The standard orogenic gold model characterizes the majority of gold deposits within the Abitibi belt. However, several examples of late mineralization are disseminated and associated with alkaline intrusions (Robert, 2001), thus differing from the standard orogenic gold model. Syenite-associated disseminated gold deposits in the Abitibi greenstone belt consist of disseminated sulphide zones with variably developed quartz stockworks, which are intimately associated with Timiskaming-age, monzonitic to syenitic porphyry intrusions (Robert, 2001). Like quartz-carbonate vein deposits, all known syenite-associated disseminated gold deposits in the southern Abitibi belt occur along a major fault. As a result of their distribution along major faults, these deposits commonly occur at or near boundaries between contrasting lithological domains. Examples of these deposits are Young-Davidson, Matachewan Consolidated, Ross, Holt-McDermott, and Lightening in Ontario; and Beattie, Douay, Canadian Malartic, East Malartic, and Barnat Sladen in Québec.

The Timiskaming-type sedimentary rocks occur along restricted segments of a major fault zones, where they are preserved as synclinal keels (Muller et al., 1991). The syenite intrusions form small stocks, commonly elongated subparallel to the overall structural trend and generally surrounded by numerous satellite dykes (Robert, 2001). Although some intrusive phases are equigranular, most are porphyritic, with K-feldspar phenocrysts in a fine-grained to aphanitic groundmass.

Syenite-associated disseminated gold deposits consist of zones of disseminated sulphides with variably developed stockworks in intensely altered wallrocks (Robert, 2001). They have sharp diffuse limits, defined by a decrease in sulphide content, gold grades, and intensity of stockwork fracturing. Owing to the abundance of microveinlet stockworking and fracturing, many orebodies take on a breccia appearance. The morphology of the deposits ranges from overall tabular to pipe-like, although many have rather irregular outlines. Most mineralized zones are steeply dipping or steeply plunging, but examples of moderately to shallowly dipping orebodies, discordant to lithological units, are also known (Robert, 2001).

The total sulphide mineral content of the orebodies is typically less than 10% by volume, and commonly a few percent (Robert, 2001). Disseminated sulphides are fine- to very fine grained and consist dominantly of pyrite, with significant arsenopyrite in a few deposits. Associated stockworks consist of millimeter- to centimeter-thick veinlets of grey to cherty quartz, with subordinate amounts of carbonate (Fe-dolomite and calcite), albite, and pyrite. In addition to pyrite and arsenopyrite, ore related metallic minerals include minor to trace amounts of chalcopyrite and hematite. Telluride minerals, molybdenite, and magnetite are common associates of this type of mineralization, whereas galena, tennantite, and bismuthinite occur at few deposits. Accordingly, orebodies are generally enriched in Cu, As, Te, with common, but variable, enrichments in Pb, o, W, Zn, and locally Sb. The gold:silver ratios of the ores generally range from about 1:1 to 5:1.

Zones of hydrothermal alteration are spatially coincident with zones of disseminated sulphide minerals and veinlet stockworks, and most intense alteration corresponds in a general way to economic mineralization (Robert, 2001). Carbonatization and albitization are significant alteration types at nearly all deposits; K-feldspar alteration and sericitization are also present in several deposits.

According to Bevan (2011), the “main” type of gold mineralization in the Duparquet deposit generally occurs within shears or fracture zones along or within the adjacent intrusive syenitic masses, and is associated with finely disseminated pyrite and minor arsenopyrite replacement. Sulphide content is generally low (0.5 to 4%), although it can be up to 10% in cases. Higher gold grades appear to be related to the finer grained sulphides (Bevan, 2011).

Mineralization at the Central Duparquet is hosted by the CDFZ (Central Duparquet Fault Zone), and is of a similar nature a the South and North zones (Bevan, 2011).

Mining of the Duparquet deposit has been designed as an open pit with a planned production of 3,650,000 tonnes per year (3.65M tpy) or 10,000 tonnes per day (tpd).

The life-of-mine (LOM) for the Duparquet Pit was based on supplying the mill with 3,650,000 tonnes of ore per year. Initially, 4.1 Mt of tailings would be reserved to supplement the mill when necessary, but it was later decided to process the tailings at a rate of 750,000 tonnes per year right from the start in order to clean the tailings area to make room for the waste stockpile #2 for the west pit.

The LOM for the Duparquet Pit will be spread over 11 years, preceded by a 4-year pre production period.

Also, the existing pit contains approximately 5.7 Mm3 of water. At the start of the projectProject, it is estimated that 2 Mm3 need to be dewatered, and that it will take approximately six (6) months.

The mining schedule will require the extraction of 39,363,029 tonnes of ore and 291,213,881 tonnes of waste rock, resulting in a LOM strip ratio of 8.26 to 1. The overburden consists of 23,398,919 tonnes. Taking into account the overburden, the LOM strip ratio is 8.93 to 1. The stripping ratio (waste to ore) will not necessarily be constant.

The drilling pattern has been selected in order to control blast induced vibrations and airblast overpressures in the Town of Duparquet. The spacing and burden for the Project were estimated at 6.5 m and 6 m respectively. Percussion drills with 215 mm bit size will be used in ore and waste. A penetration rate of 22 m/hr was used when estimating equipment requirements. A total of four (4) drills will be required to achieve the production target.

Pre-split blasting will be done in order to maximize stable bench face and inter-ramp angles along the final walls. The pre-split consists of a row of closely-spaced holes along the final excavation limit. One percussion drill with 140 mm bit size will be used for pre-shear and holes will be drilled at an interval of 1.5 m.

An explosives supplier will provide the mine with explosives. In this study, it was assumed that the explosives will be truck to Duparquet from the supplier plant located in Malartic.

The plant is designed to operate 24 h/d, 365 d/yr, and process 3.65 Mt of mineralized material (dry) annually, at a plant availability of 92%. The nominal daily throughput will be 10,000 tonnes of dry material.

The ROM ore will be trucked by mine haulage trucks to the crushing plant, dumped into the crusher feed hopper, and conveyed by apron feeder into the primary jaw crusher. The crusher product will then be conveyed by two conveyors to the crushed ore stockpile. The live capacity of the stockpile will be 10,000 tonnes with a total capacity around 36,000 tonnes. Apron feeders will then draw off the material at 452.9 t/h from below the stockpile and conveyors will deliver the material to the SAG mill feed chute.

The SAG mill is driven by a 5,670 kW (7,200 hp) motor. The SAG mill product will be discharged to the SAG mill discharge screen; the coarse material (screen oversize) will be conveyed to the pebble crusher driven by a 315 kW motor and crushed pebbles will be recirculated to the SAG mill. The screen undersize will gravity-feed the pump box and mix with the ball mill discharge. The secondary grinding circuit consists of one 6,410 kW (8,600 hp) ball mill which will operate in closed circuit with one hydrocyclone cluster. The hydrocyclone underflow will return into the ball mill while the overflow with a P80 of 100 µm will gravity feed to the trash screen prior to transfer to the conditioner tank of the flotation circuit.

The flotation circuit consists of one (1) conditioning tank, eleven (11) rougher tank cells, six (6) 1st cleaner cells, two (2) 2nd cleaner cells, and a rougher concentrate regrind mill. Reagents will be added to the conditioning tank and to different points of the rougher and cleaner flotation stages. The rougher flotation tails will be pumped to the flotation tailings pre-leach thickener feed tank ahead of the cyanidation and CIL circuits. The rougher flotation concentrate will be pumped to the regrind cyclone feed pump box. The regrind ball mill will operate in closed circuit with one cyclone cluster. The cyclone underflow will be returned to the regrind mill while the cyclone overflow with P80 of 35 µm will be pumped to the 1st cleaner cells. The 1st cleaner tails will return to the rougher feed while the concentrate will be pumped to the 2nd cleaner cells. The second cleaner tails will be pumped back to the 1st cleaner feed, while the 2nd cleaner concentrate will be the final product of the flotation circuit with 2.47% weight recovery (247 tpd concentrate), 50 g/t Au content, and 86.4% gold recovery.

The flotation tails will be delivered to the pre-leach thickener. The thickener will raise the solid density of the slurry to approximately 60-65%. The thickener underflow will then be mixed with final tailings thickener overflow to achieve a slurry with 45 50% solid density that will feed the cyanidation tanks. Lime will be added to the thickener to raise the pH level to 10.5, which will ensure that no hydrogen cyanide evolves from the first leach tank following the addition of sodium cyanide. The thickener overflow will be returned to the process and used as process make-up water.

Lime and sodium cyanide will be added to the first leach tank and the third and fourth CIL tanks to maintain a constant extraction rate. The pulp will decant from first leach tank to the second tank and from the first CIL tank to the sixth tank over the course of 24 hours. The gold in solution in the CIL tanks will be adsorbed onto the activated carbon.

The carbon will be contained by in-tank screens which will reverse the carbon flow counter current to the incoming slurry. Fresh/reactivated carbon will be fed to the last tank in the CIL train, and a safety screen will extract any broken carbon that escapes through the in tank screens. The undersize (slurry) from the screen will be pumped to the final tailings thickener.

The gold-loaded carbon will be pumped from the first CIL tank to the loaded carbon screen. The oversize (carbon fraction) from the screen will be discharged into the loaded carbon tank and will then be pumped to the elution circuit, while the undersize will be recycled back into the cyanidation tank.

In the elution circuit, the loaded carbon will be washed with recirculating hydrochloric acid in the acid wash vessel, neutralized with caustic soda (sodium hydroxide), and then washed with process water prior to being pumped to the carbon stripping vessel. In the carbon stripping vessel the gold will be stripped from the carbon using a hot caustic cyanide solution. The solution will be maintained at a temperature of 135 °C and an internal pressure of 350 kPa. The stripping solution will be heated with solution heaters and heat exchangers, and then pumped into the stripping vessel. The gold laden solution will exit the stripping vessel via the heat exchanger and flowing into the pregnant solution storage tank.

The pregnant solution will be pumped into the electrolytic (electrowinning) cells where the gold will be plated out of solution onto the steel wool cathodes. The stripped solution will exit the electrowinning cells and will gravity feed the barren solution tank for reuse during the next elution cycle once the cyanide concentration is restored.

The gold laden steel wool will be smelted in a tilting furnace with the addition of fluxes, typically sodium borate, sodium carbonate, and silica. The impurities are slagged off and the gold will be poured into the doré ingot moulds. The slag will be recycled back to the jaw crusher feed to re-enter the process to minimize the gold losses. The gold recovery from the flotation tails CIL circuit will be 47.8% of the CIL feed or 6.5% of the plant feed.

The stripped carbon will be regenerated at 650 °C in an indirectly-fired kiln for 30 minutes in the absence of oxygen. It will then be cooled in a quench tank. The reactivated carbon will be pumped into the carbon attrition tank, where the fresh carbon will be added as make-up to replace attrited carbon. Prior to return into the CIL circuit, the reactivated carbon will be screened to remove any fine carbon which may adsorb gold. If the fines are allowed back into the CIL circuit, they would cause gold losses for the process.

The final tailings will be thickened in the final tailings thickener and detoxed in the cyanide destruction circuit (CND circuit) prior to transfer to the tailings pond. The CND will be performed at 50-55% solid density using the SO2-air method.

The POX processing option also uses crushing, grinding, and flotation, but followed by pressure oxidation and cyanidation of the oxidized flotation concentrate. For the POX option, the flotation circuit is operated to maximize gold recovery to flotation concentrate instead of maximizing the gold concentrate grade as in the Concentrate option. With much less gold being rejected in the flotation tailings, cyanidation of the tailings is currently not economically feasible in light of the CAPEX and OPEX associated with a tailings CIL circuit, as well as the price of gold used ($1,350/oz).

The POX option uses the same crushing and grinding circuits as the Concentrate option. The flotation circuit will include the rougher, regrind, and 1st cleaner circuits (2nd cleaner will not be needed). After thickening, the flotation concentrate will pass through the pressure oxidation circuit to be oxidized followed by a carbon-in-pulp (CIP) circuit for cyanidation and adsorption for gold recovery. The flotation tailings will be considered as final tailings. In the CIP circuit, gold will be cyanide leached and adsorbed on the carbon and then recovered from the carbon by an elution circuit, followed by electrowinning and doré smelting as previously described for the Concentrate option tailings CIL circuit. Carbon regeneration has also been included.