moore and south of Monte Negro. Each domain accounts for 41%, 50%, and 9% of all blocks respectively. Preferential directions of continuity were defined for Moore and Monte Negro.

In the absence of robust geological domaining, a set of two discriminators or probability indicators was generated. The first indicator, 5.0 g/t Au, served to isolate the higher grade population interpreted to be associated with hydrothermal breccias (possibly feeders) that are steeply-dipping at depth and tend to flatten out and follow bedding near the surface. The second indicator, 1.0 g/t Au, was used to separate the two populations marked by a slight change in slope on a cumulative frequency curve (Figure14.4). Sectional interpretations showed these cut-off grades to be geologically reasonable.

All 10m composites were assigned either 1, 0 or -9, depending on the composite gold value being greater than or equal to the indicator grade, less than the indicator grade, or not available, respectively. The 0 and 1 indicators were then estimated by domains using inverse distance squared (ID²). A minimum of five and maximum of 13 composites were used, and a maximum of two composites per hole. This condition ensured that at least three holes were within the search range for a block to be estimated. Only composites within the same domain as the block being estimated were considered.

The resulting probabilities of a block to be greater than or equal to the indicator grade gold were back-flagged to each 10m composite and served as selection criteria for the estimate of gold grade. Only blocks with a 50% or greater chance of being greater than or equal to



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the indicator grade gold were estimated. The gold indicator estimation parameters for Moore and Monte Negro are shown in Table 14.9.


Table 14.9		Search and Sample Selection Parameters for Gold Indicator Estimates (High and Low Grade)


	Search Orientation						Search Distance (m)						Sample Selection
Estimation Pass	Bearing		Plunge		Dip		Major		Semi-Major		Minor		Mini		Max		Max per DH
Moore 5 g/t Indicator		180		0		-20		120		120		40		5		13		2
Moore 1 g/t Indicator		180		0		-20		120		120		40		5		13		2
Monte Negro 5 g/t Indicator		340		10		0		120		120		40		5		13		2
Monte Negro 1 g/t Indicator		340		10		0		120		120		40		5		13		2

For high grade areas, gold grade estimation was undertaken for using inverse distance cubed (ID³), using matching composites. A series of two passes was used, with increasing search distances and varying numbers of composites.

For low grade areas, blocks that were not previously estimated for gold grade and had a 50% or greater chance of being above 1.0 g/t were estimated by ID³ with similarly flagged composites. A three-pass estimate scheme was used, with increasing distances and varying numbers of composites. Composites were capped at 6.0 g/t Au during the estimate (values not physically capped in the database)

For waste areas, blocks not previously estimated for gold and with less than 50% chance of being greater than 1.0 g/t, were estimated by ID³ using a three-pass scheme. Composites were also capped at 6.0 g/t Au during the estimate.

Some blocks were assigned a probability during the low-grade indicator estimate, but were not estimated for gold values as they did not meet the sample and distance requirements of the various runs. Also, many blocks were not assigned a probability as they met neither the sample nor distance requirements for the probability estimate. A last gold grade estimation pass was developed to address all blocks that were within 60m of a composite, but not yet estimated. This pass covered both Moore and Monte Negro domains, used a 60 m search with anisotropy and capped composite grade to 6 g/t Au.



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Table 14.10 Search and Sample Selection Parameters for Gold Grade Estimates


Estimation Pass	Search Orientation						Search Distance (m)						Sample Selection
	Bearing		Plunge		Dip		Major		Semi-Major		Minor		Min		Max		Max per DH
All Blocks Containing DH		0		0		0		5		5		5		1		99		1
Moore High-Grade Pass 1		180		0		-20		30		30		30		2		3		1
Moore High-Grade Pass 2		180		0		-20		60		60		60		2		3		1
Moore Low-Grade Pass 1		340		10		0		30		30		30		2		3		1
Moore Low-Grade Pass 2		340		10		0		60		60		60		2		3		1
Moore Low-Grade Pass 3		180		0		-20		30		30		30		2		3		1
Moore Waste Pass 1		180		0		-20		60		60		30		2		3		1
Moore Waste Pass 2		180		0		-20		120		120		60		1		5		1
Moore Waste Pass 3		340		10		0		30		30		30		2		3		1
Monte Negro High-Grade Pass 1		340		10		0		60		60		30		2		3		1
Monte Negro High-Grade Pass 2		340		10		0		120		120		60		1		5		1
Monte Negro Low-Grade Pass 1		180		0		-20		30		30		30		2		3		1
Monte Negro Low-Grade Pass 2		180		0		-20		60		60		30		2		3		1
Monte Negro Low-Grade Pass 3		180		0		-20		120		120		60		1		5		1
Monte Negro Waste Pass 1		340		10		0		30		30		30		2		3		1
Monte Negro Waste Pass 2		340		10		0		60		60		30		2		3		1
Monte Negro Waste Pass 3		340		10		0		120		120		60		1		5		1
All Areas — Final Fill-In Pass		360		0		0		60		60		60		1		5		1
Low-Grade Structural Block		360		0		0		60		60		60		1		5		1
Polygonal Estimate Ore		360		0		0		120		120		120		1		1		1
Polygonal Estimate Waste		360		0		0		60		60		60		1		1		1

Blocks falling within the Low Grade structural domain were estimated using a single pass with similar parameters to the last described estimate. Due to the generally much lower-grade nature of those composites, it was deemed unnecessary to separate mineralized from non-mineralized domains. Some higher grade assays do occur, but are not continuous and have been interpreted as being associated with structures. In order to keep those grades from spreading too far, all composite grades were capped at 3.0 g/t Au during the estimate.

Two final estimation passes were run to define the distance and grade of the nearest composite for the low-grade domain and the combined Moore and Monte Negro domains.

The estimation parameters for all these runs are shown in Table 14.10.

As stated above, intrusive andesitic dykes have been identified in the deposits, predominantly in the Monte Negro area, but the geological model did not allow accurate flagging of both model and composites. Once the gold grades were assigned to all blocks, the final grade of each block was further diluted by assigning a 0.00 g/t grade to the portion of each block that was considered to be dyke.



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14.3.18 AMC Opinion

In AMC’s view, the use of ID³ has resulted in insufficient smoothing and a degree of conditional bias of estimated gold grades (i.e. a tendency to over-estimate high gold grades and under-estimate low gold grades). Globally, the impact is not likely to be material to the project, but because the operation will process higher grades in the early years and stockpile lower grades for later treatment, the potential local impact may be material (see later discussion).

14.3.19 Sulphur Grade Estimation

Sulphur assays were composited to 10m lengths and a high grade cut imposed at 35% S (Figure 15.10). Variography was examined, (Figure 14.11), an indicator at 3% S was used for domaining purposes, and grades were interpolated into 10m by 10m by 10m blocks using ID²rather than the ID³ used for gold grade interpolation.

Figure 14.10 Frequency Distribution of Sulphur Grades in 10m Composites


						Untransformed Azufre Statistics															Log Normal Approximation Model
	Azufre Cutoff = 0.01%					Azufre Cutoff = 1.00%					Azufre Cutoff = 2.00%					Azufre Cutoff = 10.00%							Standard		Third
	Meters			S (%)		Meters			S (%)		Meters			S (%)		Meters			S (%)		Mean		Deviation		Parameter

all zones		95,254			6.773		90,368			7.120		85,554			7.435		15,397			13.418		1.80		0.53		0.00
incr. % and grade		5.1	%		0.346		5.1	%		1.525		73.7	%		6.122		16.2	%		13.418

low cut		0.01				35 g/mt percentile					GT lost by capping					percent of GT >= 43 g/mt

						99.95%					0.02%					0.01%



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Figure 14.11 Omni-Directional Correlogram — Sulphur

14.3.20 AMC Opinion

In AMC’s view, the sulphur block grade estimates should be reasonable, particularly given the robust correlogram.

14.3.21 Resource Classification

The resource model was classified using a combination of the estimation pass used to assign the block grade and the distance to nearest composite. If a block was intersected by an assayed drill hole, then the block was considered Measured. A block was considered Indicated if it had at least two holes within 60m, or at least one hole within 20 m. Blocks that fell outside the mineralized indicator or were within the Low Grade structural zones were considered Indicated if within 30m of a drill hole or within 60m of at least two drill holes. Otherwise, estimated blocks were classified as Inferred.

The 2006 drilling campaign carried out by Barrick drilled a series of holes that were located about 100m outside of the previously drilled area, yet at the edge or just within the previous pit limit. The geographical position of those holes often placed them in a position to be classified as Indicated, yet it was deemed that too much uncertainty existed about the geological interpretation and grade continuity between the main deposit and the new holes.



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A solid was created to limit the extent of the Measured / Indicated material and all blocks falling outside the solid were reclassified as Inferred or Unclassified.

The classification criteria are shown in Table 14.11.

Table 14.11 Classification Criteria


	Distance to
	Nearest Drill	Minimum Number of Drill
Estimation Pass	Hole (m)	Holes
All Blocks Containing DH	0 to 5	1 to 99		Measured
High-Grade Pass 1 or 2	5 to 60	2 to 3
Low-Grade Pass 1 or 2	5 to 60	2 to 3
Waste Pass 1 or 2	5 to 60	2 to 3
Low-Grade Pass 3	5 to 20		1	Indicated
Waste Pass 3	5 to 20		1
All Areas — Final Fill-In Pass	5 to 30	2 to 5
Low-Grade Structural Block	5 to 30	2 to 5		Inferred
Low-Grade Pass 3	20 to 60	1 to 3
Waste Pass 3	20 to 60	1 to 3
All Areas — Final Fill-In Pass	30 to 60	1 to 5
Low-Grade Structural Block	30 to 60	1 to 5

14.3.22 AMC Opinion

In AMC’s view, the extremely restrictive criteria for classifying Measured Resources lacks logic and is inconsistent with the geology of the deposit and the interpreted continuity of gold grades as illustrated by geostatistical analyses. It has, in AMC’s opinion, resulted in a significant under-statement of Measured Resources. The criteria for separating Indicated from Inferred Resources are considered reasonable.

14.3.23 Block Model Validation

The block model gold grade estimate was validated visually against drill holes and composites in section and plan view. Grades were also compared against the nearest neighbour (composite) gold grade and a histogram of the original composite distribution was compared against the gold grade estimate (Figure 14.12).



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Figure 14.12 Composite — Model Block Gold Grade Comparison


						Untransformed Gold Statistics															Log Normal Approximation Model
	Gold Cutoff = 1.00 g/t					Gold Cutoff = 1.70 g/t					Gold Cutoff = 5.00 g/t					Gold Cutoff = 6.00 g/t							Standard		Third
	Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Meters			Au (g/t)		Mean		Deviation		Parameter

All blocks		263,921,958			2.651		195,912,584			3.084		16,403,099			6.728		7,909,116			8.127		0.43		0.65		0.00
incr. % and grade		25.8	%		1.404		68.0	%		2.751		3.2	%		5.425		3.0	%		8.127		0.08
All Comps in 450 pit		14,628,569			3.110		10,995,724			3.693		2,031,703			7.233		1,305,688			8.214
incr. % and grade		24.8	%		1.346		61.3	%		2.891		5.0	%		5.469		8.9	%		8.214		0.18

14.3.24 AMC Comment

Figure 14.12 shows two block distributions, above and below 1g/t respectively. This is to be expected given the method used to model separately high and low grade blocks. However the composites (as shown on Figure 14.4) do not show two distributions. It would be expected that high grade blocks near the 1g/t boundary have been over-estimated and low grade blocks near this boundary have been under-estimated. High grade blocks above economic cut-offs will also have been over-estimated if located near blocks that were



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modelled as low grade (and whose low grade samples did not influence the high grade estimation).

Figure 14.12 shows that the blocks above 1g/t show less variability than the composites. Furthermore these blocks average 2.65 g/t, significantly less than the 3.11 g/t of the composites. These are encouraging signs, which indicate significant smoothing of the high grade and reduced likelihood of over-estimation

14.3.25 Mineral Resource Summary

Table 14.12 summarizes the mineral resources at a 1.4 g/t Au cut-off grade inclusive of those resources converted to Mineral Reserves. The Mineral Resources are based on Measured and Indicated classifications and a Whittle pit shell generated using the following metal prices:


Gold:		US$650/oz
Silver:		US$11.50/oz
Copper:		US$2.25/lb

Table 14.12 Total Mineral Resources at a 1.4 g/t Au Cut-off Grade

(Effective Date of Mineral Resource Estimate June 30, 2007)


		Tonnes		Au		Au		Ag		Ag		Cu		Cu		Zn		Zn		S
		(M)		(g/t)		(Moz)		(g/t)		(Moz)		(%)		(Mlb)		(%)		(Mlb)		(%)
	Measured		4.6		3.3		0.5		16.9		2.5		0.07		6.7		0.63		64.5		7.5
Monte	Indicated		80.9		2.9		7.5		13.8		35.9		0.06		99.9		0.50		888.2		7.5
Negro	Total		85.5		2.9		8.0		14.0		38.4		0.06		106.6		0.51		952.7		7.5
	Measured		7.9		3.3		0.8		18.3		4.6		0.11		19.5		0.86		150.1		8.2
Moore	Indicated		155.2		2.8		13.9		12.9		64.2		0.09		301.1		0.58		1,984.3		7.8
	Total		163.1		2.8		14.8		13.1		68.9		0.09		320.5		0.59		2,134.4		7.8
	Measured		12.5		3.3		1.3		17.7		7.1		0.10		26.2		0.78		214.6		7.9
Combined	Indicated		236.1		2.8		21.4		13.2		100.1		0.08		400.9		0.55		2,872.5		7.7
	Total		248.6		2.8		22.7		13.4		107.2		0.08		427.1		0.56		3,087.1		7.7
Goldcorp Share (40%)			99.4		2.8		9.1		13.4		42.9		0.08		170.8		0.56		1,234.8		7.7
	Inferred		81.4		2.5		6.5		3.4		9.0		0.02		40.0		0.02		33.4		7.7
Goldcorp Share (40%)			32.6		2.5		2.6		3.4		3.6		0.02		16.0		0.02		13.4		7.7

Resources are stated inclusive of resources converted to reserves

14.3.26 Comparison with Placer 2005 Estimate

Table 14.13 shows a comparison between the 2005 Placer and the 2007 Barrick FS resource estimates on a 100% basis.



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Table 14.13 Comparison of 2005 Placer and 2007 Barrick FS Resource Estimates (100%)


	2005 Placer Resource Estimate										2007 Barrick FS Resource Estimate
	1.7 g/t Au cut-off grade										1.4 g/t Au cut-off grade
	Tonnes		Au		Au		Ag		Ag		Tonnes		Au		Au		Ag		Ag
	(M)		(g/t)		(Moz)		(g/t)		(Moz)		(M)		(g/t)		(Moz)		(g/t)		(Moz)
Measured		118.6		3.2		12.3		18.1		69.2		12.5		3.3		1.3		17.7		7.1
Indicated		32.0		2.9		3.0		13.3		13.7		236.1		2.8		21.4		13.2		100.1
Total		150.6		3.2		15.3		17.1		82.8		248.6		2.8		22.7		13.4		107.2
Inferred		2.2		2.9		0.2		12.6		0.9		81.4		2.5		6.5		3.4		9.0

The differences in tonnage / grade estimates between the 2005 Placer estimate and 2007 Barrick FS estimate can reasonably be explained by the impact of additional drilling and the lower cut-off grade applied in 2007. AMC has been unable to obtain a 2007 estimate at 1.7 g/t cut-off and, therefore, can not make a direct comparison between the 2005 and 2007 estimates.

The difference in the proportion of Measured to Indicated Resources in 2005 versus 2007 (79% of total M + I in 2005, 5% in 2007) is very marked and is due primarily to the very restrictive classification applied to Measured Resources by Barrick in 2007.

14.4 AMC Comment and Opinion on Barrick 2007 Feasibility Study Resource Estimate

The use of ID³ for grade interpolation is relatively unusual in feasibility study resource estimates for gold deposits. It minimises the degree of grade smoothing, thus tending to maintain the variability of grades as reflected by composite samples, but it can result in conditional bias — a bias that depends on the cut-off grade applied. The tendency is to over-estimate high grades and under-estimate low grades. This may not be a material issue when the deposit is planned to be mined at around its average grade, as the conditional biases may approximately balance out. However, Pueblo Viejo will be mined at a higher than average grade for the early years, with lower grade material being stockpiled for treatment in the later years. In this situation, a conditional bias can be a material matter.

In order to investigate this possibility, AMC re-estimated gold grades for the deposit using ordinary kriging (OK), a technique that imposes a degree of grade smoothing and that should, ideally, result in an unbiased estimate. AMC then compared the resource estimates for mill feed for the ID³ model and the OK model for the first two years of planned mine production, the years of highest gold grade mill feed (Y02 also being the highest gold production year). The OK estimate resulted in an average gold grade and contained gold around 10% lower than those for the ID³ estimate. In AMC’s opinion, this confirms that the use of ID³ may be a material issue for the early production years.

Since AMC’s estimate was a check review undertaken with limited time, the results should be confirmed before acting on the findings. AMC recommends that more detailed investigations be undertaken to assess the validity of AMC’s conclusions. If they are shown to be valid, it may be advisable to undertake infill drilling in selected parts of the Moore and Monte Negro deposits so that the issue of the gold grade interpolation method can be more thoroughly evaluated.



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In AMC’s view, the classification of Measured Resources applied by Barrick, which takes no account of continuity of mineralisation between drill holes, is not logical, is inconsistent with the geology of the deposit and has resulted in a substantial under-statement of Measured Resources (and therefore of Proven Reserves). AMC recommends that the approach to Measured Resource classification be reviewed.

14.5 Barrick End-2007 Resource Estimate

In early 2008, Barrick released an updated mineral resource estimate for Pueblo Viejo based on substantially more diamond drilling. The drilling was primarily undertaken around the margins of the planned pit areas and resulted in the discovery of the Monte Oculto mineralization and western extensions to the Moore deposit (see Section 9.2 of this report). The discoveries increased total Mineral Resources, but an increase in the cut-off grade applied to resources and other changes such as the application of geological domaining rather than grade indicator domaining resulted in an increase in contained gold in the Measured and Indicated Resources of less than 10%.



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15 ADJACENT PROPERTIES

See Section 3 of this report for information on the Montenegro Fiscal Reserve.

Information received from Barrick during the AMC Pueblo Viejo site visit in March 2008 indicated that there will be a need to cross adjacent properties with an approximately 4 km road and pipeline from the Montenegro Fiscal Reserve to the Hatillo reservoir, which will be used as a source of mine site and process water. Barrick also reported that other organizations hold mining rights for the properties over which the road and pipeline would be routed, and that there will be a need to reach agreement with these organizations over access rights.

An Australian company, Las Lagunas Ltd, was granted a limited project approval on December 27, 2006 for the Las Lagunas development area, including the tailings impoundment facility, the limestone quarry to the northeast, the borrow-material area to the southwest, the area for facilities to the south of the dam, and other areas such as access from the main road, etc. Also, the Directorate of Mining granted Las Lagunas Ltd. the right to exploit the limestone quarry for the neutralization process.

Various unresolved environmental concerns regarding the construction, operation, and closure of Las Lagunas and the overlap of development areas have led to the Pueblo Viejo project being in a position of conflict with the potential operation of Las Lagunas. These issues, as far as AMC is aware, remain unresolved.

AMC is not aware of any reason why adjacent property issues should impact materially on the project, although the effect of any substantial delays in reaching required agreements may have some effect on the project schedule.


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17 ADDITIONAL REQUIREMENTS FOR TECHNICAL REPORTS ON DEVELOPMENT PROPERTIES AND PRODUCTION PROPERTIES

17.1 Mining Operations

17.1.1 Site Conditions & Choice of Mining Method

No mining or processing has occurred at the Pueblo Viejo mine site since June 1999. Previous mining activity is clearly seen in the in two main pit areas (Monte Negro and Moore — see Figure 17.1), some smaller pit areas, several rock piles, and two old tailings impoundments (Las Lagunas and Mejita). Among the remaining infrastructure items are power station, mill and treatment buildings, water supply, and housing and recreational facilities (see Figure 17.2).

Weathering and bacteria have created a large and continuing amount of acid rock drainage that contaminates area streams and rivers. The karstic nature of the surrounding Hatillo limestone formation presents a potential problem in the way of seepage losses from waste/tailings impoundments or for tailings dam construction.

The relatively bulk nature of the deposit is obviously relevant to the proposed mining method, as is the fact that the sulphide ore is very close to surface with outcrops in several areas. The gold occurs commonly as sub-microscopic particles in intimate association with pyrite mineralization.

As with previous mining at the site, the open pit method is the most economically viable. A 10m bench height has been selected for the pits, with consideration given to the equipment to be used and the desire to be relatively selective in order to have good control of both gold and sulphur content.

Pit dewatering will be required to manage the large amounts of surface water that will affect the area in the rainy season, with diverting ditches needed to minimize surface water entering the stockpile and pit areas.

Concerns about the potentially acid-generating nature of the waste rock will be addressed by its deposition, and subsequent covering with mine tailings, in an impoundment in the El Llegal valley, located about 2km south of the mine site.

Significant amounts of limestone will be required for processing, and construction of dams and roads. The necessary limestone will be taken from a series of quarries adjacent to the ore mining operations.


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Figure 17.2 Old Processing Plant at Pueblo Viejo

17.1.2 Mine Design Factors

17.1.2.1 Ore Production Rate

The maximum production rate from mining operations will greatly exceed the capacity of the processing plant, viz. approx.40,000 t/d mining vs. 24,000 t/d processing. This will influence the project in several ways:

•		Mining operations will be completed in 16 years as compared to the 26 years required for processing.

•		Ore from several different areas of the mine will be mined concurrently and stockpiled according to both gold content and sulphur grade. Ore with a higher gold grade will be mined and processed in the earlier years to benefit project economics.

•		Similarly, stockpiling ore according to sulphur grade is intended to allow blending for processing at around an average sulphur content of 6.75%, a necessary strategy to achieve the planned daily processing output.


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17.1.2.2 Ore Processing Rate

The capacity of the processing plant is limited by the rate at which the four planned autoclaves can handle sulphur, as constrained by oxygen availability. A ‘cap’ of 407 t/d per autoclave has been stipulated to the mining operations for ore delivery to the mill. At 6.75% sulphur, which is close to the average sulphur content of the reserve, this equates to 24,000 t/d of ore containing 1630 t of sulphur being processed. This figure is matched, with plant availability taken into consideration, to the design capacity of the crushing-grinding circuit and the processing plant as a whole. Figure 17.3 illustrates the daily throughput as a function of sulphur content. AMC notes that it clearly shows the necessity of maintaining the sulphur content of the ore to be processed at, or below 6.75%. The ultimate capacity of each autoclave is somewhat above 407 t/d but, for control of mill feed purposes, that has been deemed to be the maximum. The average sulphur content of the reserve is around 6.75%. Any time, therefore, that the sulphur content of the processed ore falls below that figure means that, at some future time, ore with a higher then average sulphur content must be processed, unless that ore has had sufficient stockpile time for the sulphur content to degrade (see section 17.1.7.4). AMC notes that processing of ore with a higher than average sulphur content may result in a processing throughput less than the design figure of 24,000 t/d.

Figure 17.3 Ore Treatment Rate

17.1.2.3 In-Situ Densities

Table 17.1 summarizes the main statistics for densities. These statistics were generated from a regression curve between density assays and sulphur content that was developed by Placer Dome (PDTS) for the 2005 Feasibility Study. The regression curve was used to assign density values to every block in the resource model. Refer to Section 14.2.5 of this report for a discussion of verification of density data.


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Table 17.1 Main Statistics for Densities


	Monte Negro (t/m³)						Moore (t/m³)
Rock Type	Average		Min		Max		Average		Min		Max
Black Sedimentary		2.83		2.62		3.27		2.81		2.62		3.31
Volcaniclastic		2.78		2.62		3.21		2.83		2.62		3.21
Spilite		2.83		2.62		3.21
Default Waste Rock	2.75

17.1.2.4 Metallurgical Recovery

The ore has been divided into six metallurgical domains by Barrick (see Table 17.2), based on mathematical equations representing the results of metallurgical testwork with respect to gold recovery. AMC notes that this means of ore division, and the relative quantities and characteristics in each domain, will have a large influence on stockpiling strategy.

Table 17.2 Ore Domains and Metallurgical Recoveries


	Metallurgical Recovery (%)
		Silver		Copper
Metallurgical Domain	Gold	(%)		(%)
(1) MO-BSD; Au > 1.7	Aurec = (Au — (0.2210 * LN(Au) + 0.107)) / Au * 100
(1) MO-BSD; ELSE	Aurec = (Au — (0.01165 * LN(Au) + 0.264)) / Au * 100		84.00
(2) MO-VCL & (3) Mo- Diorite	Aurec = (Au — (0.0318 * LN(Au) + 0.157)) / Au * 100		90.00
(4) MN-BSD	Aurec = (Au — (0.0522 * LN(Au) + 0.202)) / Au * 100		84.00
(5) MN-VCL	Aurec = (Au — (0.0345 * LN(Au) + 0.188)) / Au * 100		90.00
(6) MN-SP	Aurec = (Au — (0.0212 * LN(Au) + 0.126)) / Au * 100		87.00
All Met Types	Average = 92.05%		86.80		88.05

17.1.2.5 Geotechnical Parameters

For the initial feasibility study undertaken by PDTS, Piteau Associates Engineering Ltd. (Piteau) was retained to provide geotechnical slope design and blasting criteria for the project.

Slope Stability Analysis and Design

The Piteau pit slope design considered the pit layout based on the geologic model provided by PDTS in 2004. The size of the open pit has subsequently increased and additional geological and groundwater information has been obtained. AMC notes that the pre-feasibility design level of the 2004 Piteau report is recognized in the Barrick FS, along with the need to update the design using latest information and via a geotechnical program to be undertaken in 2008.

For the Piteau design, information was gathered from an investigation program that included geotechnical drilling and mapping, documentation of existing slopes, geomechanical core logging, field point load index testing, and sampling for laboratory rock mechanics testing (direct shear and uniaxial compressive strength). Field data included


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structural information, rock mass quality and estimated blast damage. The western part of Moore and some areas of Monte Negro were not accessible. There was limited mapping of limestone exposures.

Kinematic and stability analyses were done to develop bench and inter-ramp slope design. The work included detailed assessment of possible failure modes involving discontinuities that could result in shallow failure of individual benches. For this assessment, the proposed pit was initially subdivided into 59 design sectors within which the geologic structure, lithology, and slope orientations were expected to be consistent. Inter-ramp slope angles of 38 to 50 degrees were recommended for most rock types, and 5m bench height with inter-ramp slope angle of 34 degrees for the weaker saprolite and weathered/oxide zones. The effect of earthquakes on pit wall stability was not assessed.

The Piteau report indicates that overall slope stability could not be evaluated because of lack of groundwater data. Analysis showed that depressurization via dewatering wells and sub-horizontal drain holes may be required to achieve the required stability level in some zones. The report also gives information for slope maintenance (bench scaling and berm cleaning) and on pit wall monitoring methods.

Subsequent to the Piteau report, a drilling and blasting report was prepared in February 2005. The report provides blasthole depths, angles and layout for 171 mm diameter production drill equipment and blasthole patterns for final wall development using 159 mm diameter drill equipment. Recommendations are provided for the type of explosive that will be required depending on drilling costs and groundwater conditions encountered. The report indicates that steeper bench face angles may be achievable with controlled blasting.

AMC Comment

The Piteau design was developed based on standard rock mechanics investigation and testing methods. More recent information and the 2008 geotechnical program will allow the design to be brought to a true feasibility level.

AMC observation at the mine site of previous mining by Rosario shows some areas of slope degradation but an overall relatively high degree of slope stability for areas that have been standing for at least 10 years (see Figure 17.4 below).

It is noted that any depressurization work may be time consuming and potentially disruptive to the mining schedule; however, there appears to be sufficient flexibility in the mining plan for this not to be a major issue.

Inter-ramp slope angles of 38 to 50 degrees are deemed reasonable for the majority of rock types to be encountered. Recommendations for the weaker saprolite and weathered/oxide zones of 5m bench height and inter-ramp slope angle of 34 degrees appear to be appropriate.


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Figure 17.4 Workings in Monte Negro Pit

17.1.2.6 Ore Loss and Dilution

Previous geological drilling and analysis has indicated that mineralization occurs in a reasonably continuous spatial manner with recognizable contacts between mineralized and barren rock. In addition, the presence of waste rock appears rare within the mineralized zones. This knowledge has influenced the mine planning process in that no additional dilution or ore loss factor has been added to the reserves in the block model. From a visual review of geological and block model cross sections and a limited examination of drill core, AMC believes that the nature and extent of ore / waste contacts is such that mining dilution should be considered, although it is thought unlikely to have a material impact on the mineral reserves.

17.1.2.7 Limestone Consumption

Mining and processing activities will require significant amounts of limestone for:

•		Processing

•		Tailings dam construction for the Lower and Upper Llagal impoundments. Non-acid generating dams are required, with the dams being raised as the required volume of impoundment capacity increases.


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•

Construction, such as internal roads, diversion channels, and additional dams. Again, non-acid generating material is required.

AMC notes that, in the longer term, limestone quarry sites additional to those already identified and planned will be required, but that the surrounding terrain appears to offer viable opportunities to meet the project needs

The Barrick FS expresses the limestone requirement as:

Tonnes of limestone = 3.41 x Sulphur tonnes + 0.083 x Ore tonnes + 86,882

17.1.2.8 Commodity Prices

Long-term metal prices used to establish the mineable reserves and resources are:

Reserve estimate:


•	Gold	$575.00/oz
•	Silver	$10.75/oz
•	Copper	$2.00/lb

Resource estimate:


•	Gold	$650.00/oz
•	Silver	$11.50/oz
•	Copper	$2.25/lb

17.1.3 Mine Design and Planning Process

The following methodology, which AMC notes is in-line with industry practice, has been used by Barrick for pit limit analysis, cut-off grade optimization, production sequence and scheduling, and equipment/ manpower estimation:

•		Assignment of economic criteria to the mineral resource model.

•		Calculation of economic ultimate pit limits using the Whittle 4X software package. A series of nested envelopes for a given series of economic conditions is produced using the Lerchs-Grossmann algorithm.

•		Economic extraction sequence established using the Whittle nested pits as a guide.

•		Use of an initial set of smoothed, non-operational phases to evaluate preliminary production schedules with associated production rates, metal grades, and sulphur content.

•		Operational design of ultimate pit limit and internal mining phases using Gemcom software and NCL S.A. proprietary software.



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•		Production scheduling to evaluate options and maximize economic return while satisfying plant feed and mine production constraints.

•		Waste dump design and volume estimations using Gemcom software and NCL S.A. proprietary software.

•		Quarry production scheduling using the spreadsheet bench by bench reserve estimate from the 2005 Feasibility Study but converted to represent current status maps.

•		Estimation of mine equipment fleet from production schedules and representative performance and operational indices. Use of a spreadsheet model to estimate operating hours and number of units required. Measurement of haulage distances in the scheduling software, per bench and phase, according to the mine plan and definition of the haulage network. Use of model to generate procurement schedules, manpower, capital expenses, and operating costs.

•		Application of industry standards in considering equipment selection and a safe and productive operation for personnel, equipment and installations.

•		Estimation of equipment and manpower productivity from NCL experience and industry standards. Where applicable, compared indices with Barrick operations.

17.1.3.1 Resource Block Model

Resource Model Description (see Section 14.3)

The resource model for the Barrick FS contains data items that are coded and interpolated into 10 m x 10 m x 10 m blocks. The location characteristics of the model are summarized below:

Table 17.3 Block Model Basic Parameters


						No.
Direction	Min		Max			Blocks		Block Size
East		374,300		376,900	Columns		260		10
North		2,093,900		2,096,700	Rows		280		10
Elevation		-100		500	Levels		60		10

The following tables summarize the variables contained in the model:



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Table 17.4 Metal Fields


Code	Description	Unit	Min		Max		Average
AU_ID	Gold	ppm		0		20.00		0.13
AG_PPM	Silver	ppm		0		183.21		0.49
CU_%	Copper	%		0		1.51		0.0030
ZN%	Zinc	%		0		9.33		0.01
S_%_07	Sulphur	%		0		35.00		1.38
C_%_07	Carbon	%		0		3.98		0.10

Table 17.5 Category Field


Category		Description
0		Default
1		Measured
2		Indicated
3		Inferred
4		Other

Table 17.6 Metallurgical Field


Metallurgical Type		Description
1		Moore — Black Sediments
2		Moore — Volcaniclastic
3		Moore — Diorite
4		Monte Negro — Black Sediments
5		Monte Negro — Volcaniclastic
6		Monte Negro — Spilite

Material in the block model was classified as Measured, Indicated, and Inferred.

17.1.3.2 Variables Incorporated in the Block Model

Mining Cost

Initial mining costs came from previous Barrick work in 2007. These costs were:

•		Ore material $1.32/t + (300 — Z) * $0.002/t

•		Waste Material $1.73/t + (300 — Z) * $0.002/t

		Where Z = the elevation of the bench being estimated (m).

The higher waste cost is due to haulage distance to the waste dump. The incremental cost refers to hauling material either up or down to the pit exit, considered to be at level 300.



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Processing Cost

The processing cost has a fixed component for labour and general operation of the processing plant, together with a variable component that is dependent on the sulphur grade. The Barrick FS shows a formula expressing the processing cost as:

Process Cost = 17 + 0.0138*exp(%S/100/0.0138), with capping at 46 $/t

The above equation does not consider the copper smelting/refining cost, which was provided by Barrick at an estimate of $0.44/lb of treated copper in concentrate.

General and Administrative Cost

G&A annual expenditure was estimated to be $31,273,200, or $3.57/t processed when operating at design capacity (24,000 t/d and 6.75% sulphur). The mine plan shows that these operating parameters are feasible and the G&A costs have been fixed at $3.57/t processed. AMC notes that for a processing tonnage significantly different from 24,000 t/d, the fixed figure would obviously not be valid; however, the relative weight of any G&A cost differences compared to operating cost as a whole would not materially affect project economics.

Table 17.7 Payable Metal Transport and Refining Charges


				T&R Charges
Metal	Payable Metal			$/oz
Gold		99.925	%		1.10
Silver		99.000	%		1.10
Copper		96.500	%		0.36

17.1.3.3 Royalties

A royalty charge of 3.2% of the total revenue for produced gold, silver, and copper value was applied prior to the application of treatment, refining, and freight costs.

17.1.3.4 AMC Opinion

AMC believes that the variables used in the block model, and their respective magnitudes, are a reasonable representation of the anticipated cost regime at the time of the Barrick FS. AMC also notes that estimated mining and processing costs may have risen since the Barrick FS, particularly with regard to fuel prices.

17.1.4 Application of Variables

To establish the final pit shell and the mining sequence, the profit/day/t for each block was calculated (see below). This measure allows due reference to be given to metal grade, sulphur content and time in determining the value of a given block.



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•		Profit = Revenue — Costs ($/t)

•		Revenue/t = [Gold grade (oz/t) x Gold Rec. (%) x Gold price ($/oz) x (1 — 0.032) x Payable Metal — Gold TC&RC($/t)] + [Silver grade (oz/t) x Silver Rec. (%) x Silver price ($/oz) x (1 — 0.032) x Payable Metal — Silver TC&RC($/t] + [Cu grade (lbs/t) x Cu Rec. (%) x Copper price ($/L b) x (1 — 0.032) x Payable Metal — Copper TC&RC($/t)]

•		Cost: The total cost for each block considers all standard costs — mining, processing, G&A - plus the incremental sustaining capital associated with the El Llagal impoundment dams.

Ranking Index and Profit per Day

To optimize value, a Ranking Index was applied to each block of the resource model. This allows blocks with better gold, silver and copper grades, and lower sulphur grades, to be selected for earlier mining and processing (higher sulphur means longer processing time).

Measured and Indicated blocks were treated as potential mill feed, while Inferred and unclassified blocks were treated as waste and assigned a zero value in the Ranking Index. The following is used to calculate the Ranking Index variable within the block model:

Days/Block = Block tonnage / 24.000 for %S ≤ 6.75%

Days/Block = Block tonnage x (%S/100/1630) for %S > 6.75%.

The Ranking Index is then calculated dividing the profit/tonne by days/block:

Ranking Index = Profit / days

The higher the Ranking Index, the better the daily value of the block being mined and the higher priority that should be given to the block for early processing.

17.1.4.1 Tailings Dam Sustaining Capital Cost

In general, sustaining capital costs are not applied within an economic model calculation for the determination of material as ore or waste (cut-off grade calculation). In the Pueblo Viejo case of tonnage sensitive tailings dam expansions, sustaining costs have been applied to ensure that material placed in the tailings facility can pay for any required expansion costs. Both ore and waste incur sustaining capital costs because both the amount of tailings and the amount of waste contribute to the ultimate size of the tailings facility (reference the negation of the acid rock drainage nature of the waste by covering it with tailings).



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Sustaining capital costs for the El Llagal tailings dams are shown in the following table:

Table 17.8 El Llagal Sustaining Capital Costs

$/t of Material


	Waste		Ore

Lower Llagal
Incremental Sustaining Capex		0.23		0.58
Total		0.23		0.58

Upper Llagal
Incremental Sustaining Capex		0.51		1.31
Incremental Haulage		0.30		—
Total		0.81		1.31

For incremental ore, the sustaining capital cost added to operating cost is $1.08/t which corresponds to the difference between $1.31/t and $0.23/t.

17.1.5 Open Pit Optimization and Sensitivity Analysis

17.1.5.1 General

The Lerchs and Grossman algorithm in the Whittle software package was used for pit optimization and sensitivity analysis, with a set of nested pit shell surfaces being generated by varying the revenue factor.

The current topographic surface of the site was used in the analysis. Pit shell generation was unconstrained by infrastructure as all major facilities will be outside the ultimate pit design and area of influence. Only Measured and Indicated resources were used for pit optimization and mine design. Inferred material within the mine design was only reported to estimate possible opportunities for additional mineral resources. It is considered as waste in the Barrick FS.

Unit costs and prices, as indicated in Sections 17.1.3 and 17.1.4, were used as economic assumptions.

17.1.5.2 Resource Model

The original blocks of 10 m x 10 m x 10 m, were used; no reblocking was done.

17.1.5.3 Slope Angles

A simplification of the Piteau matrix was utilized, differentiating seven regions for slope angles. Wall slope angles were reduced as appropriate for ramp width and assumed wall height before application in Whittle. Piteau-recommended angles were shallowed by 3° to account for ramps to be included in the smoothed pit.

Table 17.9 summarizes the inter-ramp slope angles used for pit optimization:



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Tonnages contained in the pit shell generated by the Whittle exercise are shown in Table 17.10. The pit lines were subsequently smoothed to create an operational pit shell from which final reserve estimates were generated (see Section 17.1.6.4).

Table 17.10 Pueblo Viejo Pit Optimization Tonnages

Total

Rock

Ore

Waste

Deposit

(kt)

(g/t)

(%)

(kt)

S.R.

Monte Negro

144,273

72,639

2.97

15.6

0.060

6.8

71,634

0.99

Moore

258,169

134,895

2.96

14.3

0.099

7.4

123,273

0.01

Total

402,442

207,534

2.96

14.8

0.085

7.2

194,908

0.94

Note: Excludes Rosario stockpiles

17.1.5.5 Sensitivity Analysis

Whittle was used to evaluate sensitivity to gold price and slope angles.

Gold Price Sensitivity

Results for the four gold prices used are shown in Table 17.11. With an increase in gold price to $675/oz, pit size and recovered gold ounces increased by 13% and 10% respectively. Gold at $475/oz reduced pit size and recovered gold ounces by 18% and 14% respectively.

Table 17.11 Pit Optimization Sensitivity to Gold Price


Gold Price	675		625		Base Case		525		475
Ore Tonnes (Mt)		240		227		208		192		168
Au (g/t)		2.81		2.87		2.96		3.04		3.17
Ag (g/t)		14.01		14.31		14.76		15.23		16.00
S (%)		7.30		7.26		7.22		7.16		7.11
Contained Au (M oz)		21.7		20.9		19.8		18.8		17.1
Relative Contained Au (%)		110		106		100		95		86
Excavation Size (Mt)		456		436		402		378		328
Relative Excavation Size (%)		113		108		100		94		82
Tonnes Mined/Contained Au (oz)		21		21		20		20		19

Note: Excludes Rosario stockpiles



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Slope Angle Sensitivity

The slope angle exercise showed no significant change in gold extracted or total ore tonnes (see Table 17.12). In the Barrick FS this is seen as an indication of the relatively shallow pits and the fact that wall slope is mainly driven by ore distribution. AMC notes that the % change in the total volume of ore caused by relatively small changes in slope angle is very small and should not, therefore, have any significant effect.

Table 17.12 Pit Optimization Sensitivity to Pit Wall Slope


	Steeper						Shallower
Slope Angle Variation	+5		+3		Base Case		-3		-5
Ore Tonnes (Mt)		209		208		208		207		207
Au (g/t)		2.96		2.97		2.96		2.96		2.96
Ag (g/t)		14.72		14.74		14.76		14.76		14.75
S (%)		7.22		7.22		7.22		7.22		7.22
Contained Au (M oz)		19.9		19.9		19.8		19.8		19.7
Relative Contained Au (%)		101		100		100		100		100
Excavation Size (Mt)		390		394		402		415		424
Relative Excavation Size (%)		97		98		100		103		105
Tonnes Mined/Contained Au (oz)		20		20		20		21		21

17.1.6 Open Pit Design and Sequencing Method

17.1.6.1 Design Parameters

The Barrick FS final pit and intermediate phase designs consider the following parameters:

•		Bench height 10 m

•		Minimum pushback width 70 m

•		Road width 26 m

•		Maximum road grade 8% in-pit (bottom benches allow 10%) and 10% out of the pit.

A simplification of the Piteau matrix of inter-ramp angle, face angle, and bench height was used for design, with the number of slope domains minimized by their classification as a function of slope value. Classifications and slope angles are listed in Table 17.13.



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Table 17.13 Slope Design Parameters based on Piteau Recommendations


			Inter-Ramp		Height		Face		Berm
	Zone		Angle		Bench		Angle		Width
MO-vi		1		38		10		75		10.1
MO-v y ii		2		42		10		75		8.4
MO-iii		5		47		20		75		13.3
MO-i		6		48		20		75		12.6
MO-iv		7		50		20		75		11.4

MN-xi		2		42		10		75		8.4
MN-vii y x		3		44		10		75		7.7
MN-viii		4		46		20		75		14.0
MN-ix		7		50		20		75		11.4

Note: It can be seen that a common face angle was utilized to simplify the design process. Therefore, the berm width actually indicated may vary somewhat based on the actual face angle of the domain. Inter-ramp angles are maintained.

The pit design considers eight mining phases, as shown in Table 17.14 below, that follow the sequence determined by the Whittle runs, of which five are in the Moore area and three in Monte Negro.

The Moore and Monte Negro inter-ramp slope angles vary between 38º and 50º, and 42º and 50º respectively. Bench configuration considers single-bench and double-bench operations, with berm widths varying between 7.7 m and 14.0 m, depending on the bench configuration and inter-ramp slope angle assigned by geotechnical domain.

AMC again notes that the geotechnical design is at a pre-feasibility level and understands that it is the intention of Barrick to use more recent information and results of future geotechnical work to arrive at final design details.

17.1.6.2 Pushback Geometry

Minimum operational width has been established at 70 m. All of the pushbacks exceed the minimum width, except for Moore Phase 0, which is close to minimum.

17.1.6.3 In-Pit Access Ramps

Internal and external roads were designed at 26 m width, adequate for medium-size trucks. In general, ramp slopes were designed at 8%, except for the last three benches of each phase where the maximum ramp slope is 10%. AMC notes that Barrick is now indicating larger reserves and more waste to be handled than that indicated in the Barrick FS, with the consequent possibility that larger equipment may also be used. These issues must, of necessity, be referenced in the final operating pit design.

17.1.6.4 Final Pit Determination

Based on the optimized pit at $575/oz gold, $10.75/oz silver and $2.00/lb copper, pit lines were smoothed to make an operational pit shell, with appropriate accesses created. Figure 17.6 is a


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section through the Monte Negro pit illustrating the relationship between the operative final pit and the Whittle optimized pit shell.

Figure 17.6 Section at 95,600 Monte Negro

17.1.6.5 Sequencing Method

Whittle nested shells were used as a guide to define mining sequence, considering minimum pushback width and economic contribution. Table 17.14 shows mining sequence and key phase parameters. The sequence follows the Ranking Index (higher means more value), except where this is not possible for geometry reasons.

Table 17.14 Phase Mining Sequence


															Total
	Ore										Waste		Au to		Tonnage/
Phase			Au		Ag		Cu		S		Tonnage		Process		oz Au		Sulphur		Ranking
Sequence	(kt)		(g/t)		(g/t)		(%)		(%)		(kt)		(k oz)		Processed		(kt)		Index
MN-1		24,338		3.80		22.80		0.06		6.58		7,064		2,972		11		1,602		349
MO-0		16,166		3.73		22.72		0.12		8.35		4,846		1,940		11		1,350		248
MO-1		12,153		3.25		18.21		0.13		7.51		3,292		1,269		12		912		237
MO-2		41,462		2.86		12.87		0.11		7.07		23,022		3,813		17		2,933		201
MN-2		18,029		2.28		15.25		0.06		6.18		27,685		1,323		35		1,114		158
MO-3		33,600		2.70		12.70		0.07		7.42		44,673		2,913		27		2,494		159
MO-4		31,727		2.69		12.08		0.08		7.30		73,045		2,740		38		2,318		161
MN-3		28,249		2.69		10.02		0.06		7.38		34,859		2,442		26		2,086		147

17.1.6.6 AMC Comment

AMC is of the opinion that the design and sequencing method described in the Barrick FS is appropriate for mining in the Pueblo Viejo pits.

17.1.7 Mine Production Schedule & Forecast

17.1.7.1 Basic Criteria

•		18,000 t/d ore processing capacity in Y01, ramping up to full capacity at 24,000 t/d in Y03


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•		Target mining rate from the Monte Negro and Moore pits at 30 Mt per year total material moved, excluding stockpile re-handle and limestone quarry production

•		Two or three active production phases at any time, with a minimum of two phases supplying ore to the plant for sulphur blending purposes. This in line with a major aim in the Barrick FS of having sufficient flexibility to create a low risk operation in terms of ability to achieve production targets.

•		Maximum production rate per phase and per period established according to the geometry of the phases and the number of shovels that can work continually within that geometry.

•		Average of seven to eight sinking benches per annum in the initial years.

•		Blocks identified and mined as per the Ranking Index parameter described above in order to maximize NPV and cash flow. Twelve different bins located in two areas. Six higher value bins at Moore Phase 4, Level 300, and six lower value bins within the mill stockpile. Target mining rate from the Monte Negro and Moore pits at 30 Mt per year total material moved, excluding stockpile re-handle and limestone quarry production

•		For stockpiling purposes, three sulphur grade ranges defined:

•		No allowance for severe climate phenomena. On average, about two days of climate related downtime per year per piece of equipment has been allowed.

17.1.7.2 Process Constraints

Sulphur Content

Sulphur content is a major factor in achieving the design average processing rate of 24,000 t/d. For sulphur content equal to or below 6.75%, the design rate is achieved. For higher sulphur values, the processing rate reduces (see Figure 17.3).

Autoclaves Start-Up

The start-up of the four autoclaves requires ore feed with gold content between 2 and 3 g/t Au and sulphur content between 8% and 9% S.

Ramp-up to full capability is scheduled through Y01 and Y02 as shown in Table 17.15:


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Table 17.15 Autoclave Ramp-Up


	4 Autoclaves		Sulphur Treatment
Period	Operation (%)		Capacity (t)
Year 1
Quarter 1		20		29,514
Quarter 2		54		80,217
Quarter 3		57		84,213
Quarter 4		65		97,139
Total Year 1		49		291,083
Year 2
Quarter 1		70		104,736
Quarter 2		76		112,607
Quarter 3		73		108,823
Quarter 4		75		111,699
Total Year 2		74		437,865
Year 3 Ahead		100		595,000

17.1.7.3 Pre-Production Mine Development

The initial pre-stripping requirement is very low as previous mining has left ore outcropping on surface. Mine activity will be performed in phase Monte Negro 1, removing 3.3 Mt of material, (2.2 Mt for stockpiling). Some material from Moore 4 will be removed in advance (originally planned for Y07) to generate a platform for the medium-term stockpile; a total of 3.5 Mt will be removed, including 1.1 Mt of ore which will be stockpiled.

The total stockpile prior to production will be 3.3 Mt with 3.61 g/t gold and 6.5% sulphur.

During this period, roads to connect both pits with the primary crusher, the two stockpile areas, the limestone quarries, and the El Llagal tailings/waste rock facility will be constructed.

17.1.7.4 Stockpiling

Ore blending for sulphur content, and early processing of high grade ore are key to maximizing NPV. Stockpile management is, therefore, a prime consideration. For the different stockpile bins, three cut-off grades have been defined for sulphur content:

•		LS Sulphur ranging between 0% and 6%

•		MS Sulphur ranging between 6% and 7%

•		HS Sulphur above 7%

For each cut-off grade, the ore is divided into 3 groups, depending on Ranking Index value.


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Stockpile stability was evaluated using accepted methodology. Analysis was performed at three sections, with results showing both static and dynamic stability, and only minor deformation under severe earthquake conditions.

Rough estimates of volumes, tonnes, and grades for existing stockpiles have been calculated. Scheduling for processing will be done according to the competitiveness of their grades. Barrick has used a decay function to calculate the effect of natural degradation in the levels of contained sulphur in stockpiles. That function is defined as:

Decayed Sulphur Grade (%) = (1 — 0.0118) ^N

Where N is the number of years the stockpile has been exposed.

The decay curve was applied to the sulphur grades, assigning N as the time difference between the moment when the section was stockpiled and the moment when the ore is reclaimed.

17.1.7.5 Mine Plan and Production Scheduling

Plan Options

Table 17.16 compares economics for four long-term mine plans developed as part of the Barrick FS.

Table 17.16 Summary of Long-Term Mine Plans


	Mine Movement			NPV @ 5% (M$)
Mine	Excl. Limestone		Blending for			First 10		First 5
Plan	(Mt/a)		Sulphur	LOM		Years		Years
5		30	Yes		3,516		2,525		1,690
6		35	Yes		3,554		2,546		1,698
7		40	Yes		3,554		2,563		1,690
8	Variable (Avg. 36 Mt/a)		No		3,429		2,516		1,675

Plan #5, with total material movement of 30 Mt/a (excluding limestone removal) was selected, as NPV differences are negligible compared to the 35 Mt/a and 40 Mt/a options. AMC notes that all options had the same basic premise of early high grade and focus on sulphur blending with mining rate being the only significant difference

Cut-off Grade Strategy

As indicated in Section 17.1.4, the Ranking Index parameter, which accounts for block value as well as treatment rate, is used to maximize NPV of cash flows. RI cut-off grades were established for each period and for each category of high, medium, and low sulphur content. An example of this is Table 17.17 below which shows cut-off grades by year for the high sulphur category.


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Table 17.17 High Sulphur Ore Cut-off Grades


High	Ranking		Mine to Stock
Sulphur	Index		High-Grade Stock				Medium-Grade Stock				Low-Grade Stock
Year	COG		From		To		From		To		From		To
Y001		600		300		600		100		300		0.001		100
Y01		600		300		600		100		300		0.001		100
Y02		400		300		400		100		300		0.001		100
Y03		300		200		300		100		200		0.001		100
Y04		300		200		300		100		200		0.001		100
Y05		300		200		300		100		200		0.001		100
Y06		300		200		300		100		200		0.001		100
Y07		150						100		150		0.001		100
Y08		150						100		150		0.001		100
Y09		150						100		150		0.001		100
Y10		150						100		150		0.001		100
Y11		150						100		150		0.001		100
Y12		150						100		150		0.001		100
Y13		150						100		150		0.001		100
Y14		100						100		100		0.001		100
Y15		100						100		100		0.001		100
Y16		100						100		100		0.001		100

Mine Life and Material Movement

Processing higher grade ore in the early years, while stockpiling lower grade ore for later processing, results in a pit life of 16 years and a processing life of 26 years. In the Barrick FS, detailed plans of material movement were generated for each year of operation. In the steady state mining years (Y02 to Y14), total material movement, including limestone, averages about 40 Mt/yr; and about 59% of ore is stockpiled for later processing. Maximum gold and silver production is in Y02 at 1,126 k oz gold and 5,205 k oz silver. Maximum limestone consumption is in Y12 at 8.8 Mt, related to tailings dam construction. The maximum medium to long term stockpile capacity requirement is 82.1 Mt in Y15.

Figure 17.7 illustrates the yearly proportion of ore to crusher direct from the mine and from medium-to-long term stockpiles. Figure 17.8 indicates daily material movement rates throughout the life of the project.



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Mining Phases

The plan shows at least two phases simultaneously active in each year. A maximum of 9 benches par phase will be mined in each year. Preferential mining for grade purposes is shown in years when one active phase has higher sinking rates than other active phase(s).

Table 17.18 shows LOM ore mined from each phase and delivered either to the crusher or stockpiles:

Table 17.18 Phase Ore Mining by Period


Total Ore (kt)
Period	1-mn1		2-mn2		3-mn3		4-mo0		5-mo1		6-mo2		7-mo3		8-mo4		Total
Y001		2,220														1,043		3,263
Y01		12,871						6,716										19,587
Y02		7,920						8,569		4,641.0		718						21,848
Y03		1,146						993		7,437.3		8,355						17,932
Y04				293								8,362		3,578				12,233
Y05				266								9,849		4,989				15,104
Y06				1,080								10,847		6,159				18,086
Y07				2,240								3,351		9,078				14,669
Y08				5,630										6,378		1,694		13,702
Y09				4,135										3,084		3,020		10,239
Y10				4,292												3,558		7,850
Y11						4,132										2,389		6,521
Y12						6,011										1,628		7,639
Y13						4,091										3,109		7,200
Y14						6,580										5,329		11,909
Y15						7,241										7,608		14,849
Y16																2,385		2,385

Total		24,158		17,936		28,054.2		16,278		12,078.4		41,482		33,267		31,762		205,016

17.1.7.6 Short-Term Planning

The early years of the project (Y-01, Y01, and Y02) are particularly important in terms of setting up the ore-mining process and then delivering on early high grade ore. To demonstrate the viability of the planned approach the first three years were planned out on a monthly basis.

To assess ore blending and short-term stockpiling requirements, a daily-based simulation was carried out for Y01 (Months 7 and 8); Y02 (Months 6 and 10), and Y07 (Month 7). Assumptions were: mill availability at 95%, daily sulphur limit at 6.75%, mill capacity at 24ktd, and typical production rates and availabilities for shovel and loader equipment. AMC notes the reference in the Barrick FS that no potential climatic events during the periods were evaluated.

Sulphur content was found to be a limiting factor on a significant proportion of operating days, with the average re-handling requirement at up to 25% per period. The results of this exercise were used to generate re-handling estimates for all years of operation, with this data then being applied to the assessment of costs and equipment needs.



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17.1.7.7 Limestone Quarries

Limestone Requirements

Total limestone required through the 26 years of processing activity is shown in Table 17.19:

Table 17.19 Project Limestone Requirements


Purpose	(kt)
Process		68,091
Tailings Dams		48,322
Road & Construction		4,948
Total		121,361

Limestone Tonnages

The identified tonnage of 77kt does not satisfy total project requirements. AMC notes that additional potential sources (see part of Table 17.20 below) require further definition work.

Table 17.20 Limestone Tonnages and Uses


	Bottom Elev.		Process		Construction		Waste		Total
Quarry	(m)		(kt)		(kt)		(kt)		(kt)
					Identified Quarry Sources
Quemados		130		29,736		12,408		4,683		46,827
Las Lagunas		180		8,508		3,550		1,340		13,398
Plant		180				14,939		1,660		16,599
Total Identified				38,244		30,897		7,683		76,824


					Assumed Quarry Sources
	Bottom Elev.		Process		Construction		Waste		Total
Quarry	(m)		(kt)		(kt)		(kt)		(kt)
Las Lagunas Extension		130
Hatillo Formation		150		3,533		1,474		556		5,563
Las Lagunas Formation		200		26,314		21,046		6,765		54,125
Total Assumed				29,847		22,520		7,321		59,688
					Total
Total				68,091		53,417		15,004		136,512

Limestone Extraction Schedule

The limestone has been scheduled to be mined according to different quality requirements. AMC notes the assumption that 20% of the material quarried will be waste material and that the basis for that is previous working experience at a 5 Mt/a limestone quarry in England. Further definition drilling may be required to substantiate this assumption, as in-situ quality will obviously affect quarry scheduling.



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17.1.7.8 Waste Dump Sequencing

The waste dump is located within the tailings dam impoundment area. Tailings will cover the waste rock shortly after its deposition and help minimize any acid rock drainage. The waste rock is to be deposited in 5m lifts, with the level of tailings generally maintained close to the advancing crest level of the waste dump.

To maintain the waste rock level only slightly above the tailings level, two levels of waste rock will generally be maintained: 1) A higher waste dump lift which acts similar to a coffer dam, and 2) a lower lift behind this higher lift. This gives the plan more flexibility to respond to variations in waste rock production rates without curtailing plant (and therefore tailings) production rates.

Waste rock will be deposited in the Lower Llagal impoundment up to the 245m level, after which deposition will be done in the Upper Llagal impoundment, starting in Y10.

AMC notes that the deposition of waste rock in the Upper Llagal impoundment will begin in the tenth year of mine life, with the deposition of tailings not beginning until Y13. This will require that the waste be kept under water prior to tailings deposition.

17.1.8 Mine Equipment

17.1.8.1 Equipment Requirements

Equipment planning has considered mine design production of approximately 40 Mt/a, including mill feed of 24 kt/d, with simultaneous mining in one limestone quarry and two different pits in up to four operating phases.

The drilling and loading equipment has to combine high productivity and low cost with high mobility to allow maximum flexibility and selectivity.

The time model used for calculating equipment hours is shown in Figure 17.9.



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Figure 17.9 Equipment Hours Model

Drilling equipment will consist of four diesel units capable of drilling 200 mm diameter holes in all materials. Pre-splitting for final phase walls will be drilled with the same equipment.

A trade-off analysis was developed for the Barrick FS to select 136t or 177t trucks, with the chosen option being 136t trucks loaded by 15 m³ hydraulic shovels and 19 m³ front-end loaders. AMC again notes that the recently reported increase in reserves and waste over and above the Barrick FS numbers has prompted a re-evaluation of equipment size.

The truck fleet has thirty-four 136t units, in line with results of a detailed analysis of hauling distances and cycle times for every type of material per phase and period. Truck speeds were determined upon the basis of typical values, with correction factors to allow for slower velocities at the benches and at the dumps, and for weather conditions. Truck hours were calculated per period, type of material, and loading unit.

Ancillary equipment includes bulldozers, wheel-dozers, graders and water trucks.

Some equipment will be purchased early to give additional flexibility and redundancy during the particularly critical, initial mining years.

17.1.8.2 AMC Comment

AMC is satisfied that the choice and quantity of equipment is generally appropriate for the Pueblo Viejo project. Selection has been made with reference to production demands and using both first principles and operating experience. The general implications of the tropical/sub-tropical nature of the climate, with significant rain storms not an unusual occurrence, have been considered as part of the equipment selection and operation estimate process. AMC notes that, at a currently operating open pit in the same region of the Dominican Republic, 872, 1,640 and 4,040 hours respectively of truck time were reported lost for the years 2005 through 2007 for a fleet of about 20 trucks. The 2005/2006 average lost time per truck was around 2.5 days. In the case of shovels over the same period, hours reported lost were 55, 101 and 610 hrs respectively for about 12 units, the average lost time per shovel for 2005/2006 being around 0.25 days. For dozers there were no registered hours lost in these years, heavy rain periods being a time when the dozer demand is highest. The Barrick FS indicates that lost time of about two days per piece of equipment per year has been assumed for estimation purposes. This does not include any allowance for major climatic events. The 2007 figures quoted above are largely a reflection of the impact of tropical storm Noel, with rainfall estimated as a one in 100-300 year event, and in December of the same year, with tropical storm Olga.


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17.1.9 Workforce Requirements

17.1.9.1 Operations

Operations workforce requirements have been estimated as a function of the equipment fleet per year and with reference to processing and plant requirements.

A summary of the total manpower per year is shown in Table 17.21.

Table 17.21 Total Mine Labour per Period


	-3		-2		-1		1		2		3		4		5		6		7		8		9		10		11		12
Overhead & Supervision		49		70		91		101		101		103		103		103		102		102		102		102		102		102		102
Mine Operations		5		16		25		27		27		31		31		31		31		31		31		31		31		31		31
Mechanical Department		8		23		24		27		27		26		26		26		26		26		26		26		26		26		26
Technical Services		36		31		42		47		47		46		46		46		45		45		45		45		45		45		45
Mine Operations		64		68		99		167		203		209		219		219		219		233		233		228		227		232		231
Mechanics		32		32		50		90		114		119		127		127		127		136		136		139		139		139		139
Total		145		170		240		358		418		431		449		449		448		471		471		469		468		473		472


	13		14		15		16		17		18		19		20		21		22		23		24		25		26
Overhead & Supervision		102		100		98		94		57		57		57		50		49		49		49		49		48		46
Mine Operations		31		29		29		25		17		17		17		14		14		14		14		14		14		14
Mechanical Department		26		26		26		26		21		21		21		19		18		18		18		18		17		17
Technical Services		45		45		43		43		19		19		19		17		17		17		17		17		17		15
Mine Operations		236		241		203		114		76		71		71		71		66		66		66		66		66		69
Mechanics		139		143		116		68		49		49		49		49		46		46		46		43		43		41
Total		477		484		417		276		182		177		177		170		161		161		161		158		157		156

17.1.9.2 Training Program and Labour Build-Up

During pioneering work, Owner personnel will be working in quarries, control, planning, and management areas. Hiring of operations and maintenance personnel will begin with initial safety training followed by training for individual functions and as equipment is available.

•		New operators, mechanics, and staffing personnel for maintenance and operations will be hired considering a training period of three months.

•		Staffing personnel for technical services will be hired 11 months before operations start, ref. the need for detailed design and planning, controls and procedures, etc.

•		The project will employ as many local people as possible during operations. Skills development programs and training initiatives are planned for local persons with the aim of minimizing the number of external employees.

•		Over $4 million has been assigned to train personnel over the life of the mine to ensure a safe and productive workplace.


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17.1.10 Mineral Reserve Estimate

Table 17.22 summarizes the Mineral Reserves. The reserves are derived from the Measured and Indicated Resources based on Whittle pit shells generated using the following metal prices:

•		Gold US$575/oz

•		Silver US$10.75/oz

•		Copper US$2.00/lb

Table 17.22 Mineral Reserves

(Effective Date of Mineral Reserve Estimate June 30, 2007)


		Tonnes		Au		Au		Ag		Ag		Cu		Cu		S
		(M)		(g/t)		(Moz)		(g/t)		(Moz)		(%)		(Mlb)		(%)
	Proven		4.3		3.3		0.5		17.9		2.5		0.07		7.0		7.0
Monte Negro	Probable		66.3		2.9		6.3		15.6		33.3		0.06		86.0		6.8
	Total		70.6		3.0		6.7		15.8		35.8		0.06		93.0		6.8
	Proven		6.9		3.4		0.8		19.6		4.4		0.12		18.0		7.7
Moore	Probable		128.2		2.9		11.9		14.0		57.8		0.10		277.0		7.4
	Total		135.1		2.9		12.7		14.3		62.1		0.10		295.0		7.4
Rosario Stockpiles	Probable		8.4		2.3		0.6										5.4
	Proven		11.2		3.4		1.2		18.9		6.8		0.10		25.0		7.5
Combined	Probable		202.9		2.9		18.8		14.6		91.0		0.09		363.0		7.1
	Total		214.1		2.9		20.0		14.8		97.9		0.09		388.0		7.1
Goldcorp Share (40%_)			85.6		2.9		8.0		14.8		39.2		0.09		155.2		7.1

17.1.11 Mineral Resource Estimate

Refer to Section 14.3.25

17.1.12 AMC Assessment and Opinion

17.1.12.1 Review of Mine Plan

The following is a description of the steps taken by AMC to examine the Pueblo Viejo proposed mine plan.

•		Review of final pit design and general plan review to verify the ROM feed accessibility and overall sequence and schedule of material movement.

•		Detailed review for years 01 and 02 with creation of a series of plan maps at every bench elevation indicated in the Barrick FS as active. For each year the blocks which met the mill feed Ranking Index COG criteria were coloured differently from the remaining material.

•		Specific checking of each of these maps to confirm the accessibility and sequencing for material planned to be mined in those years.


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From the above assessment it is concluded that the proposed mine plan should be able to meet the Barrick FS ore production schedule. Again it is noted that the requirements for the ROM contribution to the mill feed blend will necessitate careful scheduling at the operational level. The Barrick FS demonstrates a considered approach to this issue with a proposed schedule that, on a month by month basis, always has two accessible ore faces in the pit.

17.1.12.2 General Comment on Mine Design and Mining Operations

The Barrick FS describes a very detailed process of mine design, estimation, and production scheduling that is to an acceptable industry standard. Some of the key results of that exercise are discussed below.

The mining rate and overall pit production schedule are viable considering the pit design, sinking rate, mining sequence and flexibility, and estimated resource requirement in terms of equipment and personnel. The impact of normal climatic conditions has been catered for in the production calculations, but without any particular recognition of tropical storms or hurricanes that can strike the area on occasion. The potential effect of such events is somewhat mitigated by the fact that mine production rate far outweighs processing capability.

A major item with respect to gold production is the ability of the mine to produce ore at the metal grade and sulphur content levels required to satisfy the processing schedule. In this regard there is a particular risk in years 01 and 02 when autoclave capability is still building, but high gold production is projected. A categorical understanding of high grade areas and their extent, together with very selective mining and a disciplined stockpiling process, will be necessary to achieve mill feed goals. The ability to rapidly assay production drill holes will also be vital.

It is noted in the Barrick FS that the open pit limit is significantly larger than that of the 2005 feasibility study and that the geotechnical design criteria in the area of the new high walls is only of pre-feasibility study level. This issue must obviously be addressed to arrive at final design details but should not pose a major problem for the project.

Further work is required on pit dewatering to manage the large amounts of surface water that will affect the area in the rainy season. This item should not be of material significance to project viability.

Uncertainty about the longer-term limestone supply is well recognized in the Barrick FS but remains to be addressed. Further definition drilling may also be required to verify quality and substantiate assumptions about percentage waste content.

Assessment of reserves since the Barrick FS has indicated that increased volumes of ore and waste may be mined and that larger equipment may consequently be employed. It will be imperative that the implications of using such equipment be adequately incorporated into the final pit design.

17.1.12.3 Validation of Reserves

AMC used the following process to validate the reserves:


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Reproduction of Barrick FS Reserve Numbers

•		Check of the block model ASCII file provided by Barrick with data provided in the Barrick FS. The model was seen to be valid.

•		Examination and validation of the script used by Barrick to calculate the Ranking Index item in the model

•		Reproduction of the script in Minesight and appropriate coding of model items.

•		Reserve calculation using the topographical and final pit strings provided by Barrick. The reserve calculation matched to within less than 1% of the Barrick FS numbers.

Assessment of Impact of Gold Grade Estimation Method on Total Reserves

•		Creation of a second model in Minesight containing the ordinary kriging grade for gold (AUCP) instead of the ID³ grade (AUID)

•		Calculation and coding into the model of the Ranking Index items based on the ordinary kriging gold grade

•		Comparison of the estimates for the two models. The AUCP model gave 9% more ore tonnes but at a 3% lower average gold grade. The net effect of recovered ounces was a 6.4% increase (17.7Moz to 18.8Moz).

Assessment of Impact of Gold Grade Estimation Method on First Two Years of Gold Production

•		From the pit strings for the end of years 0, 01 and 02 provided by Barrick, calculation of pit ore production in years 0, 01, and 02 using the AUID model

•		Using the same strings, calculation of ore production from the pit in years 0, 01 and 02 using the AUCP model. These initial years showed the same trend as the overall reserve with the AUCP model returning more ore tonnes but at a lower average grade

•		Division of the blocks for year 02 into Ranking Index ranges of 100, from >0 to >=800. Using Ranking Index COG of 400 for Y02 (as per the procedure for the Barrick-calculated reserves) the AUCP model showed a shortfall of about 0.5Mt of ROM ore to the process plant compared with the AUID model. Lowering of the COG to 380 resulted in the AUCP Y02 pit delivering the required ore tonnes to meet the mill feed schedule, but at a potential loss of 105,000 recovered ounces or 9% of total gold production for that year.

•		Execution of a similar process for year 01 to that described above for Y02. Using the same Ranking Index COG as the AUID model, the AUCP model estimate showed enough material to meet the mill feed criteria for Y01 but at a lower gold grade.

•		Addition of a new item to the AUID and AUCP block models and coding of it with stockpile destinations based on the COG strategy outlined in the Barrick FS. This was done to enable calculation of the difference in recovered gold grade in the stockpile


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material used for Y01 and Y02 processing. The stockpile material used for mill feed in Y01 and Y02 showed 6% and 1% deficits respectively when comparing the AUCP model to the AUID model. The combination of the ore material delivered directly from the pit to the process plant with these stockpile grades resulted in a potential reduction of 11% and 9% in produced gold ounces in Y01 and Y02 respectively. This equates to a potential deficit of just less than 10% in produced gold by the end of Year 02. Table 17.23 below summarizes the results of this exercise.

Table 17.23 ID3 and OK Estimates of Mill Feed Y01 and Y02

ID³ and OK Estimates of Mill Feed Y01 and Y02


	Y01									Y02
	Feed		Recovered gold, oz/t							Feed		Recovered gold, oz/t
Source	kt		AUID		AUCP		% Diff			kt		AUID		AUCP		% Diff
HSP1		638		0.1971		0.179858		-8.75	%		813		0.162079		0.157561		-2.79	%
MSP1		0		0.149135		0.139388		-6.54	%		888		0.102091		0.102145		0.05	%
LSP1		338		0.12569		0.122713		-2.37	%		343		0.099758		0.100493		0.74	%
LSP2		0		0.082181		0.082412		0.28	%		117		0.081132		0.082007		1.08	%
HSP4		35		0.065204		0.071395		9.50	%		0		0.072731		0.079766		9.67	%
MSP4		0		0.045507		0.048281		6.10	%		14		0.046407		0.049923		7.58	%
ROM		3268		0.195291		0.171104		-12.39	%		4309		0.169454		0.147324		-13.06	%

Total		4279		0.188999		0.167771		-10.87	%		6484		0.153757		0.138554		-8.95	%

17.1.12.4 Comment on Reserves

Comment has been made in Section 14.4 on the ID³and OK methods of estimation. The approximately 10% negative difference in estimated gold production for Yrs 01 and 02 when using the OK method is cause for concern given that those years have a very significant impact on NPV. Again as indicated in Section 14.4, AMC’s OK estimate was a check review undertaken with limited time and the results should be confirmed before acting on the findings. AMC recommends that more detailed investigations be undertaken to assess the validity of AMC’s conclusions. If they are shown to be valid, AMC believes that a prudent course of action may be to drill additional diamond drill holes in and around the planned areas for early higher grade mining.

17.2 Mineral Processing and Metallurgical Testing

17.2.1 Introduction

See Section 21 (References) for a list of studies and publications on which this section is based.

17.2.2 Ore Mineralogy

Section 8 of this report contains a more comprehensive description of the orebody mineralization than the brief summary contained in this section.


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17.2.2.1. Mineralization

The Pueblo Viejo deposits will be mined in two open pits, namely Moore and Monte Negro. The ore from both pits is refractory, mainly due to gold encapsulation in the pyrite minerals, but also due to preg-robbing carbonaceous materials and to cyanide-consuming copper and zinc minerals. Pyrite and sphalerite are the two main sulphide minerals, both occurring in veins and disseminated within the host rock.

Five metallurgical ore types have been defined based on lithological and mineralization criteria, namely:

•		Moore black sediment (MO-BSD) — Fine interbeds of carbonaceous shale and siltstone. Bedding is sub-horizontal and is intersected by vertical sulphide veins. It is a main lithology and exposed within the Moore pit. This ore type is moderately preg-robbing.

•		Moore volcaniclastic (MO-VCL) — A group of volcanic (andesitic) lithology units in the Moore pit. Units include massive and fragmental volcanic flows as well as sedimentary units composed primarily of volcanic material. These units typically have lower organic carbon content. It is not exposed in the present surface and sits below MO-BSD. This ore type is not preg-robbing.

•		Monte Negro spilite (MN-SP) — Volcanic spilite (andesite) flows are found at depth. It is currently exposed only at the north end of the Monte Negro pit. This ore type is not preg-robbing.

•		Monte Negro black sediment (MN-BSD) — Interbeds of carbonaceous shale, silt stone and volcanic flows. Beds are up to 3 m thick and have a shallow dip to the south. The carbonaceous beds are similar to MO-BSD and comprise more than 50% of MN-BSD. It is currently exposed in the eastern half of the Monte Negro pit. This ore type is moderately preg-robbing.

•		Monte Negro volcaniclastic (MN-VCL) — Similar to MN-BSD except the unit is less than 30% carbonaceous beds. It is currently exposed in the western half of the Monte Negro pit. This ore type is mildly preg-robbing.

17.2.2.2 Gold Deportment

In addition to the mineralogical examinations used to identify gold association in the various ore types reported in Section 8, diagnostic leach procedures were also employed. The result of the diagnostic leach tests showed that approximately 55% to 70% of the gold is encapsulated in sulphide minerals and is not recoverable by cyanide leaching without prior destruction of the sulphide matrix. For the two black sedimentary ore types, MO-BSD and MN-BSD, 19% to 29% of the gold in the ore was preg-robbed by gold adsorption onto organic carbon.

For MO-VCL, MN-SP and MN-VCL ore types, 6% to 9% of the gold was also preg-robbed. This may be caused by gold adsorption onto sulphide minerals as these ore types contain very little organic carbon. Laboratory tests have demonstrated that the preg-robbing ability of the ore is reduced after the ore is oxidized in an autoclave. At a grind size of 80%



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passing 150 µm, less than 2% of the gold in the ore was locked up in the silicate gangue minerals.

Laboratory tests showed that pressure oxidation of the whole ore followed by CIL cyanidation of the autoclave product will recover 88% to 95% (average 91.6%) of the gold and 86% to 89% (average 87%) of the silver.

17.2.2.3 Variation in Sulphur Grade

The efficient and trouble-free operation of the pressure oxidation (POX) circuit relies heavily on maintaining relatively constant sulphur content in the autoclave feed. If the variability of the sulphur grade in the short-term is significant, it may be necessary to blend ores with different sulphur content in the mine. For this reason, the short-term variability in the sulphur content of the ore was assessed extensively in 2004. Block models for sulphur grade were developed on a block size of 50 m x 50 m x 10 m containing roughly 70,000 t of ore. The result of this exercise showed that there is a wide variation in the sulphur content of the ore as the blocks are mined sequentially. The variation in sulphur grade ranges from 3% to 20% sulphur and generally between 5% and 10%.

Blending is necessary to maintain a relatively constant sulphur grade to the autoclave feed. Blending will be practiced by the mine through mine planning and blending of ores prior to crushing. In the mill, some blending will occur as a result of the surge capacity provided for the autoclave feed. Although there will still be variation in the sulphur grade, this variation will not happen abruptly, but rather in a slow and controlled predictive manner. Therefore, adjustment in process conditions to suit the sulphur content of the feed can be anticipated.

17.2.3 Metallurgical Investigation of Process Options

17.2.3.1 Introduction

This section describes the testwork and investigations undertaken to establish the principle design criteria for the processing of Pueblo Viejo ore. This work needs to be considered in the context of the ore processing flowsheet selected for the project as shown in Section 17.2.4.

17.2.3.2 Metallurgical Studies Prior to Placer Dome (Before 2003)

The metallurgical history of the Pueblo Viejo project has been summarized by Pincock Allen & Holt (PAH). Table 17.24 is duplicated from PAH and summarizes the efforts made to develop an economic processing concept for Pueblo Viejo.

The PAH table shows that aggressive and expensive sulphide ore pretreatment routes, such as roasting or pressure oxidation, were required prior to cyanidation to achieve gold recoveries to bullion in excess of 80%.



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Table 17.24 Summary of Metallurgical Test Programs (from PAH)


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



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17.2.3.3 Placer Dome and Barrick Metallurgical Testwork (2003-2007)

Introduction

Bio-oxidation of whole ore and flotation concentrate, and ultra-fine grinding of flotation concentrates, were subsequently investigated by Placer Dome as alternative pretreatment options prior to carbon-in-leach cyanide leaching for gold and silver recovery. Ultimately a fairly straightforward process based on pressure oxidation of the whole ore followed by carbon-in-leach cyanidation (CIL) was selected for the recovery of gold and silver. Two innovations have also been incorporated into the process design:

•		A hot cure of the slurry from the autoclave to reduce lime consumption by solid basic ferric sulphate in the CIL circuit.

•		A lime boil process, involving heating the CCD washed slurry to 80-85^oC with 35kgCaO/t to release the silver in the jarosites formed in the autoclave for improved CIL silver recovery.

Testwork Samples

A number of ore samples from each of the five ore types were used for the initial metallurgical investigations. These samples were assayed in detail before being used in the various test programs. The following information is relevant to the processes considered:

•		The gold content of the ore samples ranged from 2.10 g/t to 6.60 g/t

•		The sulphur content ranged from 6.9% to 9.7%

•		The ores contained insignificant amounts of elemental sulphur and sulphates

•		The black sedimentary ore types (MO-BSD and MN-BSD) contained from 0.5% to 0.7% organic and graphitic carbon, which caused preg-robbing in the later leaching tests. The other ore types have very weak or no preg-robbing ability

•		The carbonate content varied from 0.05% to 0.37% CO2 but averaged 0.19% CO2

•		The aluminum content ranged from 7% to 10%

•		The mercury content ranged from 8 g/t to 14 g/t. The extent of mercury dissolution during pressure oxidation varied significantly according to the ore type

•		The arsenic content ranged from 260 g/t to 1,650 g/t. Most of the arsenic was dissolved and precipitated during pressure oxidation

Three ore types were used for the later metallurgical investigation. Similarly these samples were composited and assayed in detail before being used in the test programs to confirm the effectiveness of the silver enhancement process. The following information is relevant to the process evolved:

•

The gold content of the ore samples ranged from 5.3 g/t to 5.6 g/t



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•

The silver content of the ore samples ranged from 19.6 g/t to 36.1 g/t

Comminution Testwork

The original Placer Dome feasibility study circuit was designed with high pressure grinding rolls (HPGR) technology. Subsequent trade-off studies concluded that a conventional SABC circuit offered superior economics.

Work index (Wi) measurements of the five main rock types undertaken in 2004 indicated that the Bond ball mill Wi of the ore will vary from 12.8 to 16.1 kWh/t (average 14.4 kWh/t), while the rod mill Wi will vary from 14.9 to 18.6 kWh/t. Supplementary testwork undertaken on 58 different samples in April 2006 for SAG Power Index (SPI®) and Wi returned consistently higher Wi values as per Table 17.25.

Table 17.25 Comminution Testwork


	Modified Bond Wi
Ore Type	BSD		SP		VCL		All Ore Types
Average		17.05		18.17		15.62		16.73
80th Percentile		18.37		18.97		17.92		18.28

The Bond ball mill work indices (Wi) used to size the grinding mills was the average Wi for the hardest of the five ore types (MN-SP) and approximately the 80 ^th percentile Wi of all ore types.

Grinding simulations using the Minnovex proprietary program “Comminution Economic Evaluation Tool”, or CEET®, were undertaken using SPI® values from all 58 samples. The result of these simulations was almost identical to Fluor’s SAG mill power estimate when using the 18.1 kWh/t ball mill Wi.

Whole Ore Pressure Oxidation (POX)

Whole ore pressure oxidation followed by CIL was selected as the preferred process option around July, 2003 after a reasonable power cost was assured. Pressure oxidation gave higher gold recoveries for all the Pueblo Viejo ore types tested, using a technically robust, proven process. Pressure oxidation is energy intensive, more so than most other refractory processing options. Due to the complexity of the autoclaves and associated oxygen plant, pressure oxidation is also capital intensive. The higher gold recovery and associated cash flow mitigate the high energy operating cost and the high capital cost of the oxidation circuit.

Extensive batch and continuous pilot autoclave testwork was undertaken at the Barrick Technology Centre, (BTC, formerly the Placer Dome Research Centre), and at SGS Lakefield. Testwork at SGS Lakefield included a two week pressure oxidation pilot plant programme in 2004. The results of the testwork undertaken on whole ore at the design autoclave operating conditions and grind size P80 of 80 microns is summarized in Figure 17.10



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Figure 17.10 Effect of Gold Head Grade on Gold Recovery

Scale formation inside the autoclave was an issue during pilot plant operation. Most of the scale, which comprised basic ferric sulphate and lesser hematite, was formed in the first compartment and became increasingly less severe towards the end of autoclave. Severe scale formation can offer brick or liner protection inside the autoclave but it can also impair agitation efficiency, oxygen injection and dispersion, slurry flow and control of slurry levels in the autoclave. Analysis of the scale chemistry and review of scale formation, and management in other autoclave circuits, has led to design features aligned to help prevent the formation of scale. Allowance has also been made in the autoclave operating strategy and maintenance schedule for control of scale formation.

Hot Cure

At a temperature significantly below the autoclave operating temperature of 230°C, basic ferric sulphate formed during POX at this temperature dissolves to form ferric ions in acidic solution. The test program showed that by holding the autoclave flash discharge slurry for a period of 12 h at 85°C to 100°C, the basic ferric sulphate solids formed in the autoclave re-dissolves to form ferric sulphate in solution. The formed ferric ions are washed away from



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the CIL feed in the three-stage CCD washing thickener circuit. The re-dissolution of basic ferric sulphate takes place in what has been termed the hot curing step in the flowsheet.

With the addition of the hot cure, it becomes possible to remove the effects of high lime consumption in CIL and concentrate on the optimization of the pressure oxidation process. It is preferable to operate with as high as possible a temperature in POX to allow for the fastest kinetics. A temperature of 230°C was considered the maximum practical temperature for POX and was therefore chosen for all subsequent tests.

Counter Current Decantation (CCD)

Three-stage CCD washing was tested as part of the POX pilot plant operation in 2006. Based on this testwork, 99.3% wash efficiency is expected with an average thickener underflow density of 40% solids. These results confirm the three-stage CCD washing circuit testwork undertaken for each of the five ore types in the original pilot plant operation in April 2004.

Lime Boil

In 2006, a lime boil/CIL study was undertaken to improve the silver recovery.

Bench scale tests were performed using washed, CCD thickened, and underflow slurry for the tested ore composites. Liberation of silver was shown to reach completion within 2 h. No apparent improvement in gold extraction was observed with longer retention times. Variations in pulp temperature were shown to have a large effect on the amount of silver liberated. Indications from the bench scale results in Figure 17.11 show that the process was best carried out at as high a temperature as practical to minimize lime consumption and achieve the highest gold and silver extraction rates.

Figure 17.11 Effect of Temperature on CIL Silver Extraction from Lime Boil Plant Operation

Carbon-in-Leach (CIL)



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Placer Dome testwork conducted in 2003 established that destruction of the preg-robbing carbon in the black sedimentary ores during pressure oxidation was slow. Reduction in the organic carbon content through extended residence time, which also corresponds to higher sulphur oxidation, reduces the degree of preg-robbing thereby improving gold recovery. Unlike direct cyanidation (DCN), up to 0.50% organic carbon may be tolerated in the oxidized solids with CIL cyanide leaching before there is a noticeable drop in gold recovery. This effect is shown in the results from testwork undertaken on MO-BSD and MN-BSD ore types In Figure 17.12.

Figure 17.12 Relationship between Gold Recovery and Organic Carbon Content*


*		MO-BSD and MN-BSD Ore Types, batch tests, 2L autoclave, 30% pulp density, 230°C and grind size P80 of 90 to 100 microns.

CIL pilot plant runs were undertaken by Barrick in June 2006 on three ore types to determine maximum precious metal loadings on carbon and gold and silver extractions. Average gold recoveries ranged from 90.5% (MO-BSD) to 95.2% (MN-SP) and silver recoveries, 84.4% (MO-BSD) to 89.9% (MO-VCL). The Barrick FS concludes that “the performance of the gold and silver loadings was considered above expectation for the three ore types” with total loadings of 12,000 g/t (gold plus silver) being achieved.

Copper Recovery

Copper dissolution is very high under the expected operating conditions in the autoclave. The value used in the design criteria is 97.5%, but it will likely be higher than this, particularly after hot curing.

Copper recovery from autoclave discharge solutions was tested using sulphide precipitation process. Copper can be selectively precipitated as a copper sulphide (CuS) using hydrogen sulphide (H₂S). The H₂S is produced from the action of bacteria under anaerobic conditions fed with elemental sulphur, ethanol, and nutrients. The sulphide concentrates at a high grade can be sold to a third party smelter.

Initial preliminary batch testwork was carried out in May 2004, followed by a continuous pilot plant campaign in September 2004, and finally concluded by a Prefeasibility Study in



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November 2004. Encouraging results and higher metal prices required a re-evaluation with pilot plant testwork by SGS Lakefield in September 2006.

In summary, the pilot plant operated without problems and a consistently good concentrate grade was obtained. After the losses from POX, CCD, and iron precipitation are taken into account, recovery was excellent at more than 99% for the precipitation stage and 88.05% overall. Copper concentrates analysed 58.5% Cu, 26.7% S and 0.26% Zn.

Cyanide Destruction

The CIL tailings slurry generated during the pilot plant campaign in June 2004 was sent to Inco Tech to evaluate the effectiveness and economics of cyanide destruction. The conventional SO2/air cyanide destruction process was selected but confirmatory testwork was required in 2006 with the incorporation of the lime boil into the process flowsheet. This testwork was successful in reducing the residual WAD CN (weak acid dissociable cyanide) to below 1.0 mg/L in the treated Pueblo Viejo Project tailings slurry.

Neutralization of Autoclave Acidic Liquors

Significant amounts of sulphuric acid and soluble metal sulphate salts are produced during pressure oxidation. The acidic liquor generated from the pilot plant operation in April 2004 was used to determine the most cost-effective neutralization process.

A continuous seven-day pilot plant test was subsequently performed to confirm the limestone and lime to sulphur ratios determined in batch testwork. The high density sludge (HDS) neutralization process was used in the pilot plant. The HDS process removes the contained base metals in a chemically stable form by co-precipitating them with ferric iron hydroxide in the presence of limestone or lime.

The pilot plant confirmed the effectiveness of limestone neutralization removing 92.5% of the sulphate, 99.9% of aluminum and copper, effectively 100% of iron, and 86.8% of the zinc with less than 1 mg/L of the metals left in solution. The sulphate level in the clarifier overflow was 1,800 mg/L for removal of 94%. Manganese removal was 89.8% at a final concentration of 1.6 mg/L.

Limestone Grinding, Calcining and Slaking Testwork

Samples representing the limestone deposit were sent to the SGS Lakefield Research laboratory for Bond work and abrasion index measurements. The result showed that the Bond Wl of the limestone deposit ranged from 8.4 kWh/t to 10.1 kWh/t (average 9.5 kWh/t). Most of the samples tested assayed better than 96% CaCO₃.

Six limestone samples were collected in 2004 and 2005 and sent to Maerz in Switzerland for calcining and slaking tests. The results of the testwork showed that:

•		The CaO content of the kiln product ranged from 94% to 96%.

•		The burnt limestone had a high mechanical stability.

•		The burnt lime was highly reactive



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17.2.4 Recoverability

17.2.4.1 Overview

The following process description is largely taken from the Barrick December 2007 Feasibility Study Update.

The various circuits of the processing plant will be planned around the autoclave’s designed operation at 1,200 t/d of sulphide sulphur oxidation (before expansion) and 1,600 t/d (after expansion). Initially there will be three autoclaves that will process ore at rates varying from 12,000 t/d with a sulphide sulphur grade of 10% to 18,000 t/d with a sulphide sulphur grade of 6.69%. The fourth autoclave, which is planned to be operational at the end of Y02 will raise this to a design rate of 24,000 t/d. To provide the ability to handle the 24,000 t/d rate, some equipment, such as crushers, feeders/conveyors, grinding mills, thickeners, lime boil tanks, CN destruction tanks will be initially sized to handle this rate. Others, such as the autoclave feed storage tanks, the autoclaves, the hot cure tanks, the CIL tanks, the acid wash/elution vessels, and the reactivation furnaces will be sized to handle 18,000 t/d. Additional equipment will be added when the expansion is undertaken.

The proposed process plant will consist of the following major unit operations:

•		Run-of-mine (ROM) crushing

•		Grinding

•		Pebble crushing

•		Pressure oxidation

•		Hot curing

•		CCD thickener washing

•		Iron precipitation sludge circuit

•		Copper recovery

•		High density sludge neutralization circuit

•		Solution cooling

•		Silver enhancement, lime boil of oxidized solids

•		Oxidized slurry dilution and cooling

•		Carbon-in-leach (CIL)

•		Refining

•		Cyanide destruction

•		Tailings disposal

•		Water treatment

A simplified block flowsheet of the process plant design concept is provided as Figure 17.13.



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17.2.4.2 Flowsheet Description Summary

The ore is ground to its optimum grind of 80% passing 80 μm and pressure-oxidized in autoclaves for 60 min to 75 min at a temperature of 230°C and a pressure of 3,450 kPa. The autoclave product is discharged to a flash tank where heat is released, cooling the slurry to about 106°C. It is then transferred by gravity to the hot cure circuit where the slurry temperature is maintained at 100° to 105°C for 12 h to dissolve basic ferric sulphate that forms during autoclaving.

The slurry from the hot cure circuit is then washed to separate the acidic liquors from the oxidized solids. The washing is accomplished in a three-stage countercurrent wash thickener circuit to remove more than 99% of the sulphuric acid and the dissolved metal sulphates. The washed thickened slurry is contacted with steam from one of the autoclave flash vessels to heat the slurry to 95°C ahead of a two-stage lime boil treatment. Milk of lime slurry is added to the oxidized slurry to raise the pH to the 10.5 to 10.8 range to effectively break down the silver jarosites, exposing silver minerals to CIL leaching. Lime boil slurry is then diluted with reclaim water and cooled to 40°C in cooling towers. The cooled slurry is pumped to the CIL circuit.

Lime addition to the lime boil circuit provides sufficient protective alkalinity in the CIL circuit and no further addition of lime is required in this circuit. In the CIL circuit, cyanide is added to dissolve the gold and silver and is contacted with activated carbon to adsorb the gold and silver cyanide complexes. The retention time in this circuit varies from 18 h to 22 h, depending on the processing rate.

The overflow (acidic liquor) from CCD Thickener #1 is sent to the autoclave plant to quench flash steam. The quench vessel underflow is then treated with limestone in the iron precipitation circuit to remove ferric iron. The overflow from the iron precipitation thickener is forwarded to the hydrogen sulphide precipitation plant to recover the copper. In this plant, H₂S gas is added to the solution to precipitate the copper as CuS. The precipitate is thickened and filtered to produce market grade copper concentrate. The thickener overflow solution is neutralized, first with limestone and then with lime in the high density sludge (HDS) circuit where most of the remaining metal sulphates are precipitated. After neutralization, the slurry is thickened in a high rate thickener. The thickener underflow (sludge) is pumped to the tailings pond while the overflow is cooled and then recycled to the process water tank for distribution to the process, including use as wash water in the CCD circuit.

Loaded carbon from the CIL circuit is forwarded to the refinery for acid washing and stripping. The resulting pregnant strip solution goes to the electrowinning circuit for gold and silver recovery while the barren carbon goes to the reactivation kiln. A combined gold and silver sludge from the electrowinning cells is filtered, dried, retorted to remove the mercury from the sludge, and smelted to produce bullion bars. The reactivated carbon is recycled to the CIL circuit.

The CIL tailings slurry flows over the safety screens and is pumped to the cyanide destruction circuit. The conventional SO₂/air process is used to reduce the cyanide content of the CIL tailings solution from more than 100 g/t cyanide to less than the regulatory maximum of 5 mg/L (5 ppm) cyanide. The detoxified slurry together with the HDS circuit sludge is pumped to the tailings pond.


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17.2.4.3 Primary Crushing

The primary crushing station will consist of a primary gyratory crusher equipped with a hydraulic rock breaker to break oversize rocks in the dump pocket. Water sprays will be provided at the truck dump pocket and an ADS (fogging dust suppression) system will be provided at the feeder to conveyor transfer point to comply with the dust emission standards in the Dominican Republic.

From the gyratory crusher station, the ore is transferred by an apron feeder onto a stacking conveyor that discharges the ore onto a 16,000 t live capacity stockpile. The material flow rate from the crusher to the stockpile is monitored with a belt scale.

The reclaim tunnel below the stockpile will be equipped with a dust control system servicing the material transfer locations. Two variable speed apron feeders under the coarse ore stockpile reclaim the ore and feed a common SAG mill feed conveyor. The feed rate to the SAG mill is monitored by a belt scale installed along the SAG mill feed conveyor.

The proposed ore primary crusher has a rated capacity slightly higher than the design rate of 24,000 t/d, and therefore, there will be no change in the size of the crusher.

The proposed limestone primary crusher is exactly the same size as the ore primary crusher, and therefore this crusher is more than adequate for the 12,000 t/d rate.

17.2.4.4 Grinding

As it will be very difficult and prohibitively expensive to add a second SAG mill and a second ball mill or tower mills during the Phase 2 expansion, the ore grinding mills are sized to handle 24,000 t/d.

Both the ore SAG and ball mills are equipped with variable speed drives. This allows the plant operator to control the plant throughput to accommodate changes in sulphide sulphur feed grade or to reduce tonnage when an autoclave circuit is down for maintenance.

The SABC grinding circuit will consist of the following:

•		One SAG mill (9.76 m dia x 4.90 m EGL) driven by two 4,500 kW synchronous motors with variable frequency drive (VFD).

•		One ball mill (7.93 m dia x 12.40 m EGL), driven by a 16,400 kW ring (wrap-around, ring gear) motor.

•		One pebble (cone) crusher.

•		Two (one standby) SAG mill discharge screens.

•		A cluster of fifteen hydrocyclones. An additional three classifiers will be added for the 24,000 t/d ore production expansion.

The SAG mill will be equipped with pebble ports to ensure that critical size buildup in the mill will not become a problem. The discharge from the SAG mill is screened to separate


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the oversize pebbles which are transferred onto a conveyor recirculation loop feeding the material to the pebble crusher, or alternatively bypassing the pebble crusher if it is not in service. The pebble crusher product is conveyed back to the SAG mill feed conveyor. The SAG mill discharge screen undersize material is pumped to the cyclone feed pump box.

The ball mill will be in closed circuit with a cluster of fifteen cyclones, expandable to eighteen. The cyclone underflow gravity-flows back to the ball mill feed chute while the overflow flows by gravity over two vibrating trash screens. The screen undersize is thickened to approximately 50% solids in a 70 m diameter high rate thickener. The thickener underflow is pumped to the autoclave feed storage tanks while the overflow is recycled to the grinding circuit.

17.2.4.5 Pressure Oxidation

The selection of the process design criteria and the design of the pressure oxidation and ancillary processes based on the results of the various test programs were completed by Hatch Engineers.

The pressure oxidation facility comprises three autoclave circuits, expandable to four autoclave circuits, with minimal interconnections to achieve high capacity utilization. Each autoclave circuit includes a high pressure slurry feed system, slurry preheater, autoclave vessel and agitators, flash vessels, and gas handling system. Supporting the operation of the autoclaves are agitator seal systems, steam boiler (for start-up), and a high pressure cooling water system for autoclave temperature control.

The autoclave vessels are refractory lined with approximate process dimensions of an inside diameter of 4.9 m and an overall length of 37 m. The autoclaves will operate at 230°C and 3,450 kPa, with retention time between 60 min and 75 min depending on sulphur grade and feed density.

Oxygen required for the oxidation reactions in the autoclaves is provided from an on-site oxygen plant.

Two of the three autoclave circuit preheating systems are used for slurry feed heating, while the third preheating system is used for heating washed CCD underflow slurry prior to the lime boil process. The design incorporates slurry piping interconnections between these preheating systems to allow for maintenance and descaling while maintaining capacity utilization. The gas handling design will adopt a solution spray quench process providing over 90% condensation of the flash steam. Depending on the preheating requirements, a portion of the flash steam will be used to preheat autoclave feed slurry or lime boil feed slurry with the remaining steam reporting to the gas handling system. The quenching of the excess flash steam and autoclave vent gas is accomplished with CCD overflow solution. The hot CCD overflow solution then reports to the partial neutralization circuit.

17.2.4.6 Oxygen Plant

The air separation unit (ASU) is designed to supply 2,850 t/d of gaseous oxygen as well as trickle liquid oxygen. Although this is a large capacity ASU plant (compared to industry references), it is still well below the largest single ASU plant of 4,300 t/d operating at


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SASOL, Secunda, South Africa. For expansion to 24,000 t/d a second oxygen plant will be required with a capacity of 1,100 t/d contained oxygen.

The ASU plant design is based on machinery that is widely used in the cryogenic gas industry and will adopt a double column cryogenic distillation process. This is a conventional process for the air separation industry.

17.2.4.7 Hot Curing

Oxidized slurry produced from the initial 18,000 t/d capacity rate is held in 4 cascading tanks in series for a total of 12 h. Upon expansion to 24,000 t/d, an additional tank will be commissioned to ensure the optimal dissolution of basic ferric sulphate.

The slurry feed to the hot cure circuit arrives at approximately 105°C and, based on heat loss calculations, completed by Hatch, exits at approximately 100°C. The cured slurry flows by gravity to the first CCD thickener.

17.2.4.8 Countercurrent Decantation (CCD) Washing

Slurry from the last hot cure tank is treated in a three-stage CCD circuit. Each thickener will be 70 m in diameter, constructed with 316 L stainless steel walls, floor and rakes. The wash circuit is designed to treat the expanded tonnage capacity of 24,000 t/d.

The purpose of this circuit is to wash and separate acid and soluble metal salts from the gold-bearing solids phase prior to the CIL circuit. The slurry gravity flows to the first stage CCD thickener mix tank where it is diluted with overflow from the downstream CCD thickener. Overflow solution is sent to the autoclave flash steam quench vessels where it is used to condense and scrub excess steam before going to the ferric precipitation reactors. The balance of the overflow solution is fed directly to the ferric precipitation neutralization circuit ahead of copper recovery. The underflow slurry is pumped to the flash steam preheating vessels in the autoclave area prior to discharge to the silver enhancement lime boil circuit.

The nominal wash ratio in the CCD circuit is maintained throughout the tonnage expansion rate program to yield a wash efficiency of 99.0% to 99.5% at design conditions.

17.2.4.9 Ferric Iron Precipitation and Copper Recovery

The copper recovery circuit will use the hydrogen sulphide precipitation technology to precipitate the contained copper in the CCD wash solution as a copper sulphide concentrate. The process uses bacteria to convert elemental sulphur to H₂S gas, which then reacts with copper ions to precipitate copper sulphide (CuS).

Initially, the CCD overflow solution is partially neutralized with limestone in a series of two mechanically agitated, stainless steel precipitation tanks. Upon expansion to 24,000 t/d an additional two neutralization precipitation tanks will be added to the circuit. The discharge from the second limestone neutralization tank is directed by gravity to the ferric thickener in the copper recovery plant. The pH is closely controlled in the first two precipitation tanks to precipitate ferric iron from solution while minimizing the amount of copper co-precipitation.


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The iron precipitate is gravity fed to the ferric thickener, sized to handle the full flow of 24,000 t/d. The thickener underflow sludge is pumped to the neutralization circuit for completion of the neutralization process.

The ferric thickener overflow solution is pumped to the copper contactor circuit where it is contacted with H₂S gas. Three copper contactors, to handle the 24,000 t/d production rate, will be mechanically agitated, closed-top tanks to ensure adequate mixing and gas liquid mass transfer. The H₂S gas is produced by a bacterial process, which uses sulphur reducing bacteria to convert elemental sulphur into H₂S under anaerobic conditions. The two bioreactors are gas-lift loop type reactors that allow the generated H₂S gas to be drawn off the head space of the bio-reactor unit and compressed by gas blowers. The compressed gas stream, containing 8% to 10% volume H₂S, is sparged into the copper contactor vessels. The barren H₂S gas returning from the contactors, saturated with water is dewatered in a condensate knockout stage and returned to the bio-reactor.

The precipitated copper sulphide solution is degassed and fed to a 50 m diameter thickener clarifier for solids removal. The underflow is pumped to a secondary dewatering step. The sulphide filter cake is discharged onto a conveyor that delivers the concentrate to a bagging facility. Bagged concentrate will be containerized and delivered by flatbed trucks from the plant site to a port near Santo Domingo.

The copper clarifier overflow solution is pumped to the high density sludge neutralization circuit.

All of the tank head spaces containing H₂S are connected to a common header to capture and control fugitive emissions. The vapour passes through a condensate trap and emergency scrubber unit. It is then compressed by the blower and re-injected into the bio-reactor vessel.

17.2.4.10 High Density Sludge (HDS) Neutralization Circuit

Neutralization of remaining acidity and the precipitation of metals and sulphate in the CCD overflow are accomplished in the HDS neutralization circuit. The HDS neutralization circuit will comprise a three stage limestone addition followed by two stages of lime treatment to treat the 18,000 t/d production rate. An additional reactor will be added to both the limestone and lime circuit during the expansion to 24,000 t/d.

The limestone and lime reactor tanks will be arranged in a staggered cascading fashion, with step-down elevations along the train to enable gravity flow. Limestone slurry is metered into a mix tank where it is blended with some recycled HDS thickener underflow to condition the recycled material and promote the HDS precipitation seeding process. The mix tank overflows into the first neutralization tank and mixes with the cooled copper clarifier overflow solution and ferric thickener underflow product stream.

The neutralized slurry gravity flows from the final lime neutralization tank to the HDS thickener. The HDS thickener underflow is pumped to the tailings pond via the cyanide destruction tailings pump box. The overflow solution is directed to the HDS thickener overflow tank and pumped to the HDS solution cooling towers.



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17.2.4.11 HDS Solution Cooling

HDS thickener overflow solution is pumped to a bank of 6 cooling towers, with an additional pair added to accommodate the expansion to 24,000 t/d, where the temperature is reduced. The actual cooling requirements are determined by the heat balance. The cooled solution is pumped to the process water tank. It is then distributed for use as CCD wash, limestone grinding and flocculant dilution.

17.2.4.12 Silver Enhancement Lime Boil Process

The CCD circuit thickener underflow is pumped to the lime boil preheating vessel where it is reheated to 95°C using steam from the autoclave flash tanks. The reheated slurry is treated with lime to effectively break down the silver jarosites formed during the pressure oxidation and hot cure processing stages. This maximizes silver extraction during CIL leaching.

The lime boil circuit, installed to treat 24,000 t/d, will consist of 2 agitated tanks. The lime boil slurry is then cooled to approximately 40°C in 4 slurry cooling towers, arranged in parallel, with an additional unit added to treat the 24,000 t/d production rate. Cooled slurry is pumped to the CIL circuit where gold and silver are extracted using cyanide and activated carbon.

17.2.4.13 Carbon-in-Leach (CIL) Cyanidation

A CIL circuit was selected to maximize gold and silver extraction from preg-robbing carbonaceous ore sources in the deposit.

The cooled lime boil discharge is screened for the removal of trash and fed to the first of eight agitated tanks providing a retention time of approximately 20 h. For the expansion to 24,000 t/d, an additional three CIL tanks will be added to maintain the optimal leaching conditions.

A carbon loading of 2,000 g/t gold is projected from the pilot plant tests when processing ore grading 5.0 g/t gold or better. This requires an initial carbon advance rate of 48 t/d, rising to 72 t/d for the 24,000 t/d ore production rate case.

The pilot plant tests indicated average gold and silver extractions in the CIL circuit of 92% and 85% respectively. The average cyanide addition is estimated at 1.0 kg/t of CIL feed.

17.2.4.14 Cyanide Destruction

The average total cyanide level in the CIL discharge is estimated at 150 mg/L. The SO₂/air process was selected based on the results of pilot plant cyanide destruction testwork conducted by Inco Technology.

The WAD CN total cyanide and copper levels in the treated effluent produced during the 2005 pilot plant test were slightly greater than 1 mg/L. An updated laboratory evaluation of the 2006 pilot plant CIL product showed good effluent quality and met the target levels of less than 1 mg/L.



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17.2.4.15 Carbon Acid Washing and Stripping

Twelve tonne batches of loaded carbon from CIL Tank #1 are acid washed with dilute hydrochloric acid and rinsed with water before being stripped using the Pressure Zadra elution process. The pregnant solution gravity-flows to the pregnant solution tank from where it is pumped at a controlled rate to the EW circuit.

The stripped carbon is thermally reactivated in one of two electrically heated horizontal furnaces at a temperature of 700°C and at a rate of 1,000 kg/h. An additional furnace will be installed during Phase 2 expansion to handle the increased carbon tonnage. The kiln exhaust gases vent through a wet scrubber and subsequently pass through columns packed with sulphur-impregnated carbon to remove mercury.

The reactivated carbon is screened to remove carbon fines before being returned to the last CIL tank to replace the forwarded carbon. The fine carbon is forwarded to a settling pond and periodically recovered and bagged for sale.

17.2.4.16 Electrowinning (EW)

The pregnant solution or eluate is pumped from the pregnant solution tank to five parallel electrowinning (EW) trains. All EW cells will be provided with a gas extraction system connected to a mercury capture system.

Gold and silver, along with some impurities (mainly copper and mercury), are plated onto the punched stainless steel plate cathodes. The barren electrolyte, containing less than 2 g/t of gold, flows by gravity to a collection tank from where it is pumped to the barren solution storage tank for recycling to the elution circuit.

The EW cells will use non-basketed cathodes within the sludge-type EW cells to allow for high pressure cathode washing of the gold and silver sludge within the cell unit. The cells will be taken offline periodically for harvesting of cathodes and recovery of gold and silver. The resulting gold and silver sludge will be filtered in a plate and frame filter.

17.2.4.17 Refining

The gold/silver sludge from the EW circuit may contain up to 5% mercury. If not removed, it will volatilize during smelting and report to the off-gases. To comply with the air quality standard, a mercury retort will be provided to remove and recover the mercury from the sludge before smelting.

The filtered sludge is loaded into boats, dried, and heated in an electric mercury retort, which is kept under vacuum, to remove the contained mercury. The gas stream containing the volatilized mercury passes through a water-cooled condenser to condense the mercury for collection into flasks. Sulphur-impregnated carbon columns remove residual mercury from the discharge gas of the condenser.

The mercury-free sludge is fluxed and smelted in one of two induction furnaces to produce 1,000 oz bullion bars. The anticipated increase in sludge produced from the expansion to 24,000 t/d will require the installation of two further induction furnace units. These furnaces



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will be provided with a dust collection gas baghouse system to recover the gold and silver-laden dust generated during smelting, and clean the furnace off-gases before discharge to the atmosphere.

17.2.4.18 Tailings Disposal and Tailings Water Reclaim

(See also Section 17.6 Environmental Considerations.)

The detoxified leach residue is combined with the sludge from the neutralization circuit for disposal to the El Llagal tailings storage facility. Containment earth berms will be installed alongside the tailings pipeline. Any spillage will be directed toward and stored in the collection ponds.

The tailings pumping system will consist of two dual stage slurry pump trains (one operating unit and one standby unit) with variable speed drives to regulate the discharge head to match with the gradually rising tailings embankment height. Tailings will be distributed with spigots across the tailings embankment towards upstream side of the storage pond. Additional spigots will be provided to discharge along the eastern and western sides to create a small, supernatant pool in the middle of the storage pond.

From the supernatant pool, a reclaim pump barge will pump the tailings water to the process plant. Recycling of this tailings water to the process will only be implemented under extreme drought and flood conditions because of the negative impact of chloride ions in the reclaim water on gold extraction. Within the catchment of the tailings impoundment, any spillage from the reclaim pipeline will drain into the tailings impoundment. Beyond that, the reclaim pipeline will be installed alongside the tailings pipeline, inside the common containment berms for spillage control.

Seepage from the tailings storage facility will be collected in a small pond in front of the main containment embankment. A pumping and pipeline system will be installed to return the seepage to the impoundment.

In addition to the tailings, the El Llagal pond will also be used to store potentially acid-generating mine waste rock. The material will be trucked to the storage site by way of a haul road. To prevent ARD formation, the waste rock will be kept submerged.



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17.2.5 Limestone and Lime Plant Description

17.2.5.1 Design Basis

The limestone and lime plant design is based on the following estimated reagent requirements shown as Table 17.26.


Table 17.26		Limestone and Lime Plant Design Basis (Expansion 24,000 t/d Ore Processing Rate)


	Limestone		Lime
	(t/d)		(t/d)
Process Including Neutralization		4,965		1,245
ARD (1 in 200-Year Flood Event)		1,649		146
Tailings Effluent (1 in 200-Year Flood Event)				19
Subtotal (Uncorrected For Purity)		6,614		1,410
Limestone Feed to Kiln		2,300
Total (Corrected for Purity)		8,914		1,484
Design:
Limestone Crushing		9,240
Limestone Grinding		9,000
Lime Slaking				1,484

17.2.5.2 Flowsheet

Ground limestone and lime are required to neutralize acidic liquors and to control the pH in the CIL circuit. Lime is also used to adjust the pH of the effluent after water treatment. Satisfying the 24,000 t/d ore process requirement requires grinding 9,070 t/d of limestone to 80% passing 60 μm and calcining 2,785 t/d of limestone in vertical kilns to produce 1,484 t/d of lime, all of which will be slaked in a ball mill slaker. The proposed limestone plant will consist of the following unit operations: primary crushing and screening, grinding, calcining, and lime slaking.

Primary Crushing and Screening

Run-of-mine (ROM) limestone is crushed to minus 85 mm (P₈o) in a gyratory crusher equipped with a rock breaker to break oversize rocks in the dump pocket. Provision is included for the future installation of a dust control system at the primary crushing station if required to reduce fugitive dust emission. The configuration of the limestone crusher is similar to that for the ore. The crusher product is screened and the +50 mm -110 mm intermediate fraction is sent to the kiln circuit for calcination. The balance of the crusher product reports to the limestone SAG mill feed stockpile.

Grinding



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The limestone grinding circuit will consist of a SAG mill (6.70 m dia x 3.65 m EGL) driven by a 2,610 kW synchronous motor with VFD and a ball mill (4.88 m dia x 9.80 m EGL) driven by a 3,540 kW synchronous motor. The SAG mill will be in open circuit while the ball mill will be in closed circuit with a cluster of sixteen hydrocyclones. The limestone slurry is pumped to three agitated storage tanks holding approximately 6,500 t of limestone. This provides 22 h of storage capacity at peak limestone demand.

Limestone Calcining and Lime Slaking

The lime calcining plant will be designed to process 2,785 t/d of limestone to produce 1,484 t/d of lime required for the ore production rate of 24,000 t/d. The high lime requirement and the availability of high quality limestone deposits near the mine justify the installation of the lime plant.

Three 550 t/d vertical twin-shaft parallel flow regenerative (PFR) lime kilns were selected because of their efficiency. The kilns will be fed with +50 mm -110 mm intermediate screen product produced from the screening circuit.

Lime is slaked at a rate of 1,484 t/d in a ball mill operating in closed circuit with one hydrocyclone to produce a hydrated lime slurry. The lime slurry is pumped to four agitated storage tanks and is distributed from these tanks via lime loops to the lime boil and neutralization circuits, and to the effluent treatment plant.

17.2.6 Process Risk Summary

17.2.6.1 Process Design

The proposed process design for Pueblo Viejo would make it one of the most complex, large precious metals projects in operation. Considerable effort has gone into process development with input from some of the most competent and highly qualified professionals in the respective fields.

Pressure oxidation at 230°C is a very aggressive technique for refractory sulphide mineral processing, designed for rapid liberation of ultra-fine gold encapsulated within the sulphide minerals, and to minimize the preg-robbing capacity of the active carbon present in some ore types.

The extensive bench-scale and pilot plant testwork programmes undertaken in support of process and flowsheet development have been competently designed and executed, although some ‘end-to-end’ risk remains, associated with the linkage of the complex unit processes, some of which have not been used within the gold mining industry.

17.2.6.2 Equipment Selection

The equipment selection process in terms of quality and vendor capability has been rigorous with a high level of performance benchmarking. Nevada pressure oxidation experience and best practice has been incorporated into the Pueblo Viejo design. The autoclaves will have the same diameter as autoclaves in use in Nevada and conventional, well proven materials of construction are included in the design.



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17.2.6.3 Sulphur Grade

Given the very strong dependency of process plant throughput and operating cost on sulphide sulphur grade, accurate modelling of the sulphide sulphur grades, and ore scheduling to maintain optimal sulphur levels in the process plant feed, are important considerations. The Pueblo Viejo deposits exhibit a relatively high level of sulphur grade variability so effective management of a large number of plant feed stockpiles has been recognized as being an important issue.

17.2.6.4 Scale Formation

Scale formation was evident to varying degrees in all the autoclave pilot plant trials. The issue of scaling potential has been benchmarked with other pressure oxidation plants. Although scale formation has also been experienced in autoclave pilot plant testwork for other operating mines it has not proven to be a significant problem in practice. The risk of scaling will be elevated however given the slightly higher than typical autoclave operating temperatures in the Pueblo Viejo design. Should autoclave operation at lower temperatures prove necessary, increased preg-robbing may result in slightly lower precious metals recoveries for some ore types. Provision has been made to lessen the impact of scaling in other areas of the process plant through equipment redundancy and labour resources in the form of a dedicated de-scaling crew.

17.2.6.5 Process Plant Capital and Operating Costs

A high level of detail has gone into the capital cost development for each of the process equipment line items. The majority of the direct costs for the major equipment items are based on firm quotes with little use of factored costs. Labour costs are based on extensive labour market and productivity studies and indirect costs such as freight have been calculated from first principles using quoted rates for the project. Some allowance has been made for cost escalation, thus minimizing escalation risk save for a modest level of foreign exchange risk exposure.

Operating costs have also been developed to a high level of detail. Labour market studies were undertaken with input from local employers and maintenance costs were benchmarked against other pressure oxidation operations. The overriding consideration, however, is the pressure oxidation process that is very energy intensive so the sourcing of reliable, relatively inexpensive power is critical to the economic viability of the project.

17.2.6.6 Metal Recoveries

All the hydrometallurgical processes to be deployed at Pueblo Viejo have been extensively trialed at both bench and pilot plant scale and the projected metal recoveries reasonably reflect the testwork results.



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17.3 Infrastructure

17.3.1 General Infrastructure

The main road from Santo Domingo to within about 22 km of the mine site is a surfaced, four-lane, divided highway in generally good condition. Access from the main road to the site is via a two-lane road, surfaced for part of the way but gravel for the remainder. Construction of the Pueblo Viejo Project will require some upgrading of both the main and secondary roads, and particularly to accommodate trucks carrying heavy loads. The heaviest and largest items to be transported will be the autoclaves at over 700 t each. The upgrading will cover road and bridge improvements, clearing of overhead obstructions, erosion control, bypass route construction, clearing utility interferences and work permitting. Gravel surfaced, internal access roads provide access to the mine site facilities. A network of haul roads will be built to supplement existing roads so that mine trucks can haul ore, mine overburden, and limestone from the various quarries.

As well as the existing access roads, current site infrastructure includes the previous processing plant, accommodation, office and other buildings, water supply, and old tailings impoundments with some water treatment facilities. Some of these facilities will be either upgraded or renovated, as will be the case for some accommodation and other buildings, or totally demolished as in the case of the old processing plant.

The new process plant site will be protected by double and single fence systems. Within the plant site area, the freshwater system, potable water system, fire water system, sanitary sewerage system, storm drains, and fuel lines will be buried underground. Process piping will typically be left aboveground on pipe racks or in pipe corridors.

17.3.2 Power Plant

In the Barrick FS, power supply for the mine and process plant is indicated as being provided by a third party from a new, coal-fired power plant to be built on the south coast of the Dominican Republic. Distribution will be via a dual circuit 230 kV transmission line (111 km) along a corridor from the source to the mine site. The cost of building the transmission line is included in the construction cost of the Project as PVDC will be the owner of the line.

Power cost has been assumed at US$0.10/kWh during the first year of operation, and US$0.08./kWh thereafter in the operating cost model.

Information from Barrick subsequent to the Barrick FS indicates that power may not now come from a coal-fired plant but, possibly, from a combination of a new and existing HFO generation facilities. The owners recognize this item as one of the major project uncertainties, both in terms of the source of supply and its operating (and, possibly, capital) cost implications.

17.3.3 Site Electrical System

The plant will receive power from the generating plant via a dual incoming 230 kV power line. The power lines will be terminated at the mine site main substation where they will be



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fed into a single 230 kV bus system. Four main transformers will provide power for all site loads, with two being dedicated to the oxygen plant.

The average power demand at start-up of the 18,000 t/d plant will be 150 MW. The average power demand at start-up of the 24,000 t/d plant will be 195 MW.

If the supply of normal power is interrupted, the plant will operate on emergency feed. This is provided by 15 MW of diesel generation that connects to the main substation for distribution to critical areas such as lighting, communication, and computer and process equipment.

17.3.4 Process Control Facilities

The plant wide distributed control system (DCS) will use Ethernet communication links, fibre optics, Foundation Fieldbus for analog devices, conventional controls for discrete devices, and radio-links for remote sites. Three main control rooms, 13 satellite control rooms, and 3 maintenance workstations will be located throughout the site.

17.3.5 Communication Facilities

A redundant fibre communication backbone of approximately 40 km around the mine site will link and manage the data transmission of the DCS, third party PLCs, motor controls, fire detection system, Vo-IP telephone system, and computers.

17.3.6 Fuel

Two permanent fuelling stations will feed the fleet of mine vehicles. A permanent heavy fuel oil (HFO) storage will supply the lime kilns. New tanks and fuel stations will be installed for fuel storage during construction.

17.3.7 Water Supply

The Hatillo and Hondo Reservoirs will supply fresh water to the site. Reclaimed water from El Llagal tailings containment pond will only be used as a supplementary water supply under drought and flood situations. Barge-mounted pumps at the larger Hatillo Reservoir will pump fresh water to the Hondo Reservoir for make-up purposes. Fresh water will then be pumped to a freshwater/fire water tank at 400 m level and a freshwater pond, and from there will be distributed throughout the site for process, fire protection and potable needs. The potable water will be a treated system.

Initial water for earthworks and construction will be supplied largely from the Maguaca River, but also from the pipeline that connects the Hondo Reservoir and the freshwater pond. Potable water for construction offices, dining rooms, toilets, and use mainly at the plant site, will be supplied during construction from a temporary tank located north of the oxygen plant. Potable water will be delivered by trucks to another potable water tank located at the south side of the plant site.



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17.3.8 Storm Water

The plant site is located on a ridge between two drainage catchments. Where possible, runoff from the process plant will be directed to the Margajita drainage area to separate it from the storm water runoff from the old facilities. Where this is not practical, a collection pond will capture the runoff before it is returned to the process plant to serve as make-up water.

17.3.9 Waste Management

An underground, gravity sewer system will collect domestic waste water from the various site areas. Separate, underground, gravity systems will be built to serve the construction and operations camps. Clean effluent will be discharged to the local river system. Non-hazardous domestic solid waste will be sent by truck to a central handling facility. An incinerator will be installed at the non-hazardous waste dump to burn solid waste. Barrick has contracted a specialized consultant to compile an inventory of hazardous material left by the previous operation and to recommend methods for its disposal. Cost for removal of past hazardous waste remains a responsibility of the Dominican Government.

17.3.10 Sewage Treatment

The proposed sewage treatment configuration is based on three 280 m³/d plants at the construction camp, one 280 m³/d plant at the plant site, and one 61 m³/d plant for the houses. The system is based on a 3-part modular arrangement: primary settlement tank, biological treatment unit with biological rotating contactor, and final settling tank.

17.3.11 Fire Protection

Fire protection will be provided by a variety of measures, including fire walls, hose stations, automatic sprinkler systems and yard hydrants. A fresh water/fire water tank will supply fire water to the site. Buried fire water pipes will distribute fire water to protected areas.

17.3.12 Dust Control

The dust control system for the refinery furnace will be a scrubber; baghouses are assumed to be the primary technology for dust control but other options will be examined as part of the value engineering exercise during detailed engineering. Based on the individual process, dust control measures will also include water sprays and fogging systems.

Dust control on roads will include watering and use of brine solutions.

17.3.13 Landfill

Areas for hazardous and non-hazardous material storage during demolition and construction will be provided on the beaches of the Mejita tailings pond.

Landfills for historical hazardous waste (Dominican State responsibility) are proposed east of the Mejita tailings pond. Non-hazardous material will be stored in an area south of the Mejita dam for removal at a later date.



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17.3.14 Las Lagunas Project

Las Lagunas Ltd was granted a limited project approval in December 2006 for the Las Lagunas development area, including the tailings impoundment facility, the limestone quarry to the northeast, the borrow-material area to the southwest, the area for facilities to the south of the dam, and other areas such as access from the main road, etc. Also, the Directorate of Mining granted Las Lagunas Ltd. the right to exploit the limestone quarry for the neutralization process.

17.3.15 Resettlement Action Plan (RAP)

Construction and operation of the Project will result in the physical and economic displacement of households from three geographic areas, called Displacement Zones, within the overall Project Development Area. The three Displacement Zones total 1,104 ha in area with 369 households being affected. At the request of the government, Barrick provided support for the preparation of a Resettlement Action Plan. Barrick funded the assistance of expert technical personnel, local consultants, and local personnel. Approximately $1.5 million was spent in giving over a year of support to the government in the preparation of a RAP ready to implement.

On September 25, 2007, the RAP was approved and signed by the representatives of the three communities, the Dominican State, Barrick, and the Catholic Church. The last two parties participated as observers and ensured that the process followed IFC guidelines.

17.3.16 AMC Comments

Uncertainty about the source of power is a major project issue that obviously must be addressed.

Other items that may not affect project viability, but that could still cause project delays, are the Las Lagunas conflict and potential delays in permitting.

The Barrick proposed role in cleaning up existing infrastructure issues, and in particular with respect to existing ARD drainage, must be seen as a very positive aspect of the overall Barrick involvement in the area.

The Resettlement Action Plan has reportedly been well received by those affected and does not appear to be a material concern to the project. The resettlement is the responsibility of the Dominican Republic and has to be performed following World Bank Guidelines.



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17.4 Markets

17.4.1 Metal Prices

Metal prices used for the economic model are as follows:


• Gold		$700/oz

• Silver		$13.00/oz

• Copper		$2.75 Year 2011, $2.50 Year 2012, and 2013, $2.00 for remainder of mine life

17.4.2 Doré Shipping and Refining


• Gold payable %		99.925

• Silver payable %		99.00

• Refining $/oz Doré		$0.70

• Transport $/oz Doré		$0.40

17.4.3 Copper Concentrate Shipping and Refining

The copper concentrate will be stored in 3 t bags, trucked to the port Haina, accumulated until there is enough for shipment, and then loaded for transport via ship to a receiving port either in North America or South America.


• Copper payable %		96.5

• Refining charge $/lb		$0.075

• Treatment charge $/t		$75.00/t concentrate

• Price participation		10% of copper price difference above $1.20/lb up to a
		maximum of $0.10/lb copper

• Transport		$105.00/t of concentrate

17.5 Contracts

Goldcorp and Barrick are two of the world’s major gold mining companies. Terms of mining, concentrating, smelting, refining, transportation, handling, sales and hedging and forward sales contracts or arrangements, rates or charges are within industry norms.



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17.6 Environmental Considerations

17.6.1 Scope

The environmental assessment included a review of the existing conditions at the mine, environmental baseline studies undertaken, and plans for future development including construction, operation and closure of the mine. A site visit was carried out on March 18 and 19, 2008 to observe current conditions and to interview personnel involved with the environmental programs at the mine site.

17.6.2 Authorizations and Responsibilities

Several approvals, permits and licences are required prior to the commencement of construction and operation of the Pueblo Viejo Mine. The Dominican Republic Environmental Law No. 64/2000 provides the process for obtaining environmental permits. In addition, there are several sectorial permits that will be required from different agencies. Permit status for the Pueblo Viejo project as of December 2007 has been shown earlier in Section 3.4.

The principal agencies are:

•		Secretaria de Estado de Medio Ambiente y Recursos Naturales — SEMARN (Ministry of Environment).

•		Instituto Dominicana de Recursos Hidráulicos — INDRHI (Water Resources)

•		Secretaria de Estado de Industria y Comercio — SEIC (Ministry of Industry and Commerce)

•		Subsecretaria de Recursos Forestales — SFR (Under-Secretary of Forestry Resources)

•		Secretaria de Estado de Obras Públicas — SEOP (Public Works)

•		Directión General de Mineria (DGM) — (General Mining Agency)

•		Ayuntamiento (Municipalities)

Three types of permits are required, including:

•		those that must be obtained prior to the environmental licence

•		those included in the environmental licence

•		those needed after receipt of the environmental licence.

The licences or agreements discussed in the remainder of this section have been obtained, or were in the process of being finalized, as of December 2007.

17.6.2.1 Special Lease Agreement

Conditions for land use and environmental responsibilities are covered by the Special Lease Agreement (SLA) (see also Section 3.2). The Special Lease Agreement of Mining Rights for the Montenegro Fiscal Reserve (MFR) was negotiated between the Dominican State, the



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Central Bank of the Dominican Republic, Rosario and PDDC., a subsidiary of Placer, anc was approved on July 29, 2003. Barrick purchased Placer in January 2006 and thereby became owner of the PDDC. Barrick sold a 40% interest in the company that owns PDDC to Goldcorp. In December 2006, PDDC was renamed Pueblo Viejo Dominicana Corporation (PVDC)

On August 3, 2004 the MFR was modified to include the area of the El Llagal tailings and waste rock storage facility. A Project Notice, committing PVDC to construct and operate a mine at the Pueblo Viejo site under the terms of the SLA was delivered in February 2008. The Project Notice was conditionally accepted by the Dominican Government. The environmental considerations contained in the SLA include Section 7.2 — Remediation of Historic Environmental Matters, and Section 11.2 - Assignment of Responsibility for the Management of Remediation of Environmental Conditions.

PVDC is responsible for all historic environmental matters within the boundaries of the Development Areas designated in the Feasibility Study (see Figure 17.14). The main areas include the Montenegro and Moore open pit areas, the Los Quemados, Las Lagunas and Planta limestone quarry areas, the plant site, two acid rock drainage (ARD) storage ponds, camp site, El Llagal tailings and waste rock storage impoundment, and areas reserved for extraction of construction material and additional limestone requirements. PVDC is not responsible or financially liable for remediation, rehabilitation or mitigation of the historic environmental matters outside the boundaries of the Development Areas.

The terms of the SLA require the Dominican State to mitigate all historical environmental matters, except those matters within areas designated for development by PVDC in the Project Notice. The Dominican State is also responsible for the relocation of people living within the development area, including the proposed El Llagal tailings and waste storage area, in accordance with the Environmental and Social Guidelines and Policies. PVDC is not required to pay for the relocation or for land required by PVDC pursuant to the SLA.

On May 5, 2006, an Australian company Las Lagunas Ltd was granted the development area for the Las Lagunas Project including the tailings impoundment facility and adjacent borrow areas including the limestone resource to the northeast and access from the main road. It is reported that various aspects of the PVDC SLA, including legal issues, were not followed and that the Pueblo Viejo Project is in a position of conflict with the Las Lagunas Project. The conflicts are associated with unresolved environmental concerns regarding construction, operation and closure of Las Lagunas and overlap of development areas, clear definition of the development area, conflicts regarding access to resources and project infrastructure, and availability of limestone deposits.



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Figure 17.14 Development Areas of Las Lagunas

17.6.2.2 Resettlement Action Plan

(See also Section 17.3)

A Resettlement Action Plan (RAP), prepared for the government with the support of PVDC and with assistance from expert technical personnel, local consultants and local personnel, was developed in accordance with World Bank Standards. The RAP was approved and signed on September 25, 2007 by representatives of the three local communities affected by the plan, the Dominican State, PVDC and the Catholic Church.

17.6.2.3 Memorandum of Understanding

PVDC and the Dominican State signed a Memorandum of Understanding (MOU) on November 30, 2007 that covers funding for resettlement of households under the RAP, acquisition of land and mitigation of the various historical environmental liabilities. The MOU facilitates the advance of funds by PVDC to resolve the historic environmental and social liabilities that under the SLA are the government’s responsibility and requires the government to reimburse PVDC for the funds advanced.


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17.6.2.4 Environmental Licence

An Environmental and Social Impact Assessment (ESIA) was submitted to the government on November 21, 2005. Following various meetings and workshops, and upon conclusion of the government process of review and evaluation, the ESIA and the environmental management plan (EMP) were approved by the Secretariat of State for the Environment and Natural Resources on December 26, 2006 and the Environmental Licence No. 0101-06 was issued January 2007. Conditions of the Environmental Licence require submission of detailed designs for the tailings dams, installation of monitoring stations and submission for review of the waste management plan and incineration plant design.

An updated EMP including silver/copper recovery was submitted on September 30, 2007. AMC understands that formal discussions about this plan are ongoing.

17.6.2.5 Demolition Permit

The permit for demolition of the existing plant site facilities was issued on November 21, 2007.

17.6.2.6 Other Permits or Authorizations

See Section 3.4 of this report.

17.6.3 Environmental Standards

Standards for the regulation of water, effluents, air, noise, hazardous waste, and domestic waste were published by SEMARN on June 5, 2003. These standards set limits for different elements and establish waste management procedures.

17.6.4 Existing Environmental Conditions

The historic Pueblo Viejo mine operated prior to June 1999. Previous development included the mining of two main pits (Monte Negro and Moore) and several smaller pits, construction of a plantsite, and construction of two tailings impoundments (Las Lagunas and Mejita). Waste rock dumps and low grade ore stockpiles from these operations are located throughout the pit areas. Facilities from previous operations include water supply, power station, housing, office buildings and two acid rock drainage treatment plants.

17.6.4.1 Acid Rock Drainage

The major environmental issue at the mine site is related to acid rock drainage (ARD) that has impacted the surface water at various locations. The ARD has developed from exposure of sulphides occurring in the existing pit walls, waste rock dumps and stockpiles to air, water and bacteria. Untreated and uncontrolled ARD has contaminated local streams and rivers and has led to deterioration of water quality and aquatic resources both on the mine site and offsite.


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The ARD impacts on surface water and groundwater within the project area have developed from historical mining operations and are described in the feasibility study report. There have been attempts to treat the ARD water; however the water treatment plant above the Las Lagunas tailings storage facility (TSF) drainage is no longer operating and the water treatment plant above the Mejita TSF drainage appears to be only partially effective.

17.6.4.2 Tailings Storage

The Mejita and Las Lagunas tailings storage facilities were constructed during the previous mine operation. They are the responsibility of the DR government. It is reported that uncontrolled seepage has occurred from these impoundments since they were commissioned and that the geotechnical stability of the earth embankment dams and foundation suitability is questionable. The dam stability concerns are referenced in the Barrick FS report as part of the historical environmental impacts. Barrick has reviewed this issue and offered a conceptual design solution.

Discussions between PVDC and the Dominican Republic government were in progress as of December 2007 regarding runoff from the Mejita tailings impoundment, Cumbra Pit and Arroyo Hondo drainages that are the responsibility of the government and outside the defined development areas.

17.6.4.3 Hazardous Waste

Large amounts of hazardous waste materials have been identified on the mine site. The waste products include rusting machinery, hydrocarbon contaminated soils, mercury contaminated materials, asbestos, and tailings that have escaped into neighbouring watersheds. A mitigation and management plan has been prepared and will be implemented following approval from the Dominican Republic government.

17.6.5 Environmental Baseline Studies

PDDC commissioned a number of consultants to collect background data and baseline information on the existing biophysical and human environments from 2002 through 2007. The baseline studies covered the immediate project areas and also areas beyond the mine site.

Acid rock drainage (ARD) studies confirm that historic mining and current ARD generation within the mine site have severely impacted the surrounding area. Test results indicate that most of the exposed rock at the mine site is acidic and contains significant sulphide levels providing a source for additional acidity. The acid rock drainage sources include the Montenegro and Moore open pit walls, seepage from the tailings storage facilities and seepage from the waste rock dumps and low grade ore stockpiles. The Arroyo Margajita is impacted by releases of treated water and treatment sludge followed by extended periods of untreated ARD releases. Tests have been performed to determine acid generation from existing and new waste rock, acid generation from the pit walls, and parameters for limestone and lime consumption. Tests included the operation of a high density sludge (HDS) pilot plant, column tests using fresh waste rock, humidity cells using existing waste rock, field measurements and water sampling.


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Air quality baseline studies included collection of particulate matter less than 10 μm in size in the town of Maimón and at the Pueblo Viejo camp. The results indicate that the concentration of PM10 particulate matter is in compliance with the Dominican Republic daily standard. Monitoring for the new standard for particles of less than 2.5μm in size has not been undertaken. Climate and meteorology baseline studies have also been completed.

An archaeology study has been undertaken in the area of Pueblo Viejo Mine, El Llagal and the area towards the Hatillo Reservoir. The study identified 10 sites with signs of past activity of which four are located within the project boundaries. An additional archaeological site was identified in 2007. Mitigation plans are to be implemented.

The results of aquatic biology studies undertaken in local streams and the Hatillo Reservoir indicated that the health characteristics of stream invertebrate communities were higher at the Maguaca River stations relative to the Margajita River. No fish were found in the Margajita area and fish habitat is highly degraded. The absence of small fish in the Maguaca River is indicative of historical mining impacts. None of the species captured are on the International Union for the Conservation of Nature (IUCN) Red List. Fish tissue tests indicate metal concentrations were well below the Canadian Government benchmark for arsenic, lead, and mercury and therefore consumption of fish is not a risk to humans for these elements.

Terrestrial biology vegetation and fauna baseline studies have also been performed. The study area included local streams, the Hatillo Reservoir and Piedra Imán. Little vegetation cover was found in the pit areas and most of the surrounding area is forested. One vegetation species found within the Pueblo Viejo area is protected by the IUCN Red List. This species was introduced as part of the mine site reclamation. Other species are protected by draft local regulations. None of the twenty-two mammals identified during the baseline studies are listed in a protected category. Thirteen species of amphibians and thirteen species of reptiles were recorded. Based on local regulations five of the reptile species are considered threatened due to loss of habitat, hunting and impact from introduced birds. Three species of the sixty-six bird species identified are protected by local regulations. They are classified as vulnerable on the ICUN Red List due to loss of habitat, hunting and impact from introduced birds.

International consultants and Barrick personnel carried out studies for geology and geochemistry in the area of El Llagal, the mine site and Hatillo Reservoir. In the El Llagal area and the north western area of Hatillo Reservoir there was no evidence of materials with significant potential for acid generation. Sediment samples collected from the upper Maguaca, lower El Llagal and Lower Naranjo streams indicated low total sulphur content. Higher total sulphur values were found in the western area of Hatillo Reservoir, however the stream water has neutral pH indicating acid generation is not occurring. In the area of the Las Lagunas tailings facility both acid generating potential and neutralization potential was found in collected samples. Rock at the plant site area was found to have acid-generating potential. A soil geochemistry survey was undertaken to determine existing metal levels. Tests were completed for iron, arsenic, mercury, zinc, lead, cadmium and gallium.

Wastes from previous mining operation are located throughout the mine site. In accordance with the SLA the Dominican Republic is responsible for the hazardous waste cleanup, disposal and remediation. A survey of existing hazardous waste and a mitigation program for controlled disposal of the waste have been completed. The plans for waste management


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and estimated cost to implement phase 1 of the program are currently under consideration by the Dominican State government.

Hydrology conditions in the area have been studied. Surface water flows at the mine site and on the Arroyo Margajita, Rio Maguaca and Arroyo El Llagal indicate highest flows and runoff could occur between April and December with lowest flows occurring typically between January and March. Minimum flow rates were established and peak instantaneous flows were estimated. This work will be used to develop water management plans for the mine site. Stream flow measurements were obtained at several locations.

Twenty wells were drilled for hydrogeology baseline studies around the mine site and in the area of El Llagal, Maguaca and Margajita. Groundwater samples indicate that groundwater contamination is limited to the area of the Cumba Pit draining towards the Arroyo El Rey and Maguaca River, and to the area of Monte Negro. Groundwater draining towards the Cañada Hondo and Marguaca River from the Mejita tailings appears to be neutralized by the Hatillo limestone formation. In the area of the Moore and Monte Negro pits the groundwater is contaminated with acid and trace metals.

Twenty five streams have been sampled as part of baseline studies of surface water and sediment characterization. Water quality sampling is continuing at specific sites as part of the continuous monitoring program. The studies indicate that the Margajita River and Arroyo Hondo have been most severely affected by acid generation. The northern tributaries of the Margajita River have naturally low pH and low conductivity and the southern tributaries have neutral pH. Metal loadings in the Hatillo Reservoir water are highest close to the Margajita River inflow but the sediments show higher metal content closer to the dam located downstream. The upper area of the El Rey River was found to be affected by contaminants from the Cumba Pit. The Maguaca River has been slightly affected by mine site contaminants that potentially affect the lower Yuna River. Elevated sulphate concentrations in the Yuna River at the confluence with the Margajita River are slightly lower than Dominican Republic standards. Good quality water and sediment were found in the El Llagal area.

Wetland characterization studies have been carried out with three stations in the Las Lagunas wetland and two stations in the Mejita wetland being sampled. The results indicate generally higher water quality and nutrient parameters and generally lower metal parameters in the Las Lagunas wetland. Ammonia and cyanide concentrations decrease from the upper to lower ends of both wetlands. The benthic invertebrate community at Las Lagunas wetland appeared to be healthier than the Mejita wetland.

The results of a socio-economic baseline study show poverty and low levels of literacy in the towns and local communities around the mine site, together with significant unemployment. Potable water, energy and sewage systems are non-existent. Elementary and high school education is available in local towns as well as basic medical facilities. The studies found that communities are concerned about the reopening of the mine but realize the environmental and social benefits. The study identified the communities most concerned about mining activities and provided a means to address their concerns through a community relations program.


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17.6.6 Environmental Issues for Mine Operation

17.6.6.1 Water Management and Treatment

The following guidelines are used to develop the water management designs for the mine project:

•		International Cyanide Code

•		Dominican Republic Water Quality Standards

•		IFC Water Quality Guidelines

•		Barrick Water Conservation Standard (2007)

•		Barrick Principles for Tailings Management

Mine development is designed to treat the majority of surface water that has been impacted by historical mining activity, and to control water quality during mine operation and post closure so that the water released to the receiving environment will meet water quality standards established by the Dominican Republic government and the World Bank. The treated water will be discharged to the Margajita River. The compliance point for water quality monitoring is the confluence of the Margajita River and Hatillo Reservoir.

The main objectives of the water management plan are to avoid discharge of contact or tailings water during operations, treat and release excess contact water and tailings water, minimize or eliminate water treatment under closure conditions, provide an adequate quantity and quality of process water to the mill and achieve pit slope dewatering objectives.

The mine site surface water catchment areas and proposed water management plan are shown on Figure 17.14. The catchment areas associated with the Mejita tailings pond, the Las Lagunas tailings pond, the Hondo 2 Pit and Cumba Pit are part of the historical mining area and are not included in the mine development area for which PVDC has responsibility. PVDC has presented a proposal to the Dominican Republic government for collection and treatment of contaminated water in the areas for which the government has responsibility. This plan includes treatment for a period of 20 years.

PVDC is also responsible for water management associated with the El Llagal tailings and waste rock storage facility to the south of the existing mine area shown on Figure 17.15.

PVDC intend to meet compliance standards for water release from new mine development upon commencement of operations and within five years of start of construction for previously disturbed areas.

Monitoring will be undertaken at the site and the regional receiving environment during mine operations and into the post-closure period.


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Within the PVDC development area two dams are to be constructed to collect and store ARD contaminated water prior to treatment. Contaminated water from the proposed mining areas, Monte Negro Pit and Moore Pit, will be captured at Dam 1 located in the headwaters of Arroyo Margajita. ARD runoff from the low grade ore stockpile area will be captured at Dam 3 adjacent to Moore Pit in the upper Mejita drainage.

Water levels behind Dam 1 and Dam 3 will be maintained at the lowest possible level at all times to provide sufficient storage for the calculated 200 year return period storm event. At Dam 1 storage capacity will not be sufficient for the 200 year design storm event until year 7. The pond behind Dam 1 is designed with a geomembrane liner and under drains to limit seepage. Both dams will be constructed with spillways designed to pass the probable maximum flood resulting from the 24-hour Probable Maximum Precipitation. As of December 2007, foundation investigations were still to be completed at the site of Dam 1 and Dam 3 to advance the designs to a feasibility level.

Storage and pumping requirements for the ponds at Dam 1 and Dam 3 have been evaluated for return periods up to the design event of 200 years.

Limestone and lime requirements for the water treatment plant have been determined based on the results of test work at the HDS pilot plant. The pH discharge criteria used for the test was 8.5 to 9.0, which meets the Dominican Republic Standards for Mining Effluents and Receiving Water Quality applicable to mining effluents discharged to surface water (pH 6.0 to 9.0) but is slightly high for drinking water (pH 6.5 to 8.5).

17.6.6.2 Cyanide Treatment

Cyanide in the tailings stream will be routed to a partial cyanide-detoxification process to destroy most of the cyanide. The product will be blended with mill neutralization sludge prior to pumping to the tailings storage facility. Further cyanide degradation is expected to occur in the tailings impoundment to a level that will meet discharge criteria (see Section 17.2.4.14. The treatment process in the detoxification plant can be adjusted if necessary to reduce levels of cyanide.

17.6.6.3 Tailings and Waste Rock Storage

All tailings and waste rock from mine development will be deposited in the El Llagal valley, a tributary of the Rio Maguaca.

The proposed tailings and waste rock storage facility will be located in the El Llagal valley 3.5 km south of the plant site and will be constructed to store tailings from the CIL circuit blended with sludge from the neutralization circuit and also waste rock from the open pits. The impoundment is designed as two cells contained by cross valley dams. Storage of tailings and waste rock under a permanent water cover will prevent the onset of ARD. Conceptual designs for the 138 m high Lower LLagal (LL) dam and the 196 m high Upper LLagal (UL-2) dam have been developed. Foundation investigations have been undertaken to characterize the foundation conditions at the LL dam and within the impoundment. Further site investigation work is underway. The proposed rock fill dams will be constructed with a compacted saprolite core to provide an impermeable barrier to seepage and appropriate filter zones will be provided. Rock that is not susceptible to ARD generation will


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be quarried from within the lease to provide suitable material for construction of the downstream rockfill shell.

Design criteria for static and seismic stability meet the minimum safety factors for the high to very high consequence of failure classification as recommended by the Canadian Dam Association, Dam Safety Guidelines. Flood storage and spillway design have been developed based on extreme precipitation events.

Construction of a starter dam at the LL dam site will provide storage for the first 1.5 years of production. Annual dam raises will be designed and constructed to provide storage for subsequent years.

A tailings pipeline from the plant to the tailings impoundment and a return tailings pond decant water pipeline will be installed. The pipelines will be provided with secondary containment where they cross the river to minimize environmental damage in the unlikely event of a rupture at this location. Excess runoff from the tailings storage facility will be treated and released to the Arroyo Margajita.

Stabilization upgrade plans have been developed for the Mejita tailings dam. This facility is not within the PVDC mine development area and is the responsibility of the Dominican Republic government.

17.6.6.4 Low grade stockpile

A temporary low grade ore stockpile is to be constructed on flat terrain north of Moore Pit and northeast of Monte Negro Pit. The 127 Mt stockpile will provide for 10 years of processing after both pits have been mined. The feasibility study report does not provide details regarding ARD collection or mitigation. Discussions with Barrick personnel on site indicate that a rain blanket is one possibility under consideration to minimize the availability of water for ARD generation.

17.6.7 Mine Closure and Post Closure Impacts

The intent is to leave the site at closure with better water quality in the Margajita drainage system downstream than existed when the project commenced.

ARD contaminated water will eventually be collected in the mined out pits. The non-submerged areas of the pit walls may produce ARD requiring collection and possible long-term treatment.

Freshwater diversions, ARD collection ditches, ARD collection ponds and ARD pump stations will be required to remain in service during the post closure phase. These facilities will have to be maintained in good operating condition until water quality meets acceptable discharge criteria.

There is a potential to submerge waste rock, tailings and/or sludge in the pits after completion of mining.

Seepage from the El Llagal tailings storage facility will be required to be collected and pumped back to the impoundment until such time as the seepage meets acceptable


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standards for release to the environment. The water level in the tailings pond will be allowed to increase and the water will be allowed to flow over the emergency spillways once the water quality meets the discharge criteria.

17.6.8 Reclamation and Bond

17.6.8.1 Reclamation Plans

A conceptual biodiversity program has been developed for the ESIA and PVDC has made a commitment to work with others to identify potential biodiversity projects in the region and to implement studies and programs that meet the program objectives. The aim is to maintain the biodiversity resources and possibly enhance them with funding included in the operating costs.

The biodiversity program will include a forest habitat development program for site reclamation. Greenhouse facilities will be provided and the reclamation program will include open field trial plots.

PVDC plans to progressively reclaim the mine site as sections of the site become available.

17.6.8.2 Bond

The Environmental Licence No. 0101-06, signed 26 December, 2006, details the bonding requirements for the mining project. The compliance bond is RD$635,250,000 corresponding to 10% of the cost of the Environmental Adjustment and Management Plan (PMAA) of the construction phase. Once the construction phase is completed PVDC will provide a bond that corresponds to 10% of the amount of the updated PMAA defined for the operational phase. At the end of the operational phase PVDC will provide the corresponding bond at 10% of the total amount of the PMAA for the closure and post closure phases.

As part of the SLA agreement PVDC is required to create an Environmental Reserve Fund in an offshore escrow account funded at a rate equal to 5% of all operational costs, other than costs of concurrent rehabilitation, until the funds are adequate to discharge the closure reclamation obligations.

17.6.9 Risks and Liabilities

Ground disturbance from past mining practices has resulted in uncontrolled release of acid rock drainage and elevated metals originating from waste rock dumps, ore stockpiles and open pit rock walls. Two acid drainage treatment plants were constructed to treat the contaminated water; however, as of December 2007 only one treatment plant was operational. PVDC plans for mine development are designed to remediate or mitigate the majority of the existing problems within the project development area and PVDC have also proposed mitigation plans to improve environmental conditions outside the project area that are the responsibility of the Dominican Republic government. The collection and treatment of contaminated surface water should effectively improve water quality and reduce both short term and long-term environmental risks.


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Based on the information presented in the feasibility study it appears that the extent of ARD groundwater contamination and the potential for treatment, if necessary, is not well defined. The feasibility study alludes to the effectiveness of the limestone on the western side of the property to control the pH in the groundwater but also refers to low pH and elevated metal concentrations on the eastern side where limestone is not as prevalent. If long-term treatment of ARD groundwater contamination is deemed necessary the development of an effective treatment plan may be difficult due to the Karst nature of the geology at Pueblo Viejo.

The proposed El Llagal tailings and waste rock storage facility has been designed to store potentially acid generating materials including waste rock, tailings and water treatment sludge in a permanently flooded condition to minimize the potential for development of ARD. Two main earth embankment dams have been designed to meet Canadian dam safety standards. During mine operation water levels in the impoundment can be monitored and controlled to minimize the risk of ARD and to ensure dam stability. Quality assurance and quality control testing will be required to ensure the dams are built in accordance with the design and in accordance with the construction specifications. Dam instrumentation and monitoring programs will have to be developed to ensure that the dams perform as designed. Environmental risks should be low if good construction, operation and monitoring practices are followed. A suitable long-term dam monitoring and maintenance program will be required post closure as part of risk management procedures for the mine.

There is a risk that some of the potentially acid generating waste rock deposited in the tailings and waste rock storage facility would be exposed above water in the event of a premature mine closure.

The low grade ore stockpile presents a significant risk if the mine closes prematurely and the ore is exposed to atmospheric conditions. Collection and treatment of ARD runoff would require treatment for decades and this would incur significant closure costs.

Several sources of ARD have been identified that could impact long-term collection and treatment requirements after mine closure. The potential areas of ARD generation include the open pits, historical waste rock dumps, areas exposed from stripping, low grade ore stockpiles, seepage from the tailings impoundment, and other areas on the mine site where rock has been deposited. PVDC have recognised that treatment strategies for the long-term post closure period will have to be developed.

Storage capacity for ARD contaminated water behind Dam 1 will not be sufficient for the 200 year design event until year 7. There is therefore a risk of ARD release over the spillway during the first six years of mine life.



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17.7 Taxes

A Special Lease Agreement (SLA) between the Dominican State and PVDC is in place to regulate the development and operation of the Pueblo Viejo Mine. This agreement was approved on July 29, 2003, after congressional approval of an amendment to Article 19 of the Mining Law.

17.7.1 Taxes and Payments

Under the Agreement, PVDC is obligated to make the following payments:

•		$1,547,000 closing cost payment, upon approval of the SLA — completed

•		$500,000 payment at Project Notice — completed

•		A Net Smelter Return Royalty of 3.2%

A Net Profits Interest, equal to an Applicable Percentage of the Free Cash Flow. The Applicable Percentage is the sum of 5% plus the average of the London Gold Price, as defined in Section 8.2(b)(ii)(B)(IV) of the SLA, for the relevant fiscal year minus $275 (the difference cannot be less than zero), divided by 10%. In no event, however, shall the Applicable Percentage exceed 25%.

An Environmental Reserve Fund held in an offshore escrow account funded at a rate equal to 5% of all operational costs other than costs of concurrent rehabilitation, until the escrowed funds are adequate to discharge PVDC’s closure reclamation obligations.

The following general tax obligations apply:

•		25% Income Tax

•		5% W/H tax of loan interest

•		25% W/H tax for technical services

17.7.2 Special Lease Agreement Summary

(See Section 3.2)

17.8 Capital Costs

17.8.1 Basis of Estimate

The capital cost estimate considers all engineering, procurement, and construction costs for the mine development, process facilities, and support infrastructure. It includes all initial capital costs to bring the Project into production and also covers expenditures from the start of detail engineering to the point of loading ore into the crusher.



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The capital cost estimate for a 24,000 t/d plant is broken down into two separate estimates:

•		18,000 t/d Expandable: Includes a plant capable of processing 18,000 t/d and the facilities needed for the initial 18,000 t/d that are already sized for 24,000 t/d plant capacity.

•		24,000 t/d Expansion: Includes the plant expansion from 18,000 t/d to 24,000 t/d.

17.8.2 Capital Cost Summary

The total project cost is summarized in Tables 17.27 and 17.28.

Table 17.27 Project Capital Cost Estimate (as at Q3 2007)


Capital Costs	$ M
Mine		153.7
Process		951.2
Infrastructure		259.7
Indirects		929.1
Contingencies		291.7

Total Capital	$	2,585.4


*		No escalation included in costs



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Table 17.28 Project Capital Cost — 18kt/d and 24kt/d — by Responsibility

Owner/SNC-

18,000 t/d

24,000 t/d

Lavalin

Expandable

Expansion

(18 k&24 kt/d)

Total $

Fluor

1,325,799,818

66,200,016

1,391,999,834

Hatch

593,310,969

161,720,684

755,031,653

Owner — Process Plant

402,864,616

Subtotal

1,919,110,787

227,920,700

402,864,616

2,549,896,103

SNC-Lavalin — Transmission Line

42,817,428

Owner — Transmission Line

37,712,774

Subtotal

80,530,202

Total Project ($)

2,630,426,305

Operational Spares 2 Years (Included in above Capital Cost)

Fluor

7,743,546

483,165

8,226,729

Hatch

12,346,147

3,632,734

15,978,881

Owner

20,780,190

Total

20,089,711

4,115,899

20,780,190

44,985,800

Total Project — Excluding Operational Spares 2 Years

1,899,021,076

223,804,801

462,614,628

2,585,440,504

17.8.3 Estimate Base Date and Exchange Rates

The estimate base date is the 3rd quarter of 2007 and all costs are expressed in United States dollars ($) with no allowance for escalation beyond the base date.

The currency conversion rates per US dollar are as follows:


•	DOP (Dominican Republic Peso)	33 = US$1.00
•	EUR (European Euro)	0.71 = US$1.00
•	MXN (Mexican Peso)	11 = US$1.00
•	CAD (Canadian Dollar)	1.05 = US$1.00

17.8.4 Contingency

A contingency of 12.8% is included in the total project capital cost of $2.585M.

17.8.5 Exclusions

The following items are specifically excluded from this capital cost estimate:

•		Escalation beyond the third quarter of 2007

•		Removal of existing contaminated materials (Government responsibility)



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•		Asbestos removal not associated with demolition (i.e. asbestos roof tile disposal to landfill site is included in the demolition cost)

•		Unless existing facilities are in the area of the new plant layout, no provisions are included for additional scope of demolition or remediation

•		Salvage value from demolition of equipment and materials Taxes (ITBIS)

•		BZS-SX/EW zinc recovery facilities

•		Import duties

•		Sunk costs, including studies

•		Project interest, and financing cost during construction

•		Cost due to currency fluctuation

•		Provision for force majeure or unusually severe weather conditions

•		Government water and remediation

•		Heliport

17.8.6 Sustaining Capital Costs

During the mine life, additional capital expenditures will be required. These post-construction expenditures are summarized in Table 17.29.

Table 17.29 Sustaining Capital Cost Summary*


Sustaining Capital Mine Life	M$ US
Mining Equipment		144
Service Mobile Equipment		32
Plant & Shop Equipment		—
Mill & Maintenance Projects		79
Tailings Dam Lifts		298
Freight		11
Warehouse Inventory		5

Total	$	569


		No escalation included in costs

*		Q3 — 2007 Dollars

17.8.7 AMC Comment

Barrick completed the capital cost estimate with input from the following major consulting firms and their subcontractors:

•		Fluor (general plant site and infrastructure)

•		Hatch (pressure oxidation and oxygen plant)

•		SNC-Lavalin (power and transmission lines)



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•		NCL (mining studies)

•		BGC Engineering (geotechnical)

Personnel from PVDC and Barrick Gold provided inputs to the area of Owner’s costs.

Fluor had the responsibility of compiling the estimates from the various parties to produce the overall project estimate.

The above parties are all respected organizations with a great deal of specialist experience in major project estimation. The project estimates have been completed in a generally very detailed manner. Although not an expert in the infrastructure field, AMC is satisfied that these estimates are a reasonable, feasibility level reflection of project costs that would be incurred if the project were executed in accordance with the construction plan, schedule, and implementation plan as described in the Barrick FS.

The total capital cost estimate is M$2,600, to be spent over a 5-year period from 2008 through 2012. The project schedule is ambitious, with the motivation to commence mining and processing as soon as possible. AMC is satisfied that the schedule is realistic but issues such as permitting and power could certainly affect the rates of project execution and capital expenditure.

Provision of the operating power supply is an area of significant uncertainty with the associated potential for increased capital requirements. The capital estimate provides for construction of a 111km transmission line from Palenque to the mine site but does not include any capital cost associated with establishment of power generation capability. At the time of the Barrick FS, power generation was assumed to be provided from a new, coal-fired power plant to be built near Palenque by a 3rd party. Since then, AMC understands that other power supply options have been, and continue to be, considered.

17.9 Operating Costs

17.9.1 Operating Cost Summary

The average site and post-site operating costs from the Barrick FS are as follows:

•		Direct Site Costs: $33.68/t milled

•		Royalties, NPI, Refinery/Marketing: $6.56/t milled

•		Copper and Silver Credit: $8.28/t milled

•		Total Operating Cost: $31.96/t milled

Processing cost makes up about 75% of the total operating cost with electrical power costs being approximately 44% of processing. Power is estimated to be purchased at a cost of $0.08/kWh, except for the first year when $0.10 /kWh is projected.



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17.9.2 Operating Cost Areas

Annual operating costs will vary during the mine life. Major operating costs areas are:

•		Mine

•		Process

•		Environment

•		Site Services

•		General and Administration (G & A)

17.9.2.1 Mining Operating Costs

Mine operating costs were estimated on a monthly basis for Years 2010, 2011 and 2012, and on an annual basis for other years. The estimation considers required equipment operating hours, unit rates applied to the different equipment types, personnel needs, and unit costs for materials, services, and labour.

•		Fuel costs have been estimated at $0.63/L.

•		Operations supplies and consumables were estimated separately for each equipment type.

•		In the operations area, contractors have been considered only for blasting operation.

•		Areas included as Overheads are assays, office supplies, survey, ore control, light vehicles, etc.

•		Labour costs are estimated from required personnel numbers and by applying an annual cost per person.

•		The open pit operations can be divided into three main work areas:

•		Open pit mining ore and waste

•		Quarries mining limestone

•		Road maintenance

The open pit operation is responsible for movement of ore to the primary crusher, removal of waste to the tailings pond, and maintenance of haul roads. Also included in the open pit cost is the cost of stockpiling and reclaiming high-grade ore (for sulphur blending) and reclaiming low-grade ore in the latter years of the mine life.

The limestone from the quarries is used in the metallurgical process to produce lime and limestone slurry, and also as construction material to raise the tailings dams during the mine life. The costs of mining the limestone are distributed to the processing costs and to the cost of raising the tailings dams.

17.9.2.2 Process Operating Costs

Included in the process costs are:



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•		Process operation overheads

•		Operations labour

•		Operations staff

•		Operations hourly labour

•		General Operations supplies and other

•		Reagents and consumables

•		Fuel

•		Power

•		Maintenance labour

•		Maintenance staff

•		Hourly labour

•		Supplies and services

•		Contractors

The process plant operations and maintenance labour are distributed. Therefore, the costs for each area include labour (operations and maintenance), maintenance parts and services, power, reagents and consumables, and other costs.

17.9.2.3 Environmental Operating Costs

These costs include personnel for activities such as monitoring air emissions and water analysis. Licence fees and permits as well as various studies are also included in this cost area.

17.9.2.4 Site Services Operating Costs

Areas covered under Services are:

•		Yard, fences and roads

•		Waste management

•		Domestic water treatment

•		Site Power supply and distribution

•		Buildings maintenance (and minor construction)

•		Site water supply

•		Camp and accommodations

•		Site services mobile equipment

•		Service personnel and contractors



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17.9.2.5 G & A Operating Costs

•		General and Administration areas are:

•		Cost of operating the Santo Domingo and satellite offices

•		Senior management

•		General administration

•		IT and systems

•		Warehouse operations

•		Safety and training

•		Emergency response team (ERT)

•		HR Department operating costs

•		Community/Government relations and communications

•		Security and security systems

17.9.3 AMC Comment

Operating cost estimates have been largely generated from first principles and in a very detailed fashion. AMC accepts that these estimates are done to industry standard and represent a reasonable projection of expenditures for operation of the Pueblo Viejo site as envisaged in the Barrick FS.

Processing costs are the major component of total operating cost at about 75%. In the feasibility study, power is about 44% of the processing cost and 36% of the total. Since the generation of the Barrick FS estimates, the price of oil has risen quite significantly. This obviously affects equipment fuel costs, but could also have a major effect on the price of power should heavy fuel oil be the energy source for the eventual generation facility or facilities.

Table 17.30 is a summary of average operating costs for distinct phases of the project: the first three individual mining years, the remainder of mining years (Y04-15) and finally, the solely processing years (Y16-Y22 and Y23-25). AMC notes increased vulnerability to lower metal prices in the final 10 years of the project when mining is completed and the remaining ore to be processed comes entirely from the low grade stockpile. Unit operating cost declines from around $300/tonne to about $240/tonne, but cash cost/oz increases from around $370 to the range of $450 to $550.



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Table 17.30 Operating Cost Summary

Operating Cost Summary (US$)


	Y01		Y02		Y03		Y04-Y15		Y16-22		Y23-25		Y01-25
Average Operating Cost/t		207		218		257		305		249		231		271
Cost/t Milled		48		34		29		35		28		26		32
Cash Cost /oz		266		194		229		368		450		546		396

17.10 Economic Analysis

17.10.1 Revenue

The Pueblo Viejo Project will produce gold and silver as doré bullion, and copper in concentrate. With Base Case assumptions of gold at $700/oz, silver at $13.00/oz, and copper at $2.75/lb for 2011, $2.50 for 2012/2013, and $2.00 thereafter, the revenues, after transport and refining of ore, are projected to total $14,655 million over the life of the mine.

Table 17.31 Total Projected Revenues by Product


Product	$ M
Gold		12,896.4
Silver		1,093.7
Copper		674.9
Total	$	14,665.0

17.10.2 Capital Expenditure

Figure 17.16 shows the flow of project capital for the 18,000 t/d process plant installation and through the phased expansion to a 24,000 t/d operating capacity. Amounts shown are incurred costs (value of material and equipment manufactured, and value of work installed or performed). Total project capital is projected at M$2,585 (US dollars).



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Figure 17.17 Project Net Cash Flow

17.10.4 Sensitivity Analysis

A Sensitivity Analysis in the Barrick FS shows the project to be very sensitive to metal prices and operating costs, as summarized in Tables 17.32, 17.33, 17.34 below for what are termed the Base, Downside, and Optimistic cases.

The metal prices used for the Downside Case are:

•		Gold: $600/oz

•		Silver: $10.50/oz

•		Copper: $2.25/lb for Yr 2011, $2.00 for Yrs 2012 and 2013, and $1.50 thereafter

•		Crude Oil: $70/barrel (equivalent to $0.60/L diesel: HFO at $0.34/L)

The metal prices used for the Optimistic Case are:

•		Gold: $800/oz

•		Silver: $15.50/oz

•		Copper: $3.25/lb for Yr 2011, $3.00 for Yrs 2012 and 2013, and $2.50 thereafter

•		Crude Oil: $85 per barrel (equivalent to $0.69/L diesel: HFO at $0.39/L)



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Table 17.32 Cash Flow Summaries Base Case


					Copper		Pre-Financing Net
	Gold Price		Silver		Price		Present Value @		Pre-Financing		Payback
	$/oz		Price $/oz		$/lb		5% $ M		NPR		*Years ()**
Base Case		700		13.00		2.00		411.4		7.41		6.7
Operating Costs +10%		700		13.00		2.00		201.7		6.23		7.4
Operating Costs -10%		700		13.00		2.00		619.8		8.50		6.1
Capital Cost +10%		700		13.00		2.00		214.7		6.15		7.8
Capital Cost -10%		700		13.00		2.00		604.9		8.89		5.8

Table 17.33 Cash Flow Summaries Downside Case


					Copper		Pre-Financing Net
	Gold Price		Silver		Price		Present Value @			Pre-Financing		Payback
	$/oz		Price $/oz		$/lb		5% $ M			NPR		*Years ()**
Base Case		600		10.50		1.50		(260.1	)		3.34		11.3
Operating Costs +10%		600		10.50		1.50		(487.5	)		1.73		14.0
Operating Costs -10%		600		10.50		1.50		(42.4	)		4.74		9.3
Capital Cost +10%		600		10.50		1.50		(473.7	)		2.28		13.5
Capital Cost -10%		600		10.50		1.50		(53.4	)		4.62		9.1

Table 17.34 Cash Flow Summaries Optimistic Case


					Copper		Pre-Financing Net
	Gold Price		Silver		Price		Present Value@		Pre-Financing		Payback
	($/oz)		Price ($/oz)		($/lb)		5% ($ M)		NPR		*Years ()**
Base Case		800		15.50		2.50		1,048.1		10.71		5.1
Operating Costs +10%		800		15.50		2.50		839.0		9.73		5.4
Operating Costs -10%		800		15.50		2.50		1,255.5		11.63		4.9
Capital Cost +10%		800		15.50		2.50		857.4		9.30		5.7
Capital Cost -10%		800		15.50		2.50		1,236.2		12.38		4.5


*		Payback is calculated on undiscounted cash flow basis.

17.10.5 Payback Period

Table 17.32 indicates a Payback Period of 6.7 years for the Base Case on an undiscounted cash flow basis. Reference to Figure 17.17 shows that, on a discounted basis, the Payback Period is approximately 10 years.

17.10.6 AMC Analysis and Comment

AMC has expanded upon the sensitivity analysis described in the Barrick FS, with particular reference to areas of possible or recognized uncertainty. Net Cash flow (NCF), Internal Rate of Return (IRR), and Net Present Value (NPV) are, as in the Barrick FS, calculated on a pre-financing basis.



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17.10.6.1 Economic Model

The Barrick economic model is similar to those employed by other major mining companies for major project assessment. AMC is satisfied with the general validity of the model.

17.10.6.2 Metal Prices

As recognized in the Barrick FS, the project is very sensitive to changes in metal price.

An increase of 20% in all metal prices, with gold at $840/oz, silver at $15.60/oz, and a long-term copper price of $2.40/lb results in a project NPV_(5%) of M$1,543 (Base Case M$411), IRR at 12.86% (Base Case 7.41%), and NCF at M$4,102 (Base Case M$1,982). All other parameters for this scenario are as per the Base Case.

A decrease of 20% in all metal prices from the Base Case, with gold at $560/oz, silver at $10.40/oz, and a long-term copper price of $1.60/lb results in a negative NPV₍₅%) at -M$750, IRR at -0.72%, and NCF at -M$136. Again, all other parameters are as per the Base Case.

With all parameters other than gold as per the Base Case, a zero NPV₍₅%), IRR at 5.0% and NCF at M$1,227 results from a gold price at $644/oz.

17.10.6.3 Gold Grade and Processing

A key strategy of the project is the mining and processing of higher grade gold in the early years. A failure to deliver this early high grade, or a processing capability less than envisaged in the Barrick FS, would obviously have a negative effect on the project economics. AMC has examined the impact of a 20% reduction in gold production in years 1 and 2. The Base Case processed ore gold grade for these years is 5.94g/t and 5.72g/t respectively, compared to the average processed gold grade for the project at 2.92g/t. A reduction in year 1 and 2 grades to 4.75g/t and 4.57g/t respectively, and with all other parameters as per the Base Case, results in the project NPV₍₅%) reduced by 40% to M$245, IRR at 6.37%, and NCF of M$1,790.

A project overall reduction of 10% in gold production for the same tonnes of ore mined (90% of Base Case gold produced each year), gives a negative NPV₍₅%) at -M$98, IRR of 4.37%, and NCF of M$1,047. A zero NPV₍₅%) is given with overall gold production per year at 92% of Base Case values.

17.10.6.4 Exchange Rates

The Barrick FS estimate base date is the third quarter of 2007 and all costs are expressed in US dollars with no allowance for escalation beyond the base date. 63.2% of costs are estimated directly in U.S. dollars, 19.8% in Dominican Republic Pesos, 2.3% in Mexican Pesos, 5.2% in Euros, 9.1% in Canadian dollars, and 0.4% in Australian dollars.

Comparison of currency exchange rates of April 2008 with those used in the Barrick FS indicates a small increase in project capital of M$14, or 0.5%. The project is probably relatively immune from exchange rate fluctuations because of the high proportion of direct U.S. dollar costs.



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17.10.6.5 Capital Cost

See Section 17.10.6.7 below for the economic implications of a 20% increase in project capital cost.

17.10.6.6 Operating Cost

Section 17.10.6.7 discusses the effect of a 20% increase in power supply cost. A 20% increase in overall operating costs, with all other parameters as per the Base Case, results in a negative NPV₍₅%) at- M$136, IRR of 4.08%, and NCF of M$902.

Also of note with reference to the mine being potential vulnerability to lower gold prices, is the reduction in operating cost/tonne but increase in cash cost/oz after mining is completed and processing is only from the stockpile. Unit operating cost declines from about $300/tonne to around $240/tonne but cash cost/oz increases from around $370 to the range of $450 to $550.

17.10.6.7 Power Cost

Significant uncertainty is associated with the supply of power to the project, from both the source and operating cost points of view.

Power Capital Cost

The Barrick FS indicates power being supplied by a third party, but with off-site capital implications to the project of M$118 for construction of a transmission line from the source to the mine site. One possible scenario to address the power supply situation, as per that outlined in the PDFS of 2005, would be for the owners to take on the additional capital cost associated with building and/or purchasing a power generating facility or facilities. In the 2005 study, total capital for power, including that for constructing a generating facility, was projected at M$329. The current plan envisages a higher processing throughput with additional power requirements. This, along with increased construction costs since 2005, would result in a major increase in total power capital over the 2005 projection. AMC has not researched or examined any power supply construction or plant purchase options but has tested the implications of a major increase in project capital requirements. A 20% increase (M$520), expended in 3 equal instalments over the years 2009 to 2011, reduces the project NPV_(5%) to M$39, with an IRR of 5.2%. The NCF reduces to M$1,593.

Power Purchase Price

Power makes up about 44% of the processing cost and about 36% of the total operating cost. The project economics assume a power purchase price of $0.10/kWh for the first year of operation and $0.08/kWh thereafter. A power cost increase of 20%, with all other parameters as per the Base Case, shows the project maintaining a positive NPV₍₅%) at M$220 with an IRR of 6.35% and NCF of M$1,590.

It is also noted that one of the options for power generation would be via a Heavy Fuel Oil (HFO) facility. The Base Case assumes an oil price of $75/barrel. Using the Barrick FS



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relationship of barrel price to HFO price, an increase of 20% in HFO price equates to an oil price of about $98/barrel.

Also see Section 17.10.6.8 below regarding further implications of an increase in fuel price.

17.10.6.8 Fuel Price

As indicated above, the Barrick FS assumes a diesel price of $0.63/L (HFO $0.36/L). The economic model shows that an increase of 20% in fuel price (with no consideration of any power cost implications) reduces NPV_(5%) to M$351, with IRR of 7.08% and NCF at M$1,865.

A combination of a 20% increase in fuel cost and a power supply system dependent on HFO has further negative implications, showing a 61% reduction in NPV₍₅%) to M$159, IRR of 5.99%, and NCF of M$1,472. Again, all other parameters in the economic model are as per the Base Case.

17.10.6.9 Increased Gold Price, Operating & Capital Costs

A 20% increase over the Base Case in each of gold price, capital cost, and operating cost, results in NPV₍₅%), IRR and NCF values of M$542, 7.66%, and M$2,437 respectively.

17.10.6.10 Summary

Table 17.35 is a summary of the main results of the AMC sensitivity analysis. As with the Barrick FS assessment, the project is seen to be very sensitive to metal prices and operating costs, but also to be vulnerable to a reduction in gold production of less than 10% with the gold price at or below the Base Case level of $700/oz. The risk of a less than anticipated gold production rate, which is probably highest in the early years of operation, has specific relevance for both delivery of grade to the mill, and to the capability of the processing plant to operate as projected. Provision of power, the means of which is still uncertain, could have decidedly negative effects for both capital and operating costs, particularly at a time of high oil prices. On the positive side, sustained metal prices around $800/oz., or greater, show high economic returns and resilience against a significant increase in both capital and operating costs; and a very positive net cash flow is the result for all considered scenarios other than where there is a 20% reduction in all metal prices or a 20% increase in total operating cost (all other parameters as per the Base Case).



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19		INTERPRETATION AND CONCLUSIONS

19.1 Introduction

Goldcorp owns 40% of the Pueblo Viejo Gold Project, the remaining 60% of which is owned by Barrick. Goldcorp has reported on its share of the Pueblo Viejo mineral resources and mineral reserves in its 2008 AIF. The mineral resources and reserves for the entire project have been defined in the Barrick FS. That study provides comprehensive details on the location, accessibility, climate, history, geology, exploration data, metallurgical testing, mine design, production plans, ore processing, infrastructure and energy requirements, equipment, manpower, environmental studies, capital and operating costs, and economic analysis for the project.

The conclusions and recommendations that follow should be read in conjunction with the body of the report.

19.2 Conclusions

AMC has reviewed the details of the Barrick FS, visited the Pueblo Viejo site, had discussions with relevant personnel from both Barrick and Goldcorp, and conducted specific analysis in various, key project areas. The following summarizes AMC’s general conclusions:

The geology of the Pueblo Viejo property is well understood. Barrick has conducted an extensive exploration program that supplements knowledge from previous mining at the site and exploration activities by several other companies.

The resource database is free from major defects and is of an acceptable quality to support a feasibility study. Any remaining deficiencies are unlikely to materially affect global resource estimates, but may impact in places on local estimates.

Most of the components of the resource estimate are of a good standard, but two give cause for concern.

	1)		The use of ID³ for grade interpolation potentially results in conditional bias, with the possibility of over-estimating high grades and under-estimating low grades. This may be a material issue because Pueblo Viejo will be mined at a higher than average grade for the early years, with lower grade material being stockpiled for treatment in the later years. AMC’s check re-estimation using ordinary kriging of gold grades for the direct mill feed component of the first two years of planned mine production resulted in an average gold grade and contained gold around 10% lower than for the ID³ estimate, confirming that the use of ID³ may be a material issue associated with the early years of production.

	2)		The classification of Measured Resources applied by Barrick, which takes no account of continuity of mineralisation between drill holes, is not logical, is



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inconsistent with the geology of the deposit and has resulted in a substantial under-statement of Measured Resources (and therefore of Proven Reserves).

4.		Mineral Resources have been reported in accordance with CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines (2000) and therefore comply with NI 43-101.

5.		The mine plan, based on the mining rate (approximately 40,000 tonnes per day) substantially exceeding the processing rate (approximately 24,000 tonnes per day), and higher gold grades being mined and processed in the earlier years, is sound. However, stockpile control of sulphur grades will be critical.

6.		All aspects of the mine design, planning and optimization process have been undertaken to normal industry standards and all areas of material risk identified. The concern about the use of ID³ for gold grade interpolation has been discussed above. There is a particular risk in years 1 and 2 when autoclave capability is still building, but high gold production is projected, necessitating a good understanding of the location and extent of high grade areas, together with very selective mining and a disciplined stockpiling process.

		Further geotechnical work is recognized by Barrick as being necessary in order to bring the geotechnical design aspects to a true feasibility level. Appropriate operating and, where necessary, sustaining capital costs have been applied. Sensitivity analyses showed the pit sizes and recovered ounces to be moderately sensitive to gold prices and insensitive to pit slopes.

		Although insufficient limestone tonnages have been delineated at this stage for total project requirements, additional potential sources should yield sufficient tonnages of appropriate quality.

7.		Proven and Probable Reserves have been developed using Measured and Indicated Resources and appropriate mine designs. Mineral Reserves have been reported in accordance with CIM Standards on Mineral Resources and Reserves, Definitions and Guidelines (2000) and therefore comply with NI 43-101.

8.		While the proposed process design for Pueblo Viejo would make it one of the most complex, large precious metals projects in operation, considerable effort has gone into process development and it represents a technically and economically viable treatment route that is not dependent on unproven technology.

		The very strong dependency of process plant throughput and operating cost on sulphur grade means that careful ore scheduling to maintain optimal sulphur levels is very important.

		There is some risk that excessive scale formation will occur at the slightly higher than typical operating temperatures. This may lead to a need to reduce the operating temperature and result in slightly lower precious metals recoveries for some ore types.



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9.		The Resettlement Action Plan (RAP), covering the physical and economic displacement of 369 households from the Project Development Area, has been well conceived and funded.

10.		There are a number of environmental issues at Pueblo Viejo that are the responsibility of the government of the Dominican Republic, including dealing with the current uncontrolled release of acid rock drainage (ARD) and elevated metals originating from waste rock dumps, ore stockpiles and open pit rock walls, The project development plans should remediate or mitigate the majority of the existing problems within the project area and also improve environmental conditions outside the project area. The collection and treatment of contaminated surface water should effectively improve water quality and reduce both short-term and long-term environmental risks. Barrick has recognised that treatment strategies for the long-term post closure period will have to be developed.

		It appears that the extent of existing ARD groundwater contamination and the potential for treatment, if necessary, is not well defined. While the responsibility of the Dominican Republic government, it could have implications for the Project in later years.

		The planned tailings and waste rock storage facility (El Llagal), designed to store potentially acid generating materials in a permanently flooded condition, should effectively minimize the potential for ARD development. Environmental risks should be low if good dam construction, operation and monitoring practices are followed. A suitable long-term dam monitoring and maintenance program will be required post-closure as part of risk management procedures. Should the mine close prematurely, there is a risk that some of the potentially acid generating waste rock deposited in the storage facility would be exposed above water, and also that the low grade ore stockpile would not be fully treated and present an ARD issue. There is no reason at this stage to anticipate such a scenario.

		There is a risk that water quality may not comply with criteria for release to the receiving environment after the 75-year post closure period.

11.		The power supply for the mine and process plant has yet to be fully resolved and is recognized as being one of the major project uncertainties, both in terms of the source of supply and its operating (and, possibly, capital) cost implications. Alternatives being considered are a combination of a new and existing heavy fuel oil generation facilities with or without supply via a third party from a new, coal-fired power plant located on the south coast of the Dominican Republic.

12.		Issues surrounding the exploitation of the Las Lagunas tailings facility by a third party remain unresolved.

13.		The capital cost estimate for the project of $2.59 billion, covering initial capital costs to bring the project into production and expenditures from the start of detailed engineering to the point of loading ore into the crusher, is reasonable. The project schedule outlined in the Barrick FS is ambitious but realistic, although issues such as permitting and power supply could affect the rates of project execution and capital expenditure.



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14.

Operating cost estimates, which have been largely generated from first principles, are reasonable. Processing costs are the major component of operating cost at about 75% of the total, with power alone representing 36%. The recent increase in the price of oil affects equipment fuel costs, but could also have a major effect on the price of power should heavy fuel oil be the energy source for the eventual generation facility or facilities. The project has increased vulnerability to lower metal prices in the final 10 years of the operation when processing only stockpile material, as cash cost/oz increases from around $370 to the range of $450 to $550.

15.

The project is very sensitive to metal prices and operating costs, and is vulnerable to a reduction in gold production of less than 10% with the gold price at or below the Base Case level of $700/oz. The risk of a less than anticipated gold production rate, which is probably highest in the early years of operation, has specific relevance for both delivery of grade to the mill, and to the capability of the processing plant to operate as projected. Provision of power, the means of which is still uncertain, could have decidedly negative effects for both capital and operating costs, particularly at a time of high oil prices. On the positive side, sustained metal prices around $800/oz., or greater, show high economic returns and resilience against a significant increase in both capital and operating costs; and a decidedly positive net cash flow for all scenarios considered by AMC other than where there is a 20% or more reduction in all metal prices or a 20% or more increase in total operating cost.



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20		RECOMMENDATIONS

The following is a list of AMC recommendations for the Pueblo Viejo project. Note that, other than for items ‘1.’ and ‘2.’, all recommendations are in concurrence with what AMC understands is the intention of the project operators and within the project scope of work.

	1.		More detailed investigations should be undertaken to assess the validity of AMC’s conclusions with respect to the use of ID³ rather than ordinary kriging for resource estimation. If they are shown to be valid, it may be advisable to undertake infill drilling in selected parts of the Moore and Monte Negro deposits so that the issue of the gold grade interpolation method can be more thoroughly evaluated.

	2.		The approach to classifying Measured Resource should be reviewed.

	3.		Further assessment should be undertaken of any available data on existing sulphide ore stockpiles, particularly with respect to original sulphur grade and time of deposition. Additional drilling of stockpiles may be required to investigate and corroborate assumptions about the stockpile sulphur decay rate.

	4.		Further geotechnical work should be undertaken to bring the geotechnical pit design aspects to a true feasibility level.

	5.		Additional drilling for both supply and quality purposes should be undertaken at the appropriate stage of project development in existing and potential limestone source areas.

	6.		Water-treatment strategies should be developed for the longer-term post closure period.

	7.		A long-term dam monitoring and maintenance program should be developed.


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21		REFERENCES

The information provided and reviewed in this report is substantially based on the following sources:

•		Barrick Gold Corporation, Pueblo Viejo Project Feasibility Study Update, Volumes 1 to 6, December 2007

•		Placer Dome Technical Services, Pueblo Viejo Feasibility Study, Volume 3, Metallurgy, July 2005.

•		Placer Dome Inc. filing “Pueblo Viejo Project, Province of Sanchez Ramirez, Dominican Republic 43-101 Technical Report and Qualified Person’s Review”, AMEC, 26 October, 2005.

•		Pueblo Viejo Dominicana Corporation, Pueblo Viejo Dominican Republic, Monthly Progress Report, Period: Month Ending December 31, 2007

•		Placer Dome Inc., Pueblo Viejo Project, Feasibility Pit Slope Design, Geotechnical Investigations and Slope Design Recommendations for the Proposed Open Pit, Prepared by Piteau Associates Engineering Ltd., December 2004.

•		Barrick internal report: Pueblo Viejo Project — Slope Design Assessment for Mine Plan FS2007 rev00 (Final Report), September 19, 2007

•		Interviews on site with PVDC personnel

•		Interviews with Goldcorp and Barrick personnel

•		Site tour of Montenegro and Moore open pits including drill hole sites, Mejita and Las Lagunas tailings storage facilities, Mejita and Las Lagunas water treatment facilities, UM-1 dam, plantsite and proposed contaminated water storage areas at Dam 1 and Dam 3.

•		Site tour of location of the proposed El Llagal tailings and waste rock storage facility.

•		Examination of Pueblo Viejo geological logs, plans and sections, and inspection of cores

•		E-mails from Goldcorp and Barrick personnel


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23		QUALIFIED PERSON’S CERTIFICATES

CERTIFICATE OF QUALIFIED PERSON
Pueblo Viejo Gold Project, Dominican Republic,
Technical Report, Goldcorp Inc. effective date May 01, 2008
(the “Technical Report”)

Herbert A. Smith, P.Eng.

I, Herbert A. Smith, P.Eng., do hereby certify that:

1.		I am a Principal Mining Engineer of AMC Mining Consultants (Canada) Ltd. located at Suite 1040, 609 Granville Street, PO Box 10327, Pacific Centre, Vancouver, BC, V7Y 1G5, Canada.

2.		I graduated with a degree of B.Sc. in Mining Engineering in 1972 and a degree of M.Sc. in Rock Mechanics and Excavation Engineering in 1983, both from the University of Newcastle Upon Tyne, England.

3.		I am a member of the Association of Professional Engineers, Geologists and Geophysicists of Alberta, Professional Engineers Ontario, and the Canadian Institute of Mining, Metallurgy and Petroleum.

4.		I have worked as a Mining Engineer for a total of 30 years since my B.Sc. graduation from university.

5.		My relevant work experience for the purpose of the Technical Report is:

	•		3 years as Chief/Head Mining Engineer of an operating mine for Xstrata Nickel Ltd. (previously Falconbridge Ltd.)

	•		3 years as Head of the Life of Mine Group for 3 Xstrata Nickel Ltd. (previously Falconbridge Ltd.) operating mines

	•		9 years as a Senior Mining Engineer for operating mines for Falconbridge Ltd

	•		5 years as Chief Mining Engineer R&D projects for operating and developing mines for Falconbridge Ltd.

6.		I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“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 fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

7.		I am responsible for the preparation of Section 15, Section 17 other than Processing and Environmental aspects, and, jointly with Mr. Patrick R. Stephenson, Sections 18 to 21 and the Executive Summary, of the Technical Report.

8.		I visited the Pueblo Viejo Project on March 18, 2008 for 2 days.

9.		I have not had any prior involvement with the Pueblo Viejo Project.

10.		I am independent of Goldcorp Inc. as described in Section 1.4 of NI 43-101.

11.		I have read NI 43-101 and the Technical Report has been prepared in compliance therewith.

12.		As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this13th day of May, 2008.


Herbert A. Smith, P.Eng.


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CERTIFICATE OF QUALIFIED PERSON
Pueblo Viejo Gold Project, Dominican Republic,
Technical Report, Goldcorp Inc. effective date May 01, 2008
(the “Technical Report”)

Patrick R. Stephenson, BSc (Hons), FAusIMM (CP), MCIM, FAIG

I, Patrick R. Stephenson, BSc (Hons), FAusIMM (CP), MCIM, FAIG, do hereby certify that:

1.		I am Principal Geologist, Regional Manager and Director of AMC Mining Consultants (Canada) Ltd., Suite 1040, 609 Granville Street, PO Box 10327, Pacific Centre, Vancouver, BC, V7Y 1G5, Canada.

2.		I graduated with a BSc (Hons) in Geology from the University of Aberdeen, Scotland in 1971.

3.		I am a Fellow of The Australasian Institute of Mining and Metallurgy (Chartered Professional), a Member of the Canadian Institute of Mining, Metallurgy and Petroleum, and a Fellow of the Australian Institute of Geoscientists.

4.		I have worked as a geologist for a total of 36 years since my graduation from university.

5.		My relevant work experience for the purpose of the Technical Report is:

	•		7 years as Principal Geologist specializing in mineral resource/reserve estimation for AMC Consultants Pty. Ltd.

	•		12 years as Principal Consulting Geologist specializing in mineral resource /reserve estimation for P R Stephenson Pty Ltd.

	•		8 years as Chief Geologist for a series of Australian mining companies.

	•		2 years as Resident Manager for exploration for Renison Gold Fields in Papua New Guinea.

6.		I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“NI 43-10 1”) 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 fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

7.		I am responsible for the preparation of Sections 1 to 14 and, jointly with Mr. Herbert A. Smith, Sections 18 to 21 and the Executive Summary, of the Technical Report.

8.		I visited the Pueblo Viejo Project on March 18, 2008 for 2 days.

9.		I have not had any prior involvement with the Pueblo Viejo Project.

10.		I am independent of Goldcorp Inc. as described in Section 1.4 of NI 43-101.

11.		I have read NI 43-101 and the Technical Report has been prepared in compliance therewith.

12.		As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 13th day of May, 2008.


Patrick R. Stephenson, BSc (Hons),
FAusIMM (CP), MCIM, FAIG


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CERTIFICATE OF QUALIFIED PERSON
Pueblo Viejo Gold Project, Dominican Republic,
Technical Report, Goldcorp Inc. effective date May 01, 2008
(the “Technical Report”)

Christopher A Carr, P.Eng.

I, Christopher A Carr, P.Eng. do hereby certify that:

1.		I am a Senior Geotechnical Engineer of Rescan Environmental Services Ltd., 101 — 770 Cormorant Street, Victoria, BC, V8W 3J3, Canada.

2.		I graduated with a B.Sc. (Hons) degree in Engineering Geology and Geotechnics from the University of Portsmouth, England in 1971.

3.		I am a member of the Association of Professional Engineers and Geoscientists of BC and member of the Canadian Dam Association.

4.		I have worked as a Geotechnical/Geological Engineer for a total of 31 years since my graduation from university.

5.		My relevant work experience for the purpose of the Technical Report is:

6.		4 years as Senior Geotechnical Engineer with Environmental and Engineering consulting companies working on mining projects in Alberta and British Columbia

7.		7 years as Manager of Geotechnical Engineering and Mines Inspector with British Columbia Ministry of Energy, Mines and Petroleum Resources

8.		9 years as Senior Geological Engineer for Syncrude Canada Limited

9.		2 years as Exploration Geologist with RST in Botswana.

10.		I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“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 fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

11.		I am responsible for the preparation of the environmental aspects of Section 17 of the Technical Report.

12.		I visited the Pueblo Viejo Project on March 18, 2008 for 2 days.

13.		I have not had any prior involvement with the Pueblo Viejo Project.

14.		I am independent of Goldcorp Inc. as described in Section 1.4 of NI 43-101.

15.		I have read NI 43-101 and the Technical Report has been prepared in compliance therewith.

16.		As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 1st day of May, 2008.




Christopher A Carr, P.Eng.



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CERTIFICATE OF QUALIFIED PERSON
Pueblo Viejo Gold Project, Dominican Republic,
Technical Report, Goldcorp Inc. effective date May 01, 2008
(the “Technical Report”)

Murray (Guy) Butcher, MAusIMM

I, Murray (Guy) Butcher, MAusIMM, do hereby certify that:

1.		I am currently employed as Group Metallurgist of Goldcorp Inc. located at Suite 3400 — 666 Burrard Street, Vancouver, British Columbia, Canada V6C 2X8.

2.		I graduated with a Bachelor Degree in Applied Science from the Queensland University of Technology in 1974.

3.		I am a Member of The Australasian Institute of Mining and Metallurgy.

4.		I have practiced my profession continuously since 1975.

5.		My relevant work experience for the purpose of the Technical Report is:


•	Group Metallurgist, Goldcorp Inc.	2006 — 2008
•	Group Metallurgist, Placer Dome Inc.	1997 — 2006
•	Various positions, Placer Dome Inc.	1985 — 1997

6.		I have read the definition of “qualified person” set out in National Instrument 43-101 Standards of Disclosure for Mineral Projects (“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 fulfil the requirements to be a “qualified person” for the purposes of NI 43-101.

7.		I am responsible for the preparation of the processing aspects of Section 17 of the Technical Report.

8.		I have not visited the Pueblo Viejo Project.

9.		Prior to working for Goldcorp Inc. or Placer Dome Inc., I have not had any prior involvement with the Pueblo Viejo Project.

10.		I am not independent of Goldcorp Inc. as described in Section 1.4 of NI 43-101.

11.		I have read NI 43-101 and the Technical Report has been prepared in compliance therewith.

12.		As of the date of this certificate, to the best of my knowledge, information and belief, the Technical Report contains all scientific and technical information that is required to be disclosed to make the Technical Report not misleading.

Dated this 05th day of May, 2008.




Murray (Guy) Butcher, MAusIMM



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Figure 8.1		Alteration and Mineralization, Moore Deposit Section 94600 N, West Flank and Vein Zone (Looking North)


Figure 8.2		Alteration and Mineralization, Monte Negro Deposit Central Zone, Section 95650N (Looking North)

•		define Inferred Resources containing 2-3 Moz gold within target areas along the edges of the planned pits

•		Intersect potentially economic mineralization within high priority exploration targets near the pits


Figure 14.3		Frequency Distribution of Gold by Drilling Campaign: All Drill Holes vs. Rosario RC


PDTS Domain		Inter-Ramp Angle (°)
1		35
2		39
3		41
4		43
5		44
6		45
7		47

•		Discount Rate: 5%

•		Power Cost: $0.08kWh

•		Income Tax Rate: 25%

•		Depreciation Rate: 39%

•		Diesel Price: $0.63/L (oil at $75/barrel)

•		HFO Price: $0.36/L


				NPV_(5%) (M$)		IRR %		NCF (M$)
Base Case					411		7.41		1,982
Metal prices	All	+20	%		1,543		12.86		4,102
	All	-20	%		-750		-0.72		-136
	Gold	-8	%		0		5.00		1,227
Gold Production	Y1 & Y2:	-20	%		245		6.37		1,790
	Project	-10	%		-98		4.37		1,047
	Project	-8	%		4		5.03		1,234
Capital		+20	%		-39		5.20		1,593
Operating Cost		+20	%		-136		4.08		9
Power Cost		+20	%		220		6.35		1,590
Fuel Cost		+20	%		351		7.08		1,865
Power + Fuel		+20	%		159		5.99		1,472
Metal prices, capital & operating costs		+20	%		542		7.66		2,437


Signed by:
Herbert A. Smith, P.Eng.
Principal Mining Engineer, AMC Mining Consultants (Canada) Ltd
On 13 May 2008
Effective date of report 1 May 2008

Signed by:
Patrick R. Stephenson
Principal Geologist, Regional Manager & Director, AMC Mining Consultants (Canada) Ltd
On 13 May 2008
Effective date of report 1 May 2008

Signed by:
Christopher A. Carr, P.Eng.
Senior Geotechnical Engineer, Rescan Environmental Services Ltd.
On 1 May 2008
Effective date of report 1 May 2008

Signed by:
Murray (Guy) Butcher
Goldcorp Inc.
On 5 May 2008
Effective date of report 1 May 2008