EX-99.1 2 o40716exv99w1.htm PUEBLO VIEJO GOLD TECHNICAL REPORT Pueblo Viejo Gold Technical Report
Exhibit 99.1
(AMC CONSULTANTS)

 


 

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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Pueblo Viejo Technical Report
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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GOLDCORP INC
Pueblo Viejo Technical Report
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 230oC 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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Pueblo Viejo Technical Report
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  
     
Pueblo Viejo Gold Project technical Report — Goldcorp Inc.   xi

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
                     
    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  
     
Pueblo Viejo Gold Project technical Report — Goldcorp Inc.   xii

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
                     
 
      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  
     
Pueblo Viejo Gold Project technical Report — Goldcorp Inc.   xiii

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
                     
 
      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  
     
Pueblo Viejo Gold Project technical Report — Goldcorp Inc.   xiv

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
                     
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  
     
Pueblo Viejo Gold Project technical Report — Goldcorp Inc.   xv

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
             
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  
     
Pueblo Viejo Gold Project technical Report — Goldcorp Inc.   xvi

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
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  
     
Pueblo Viejo Gold Project technical Report — Goldcorp Inc.   xvii

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
             
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
     
Pueblo Viejo Gold Project technical Report — Goldcorp Inc.   xviii

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
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:
  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 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.

 


 

GOLDCORP INC
Pueblo Viejo Technical Report
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   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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2   RELIANCE ON OTHER EXPERTS
The Qualified Persons have not relied upon the work of any Experts who are not Qualified Persons as listed in Table 1.1 in the preparation of this report.
     
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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
(GRAPHIC)
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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Figure 3.2 Montenegro Fiscal Reserve
(GRAPHIC)
The Montenegro Fiscal Reserve is centred at 19°02’ N, 70°08’ W in an area of moderately hilly topography (see Figure 3.2) It covers an area of 4,880 ha and encompasses all of the areas previously included in the Pueblo Viejo II concession areas, which were owned by Rosario Dominicana until March 7, 2002, and the El Llagal area. On March 7, 2002, the
     
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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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Permit #   Permit or Authorization   Status
27
  Authorization to Store Explosives   In progress — Early Works Package.
28
  Certification of Blasters   In progress — Early Works Package.
29
  Authorization for Construction of Transmission Lines and Substations   In progress
30
  Construction Permit — Haul Road Crossings   In progress
31
  Construction Permit — Haul Road Crossings   In progress
32
  License for Operation of First Aid Facilities   In progress
33
  Demolition   Obtained
37
  Construction Permit — Road Access Connection   In progress
     
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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
(GRAPHIC)
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
(GRAPHIC)
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
(GRAPHIC)
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
(GRAPHIC)
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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Figure 8.1   Alteration and Mineralization, Moore Deposit Section 94600 N, West Flank and Vein Zone (Looking North)
(GRAPHIC)
Faults create centimetre-scale displacement of bedding, and pyrite-sphalerite veins occur along steep north-north-east trending faults and westerly-dipping thrusts. Two main north-north-east faults were mapped across the West Flank, sub-parallel with the Moore dacite porphyry contact. Displacement of veins preserves evidence for a lateral, sinistral component of movement.
     
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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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Figure 8.2   Alteration and Mineralization, Monte Negro Deposit Central Zone, Section 95650N (Looking North)
(GRAPHIC)
     
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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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within the boundaries of advanced argillic alteration. The outer boundary of advanced argillic alteration, combined with lithological and veining zones were used to generate domains for resource estimation.
     
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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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Figure 9.1 Main Results of 2006 Phase 1 Drilling Program
(GRAPHIC)
9.1.2 2006 Phase 2 Drilling Program
The main aims of the Phase 2 drilling program were to
  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
Both aims were again successfully achieved. Significant mineralization was encountered in all four areas drilled (Figure 9.2).
     
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Figure 9.2 Main Results of 2006 Phase 2 Drilling Program
(GRAPHIC)
     
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Figure 9.3 Main Results of 2007 Drilling Program
(GRAPHIC)
     
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Figure 9.4 Monte Oculto Discovery
(GRAPHIC)
9.2 AMC Opinion
AMC reviewed the descriptions of exploration procedures, visited drill hole sites, reviewed the results of exploration programs and held discussions with site geologists. AMC is satisfied that the Barrick exploration was undertaken to good industry standards and that the results have been interpreted appropriately.
     
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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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Figure 10.1 Location of all Drill Holes, Moore Deposit
(MAP)
     
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Figure 10.2 Location of all Drill Holes, Monte Negro Deposit
(MAP)
     
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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 30th 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
(PERFORMANCE GRAPH)
         
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Figure 12.2 Barrick 2006 / 07 QA/QC Results — Standards PV2, PV4, PV5
(GRAPHIC)
     
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Figure 12.3 Barrick 2006 / 07 QA/QC Results — Standards PV1, PV7
(GRAPHIC)
The actions/corrections taken are listed in the in-house system
Custom and commercial standards were used, with different grade ranges.
(GRAPHIC)
     
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Figure 12.4 Barrick 2006 / 07 QA/QC Results — Core Duplicates
(GRAPHIC)
     
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Figure 12.5 Barrick 2006 / 07 QA/QC Results — Sample Grain size
(GRAPHIC)
  Each sample is weighted
 
  Granulometric control is done on 3% of the samples at the crusher stage and 3% on the pulverization stage.
(GRAPHIC)
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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holes of known quality. These comparisons were generally satisfactory and are described in Section 13 of this Technical Report.
AMC observed Barrick’s QA/QC procedures during its site visit.
12.4 AMC Opinion
In AMC’s opinion, the QA/QC procedures and results for the Placer and Barrick drilling programs were generally satisfactory and provided sample results that may be relied upon for resource estimation purposes. See Section 13 for commentary on the reliability of pre-Placer data.
     
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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
(GRAPHIC)
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
(GRAPHIC)
(GRAPHIC)
                                                                                         
                    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
(GRAPHIC)
(GRAPHIC)
                                                                                         
                    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
(GRAPHIC)
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
(GRAPHIC)
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                   Avg                   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  
RC
                                                    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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Figure 14.3   Frequency Distribution of Gold by Drilling Campaign: All Drill Holes vs. Rosario RC
(BAR CHART)
(LINE CHART)
14.3.5 Topography
A graphics database provided by site personnel was used to create a Vulcan surface wireframe. That surface was cropped to an area larger than Placer optimal pits to limit the extent of the block model created. A copy of the topography surface was then raised by 20m and used to limit the vertical extent of the block model to ensure that at least a full row of blocks would be generated above the actual topography. Blocks were then loaded with the percentage of the block that sits below topography.
     
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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
(PERFORMANCE GRAPH)
(PERFORMANCE GRAPH)
                                                                                         
                    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
(PERFORMANCE GRAPH)
                                                                                         
                    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
(PERFORMANCE GRAPH)
(PERFORMANCE GRAPH)
(PERFORMANCE GRAPH)
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
(PERFORMANCE GRAPH)
(PERFORMANCE GRAPH)
                                                                                         
                    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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14.3.16 Variography
Down-hole and omni-directional semi-variograms, correlograms and indicator variograms were calculated using the 10m composites. All variograms were fitted with spherical models. All show very continuous mineralization, and the omni-directional correlogram tends to show ranges of 65m and 120m, which correspond to 80% and 90% of the total variance, respectively. Down-hole and omni-directional correlograms are shown in Figures 14.8 and 14.9 respectively.
Figure 14.8 Down-Hole Correlogram Gold
(PERFORMANCE GRAPH)
         
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Figure 14.9   Omni-Directional Correlogram Gold
(GRAPHIC)
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 ID2 rather than the ID3 used for gold grade interpolation.
Figure 14.10 Frequency Distribution of Sulphur Grades in 10m Composites
(GRAPHIC)
                                                                                         
                    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
(PERFORMANCE GRAPH)
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
(PERFORMANCE GRAPH)
(PERFORMANCE GRAPH)
                                                                                         
                    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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16 MINERAL PROCESSING AND METALLURGICAL TESTING
Refer to Section 17.2 of this report.
         
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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 17.1 Moore Pit from Monte Negro
(GRAPHIC)
         
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Figure 17.2 Old Processing Plant at Pueblo Viejo
(GRAPHIC)
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
(PERFORMANCE GRAPH)
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
(GRAPHIC)
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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Table 17.9 Pit Optimization Slope Angles
     
PDTS Domain   Inter-Ramp Angle (°)
1   35
2   39
3   41
4   43
5   44
6   45
7   47
17.1.5.4 Pit Optimization Results
As indicated earlier, metal prices used to determine the economic reserve were: gold: $575/oz, silver: $10.75/oz, copper: $2.00/lb. Figure 17.5 shows the resulting pits from the Whittle exercise for revenue factor one (gold price $575/oz).
Figure 17.5 Pit Optimization Illustration
(GRAPHIC)
     
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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   Au   Ag   Cu   S   Waste    
Deposit   (kt)   (kt)   (g/t)   (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
(GRAPHIC)
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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Figure 17.7 Ore to Crusher
(PERFORMANCE GRAPH)
Figure 17.8 Mine Daily Movement (excluding Quarries)
(PERFORMANCE GRAPH)
     
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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)