Pueblo Viejo is a Cretaceous high sulphidation epithermal gold, silver, copper and zinc deposit. High sulphidation deposits are typically derived from fluids enriched in magmatic volatiles, which have migrated from a deep intrusive body to an epithermal crustal setting, with only limited dilution by groundwater or interaction with host rocks. Major dilatant structures or phreatomagmatic breccia pipes provide conduits for rapid fluid ascent and so facilitate evolution of the characteristic high sulphidation fluid.

The PVDC interpretation is based on geological evidence observed within the Pueblo Viejo deposit and is not a regional interpretation as presented by Sillitoe et al. (2006). However, PVDC believes that uncertainty with respect to the deposits origin has no practical impact on exploration at the levels that may be mined by open pit methods. The areal extent of the deposits has been constrained by drilling and the vertical extents are reasonably well known, although additional drilling is required to define the deepest parts of the deposit.

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

Studies have determined that there were three stages of advanced argillic alteration associated with precious metal mineralization:

1. Stage I alteration produced alunite, silica, pyrite, and deposited gold in association with disseminated pyrite.

2. Stage II overprinted Stage I and produced pyrophyllite and an overlying silica cap.

3. Stage III of mineralization occurred when hydro-fracturing of the silica cap produced pyrite- sphalerite-enargite veins with silicified haloes. Syntaxial vein growth preserves evidence for pyrite-enargite-sphalerite-grey-silica paragenesis.

Individual Stage III veins have a mean width of four centimetres 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-northwest and can be 250 m long, 100 m wide, and 100 m vertical. Stage III veins contain the highest precious and base metal values and are more widely distributed in the upper portions of the deposits.

Veins tend to be parallel and follow a number of local structures that crosscut the deposit. Those structures have a northerly trend at Monte Negro and Moore, with a northwestsoutheast 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, tetrahedritetennantite, geocronite, galena, siderite, and tellurides are also found in veins.

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

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

The Pueblo Viejo Mine consists of two open pits: Moore and Monte Negro. Mining operations are undertaken by a conventional truck and shovel method.

Pit dimensions: 2.5km long x 1.5km wide x 300m deep (no backfill) and Typical bench height - 10m.

The process plant is designed to process approximately 24,000 tpd of ROM refractory ore. The plant bottleneck is the supply of oxygen. If the ROM feed has a low sulphide content, the plan can process 30,000 tpd. The design basis for the oxygen plant is to provide the oxygen required to oxidize approximately 80 tonnes per hour (tph) of sulphide sulphur. This is equivalent to 1,200 tph of feed containing 6.79% sulphide sulphur, assuming a design factor of 2.2 tonnes O2 per tonne sulphide sulphur. The overall facility consists of the following unit operations:
• Primary crushing
• SAG mill and ball mill grinding with pebble crushing (SABC circuit)
• POX
• Hot curing
• CCD washing
• Iron precipitation
• Copper sulphide precipitation and recovery
• Neutralization
• Solution cooling
• Lime boiling for silver enhancement
• CIL circuit
• Carbon acid washing, stripping and regeneration
• Electrowinning (EW)
• Refining
• Cyanide destruction
• ARD treatment
• Limestone crushing, calcining, and lime slaking.

The ore is ground to an optimum size of 80% passing 80 µm to120 µm and oxidized in autoclaves at a temperature of 210°C to 230°C and a pressure of 3,100 kPa to 3,450 kPa for 60 minutes to 75 minutes. The product from each autoclave is discharged to a flash vessel where heat is released, cooling the slurry to approximately 106°C. It then flows by gravity to the hot cure circuit where the slurry temperature is maintained between 100°C and 105°C for 12 hours in order to dissolve the basic ferric sulphate (BFS) that forms during the pressure oxidation process. This process overall temperature will be deliberately lowered in 2018 to approximately 96°C to protect equipment. The resulting BFS dissolution will essentially be the same.

The next step in the process is to separate the base metal rich acidic liquors from the oxidized solids within the slurry. This is accomplished in a three-stage CCD wash thickener circuit to remove more than 99% of the sulphuric acid and the dissolved metal sulphates. The washed thickened slurry is then contacted with steam from one of the autoclave flash vessels to heat the slurry to 95°C ahead of a two-stage lime boil treatment. Adding milk of lime slurry to the oxidized slurry effectively raises the pH to the 10.5 to 10.8 range breaking down the silver jarosites, making it possible to recover the silver minerals in the CIL circuit. Following the lime boil circuit, the slurry is diluted with reclaimed water and cooled to 40°C in cooling towers. The cooled slurry is pumped to the CIL circuit.

The addition of lime to the lime boil circuit provides sufficient protective alkalinity in the CIL circuit. No further addition of lime is required in this circuit. In the CIL circuit, cyanide is added to solubilize the gold and silver into solution which is contacted with activated carbon to adsorb the gold and silver cyanide complexes. Retention time in this circuit varies from 18 hours to 22 hours, depending on the processing rate.

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

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

The tailings from the CIL circuit flow by gravity across the carbon safety screens and are pumped to the cyanide destruction circuit. The conventional SO2/air process reduces the cyanide content of the CIL tailings to less than 5 mg/L cyanide. The detoxified slurry is mixed with the HDS and pumped to the TSF.