Two types of gold mineralization styles have been identified to date on the Property. The first, host to the bulk of mineralization in the Detour Lake area (Detour Lake and West Detour deposits), occurs as orogenic greenstone-hosted hydrothermal lode gold deposits.

A second style of gold mineralization was identified in the Lower Detour area, which is localized into three spatially related mineralized horizons known as the 58, 58N, and 75 Zones. These mineralized lenses share characteristics with both syenite-associated oxidized intrusion-related deposits of the Kirkland Lake area (Robert, 2001) and the Sigma-Lamaque deposits of the southern Abitibi (Robert, 1986), the latter of which are thought to be borne of the orogenic model described above. Still in the early stages of exploration and research, a clear genetic model has not yet been determined.

At the Detour Lake mine, the gold mineralization is characterized into two main zones: hangingwall mineralization and footwall mineralization.

The mineralization forms a 200 metre wide (locally up to 350 metres) corridor within a broad assemblage of mafic volcanics with an overall east-west trend. To date, Detour Gold has drill tested this mineralized corridor for a distance of over five kilometres, west from the former Campbell pit. The bulk of the mineralization within this corridor is concentrated along a highly strained corridor of a moderate to strong potassic alteration envelope at the contacts between the pillowed mafic flows and massive flows.

At the eastern end of this corridor at the Detour Lake mine, the main mineralized zone is referred to as the Main Zone (terminology of the former Detour Lake mine). The Main Zone was the largest gold-bearing mineralized zone and consists of gold mineralization occurring in the CMH or in quartz and quartz-carbonate vein systems splaying from the SLDZ. Placer’s Quartz Zones Q50, Q70, Q100, and Q120 zones are typical quartz veins arrays splaying off the SLDZ into the hangingwall (Barclay, 1993). These quartz veins are typically less than one metre in width. Gold occurs generally as free gold with these veins. The quartz veins commonly occur with a frequency of greater than one vein per metre.

West of section 19,620E, the mineralization mainly straddles the lower massive flow-pillow flow contact and is commonly associated with increased biotite alteration, shearing, narrow quartz veining and minor pyrite or pyrrhotite. Local zones of strong brecciation with sulphide infilling have also been recognized along with minor chalcopyrite, telluride minerals and visible gold. Visible gold is usually found as specks, clusters or fracture coatings. The gold mineralization is associated with a series of sub-vertical to arcuate deformation zones characterized by enhanced strained fabrics, well defined open-space breccias, and to a lesser degree sheeted shear-hosted veins and extensional veins.

From sections 16,160E to 17,160E, Placer intersected the QK Zone along a massive flowpillow flow contact between 600 metres and 750 metres below surface. This mineralization is associated with narrow parallel to sub-parallel quartz veins, quartz boudins and sulphide rich veins/breccias with adjacent silicification and potassic alteration envelopes. Gold is directly associated with sulphides.

The footwall mineralization (Talc Zone) commonly contains boudins of quartz veins and felsic as well as intermediate dyke slivers. It is generally intensely sheared, especially in the area adjacent to the contact with the volcanoclastic sediments. Gold in the Talc Zone occurs dominantly within pyrite-pyrrhotite and minor chalcopyrite along foliation planes, narrow discrete shears or strain zones and in irregular lenses (Barclay, 1993). The zones also contain short deformed lenses or boudinaged quartz veins. In some cases, it appears that the mineralization is controlled by strong fault structures containing several centimetres of gouge material.

Mining at Detour Lake is currently conducted using conventional open pit mining methods. The drill and blast, load and haul cycle uses an ‘Ultra Class’ mining fleet. Waste material is categorized then stored in various locations, typically segregated by material type (NAG, PAG, overburden, mineralized waste (low-grade) stockpiles). Ore is fed directly to the processing plant or stored in ROM stockpiles for processing at a future date.

Mining at the West Detour is planned to employ similar conventional mining methods with the initial use of smaller equipment for the pre-stripping phase, especially in overburden (starting in 2018). Once the majority of working faces have been advanced into hard rock, a mixture of Ultra-Class and smaller mining equipment will be used at West Detour.

The processing plant was designed to process ore at an average throughput of 55,000 tpd or 20 Mt per year, equivalent to milling rates of 2,500 tpoh with operating time of 92% in a 24 hour day.

Crushing and Grinding
The plant throughput was designed to average 55,000 tpd at 92% availability at a grind size of 95 µm.

The selected primary crushing system was a single stage, open circuit, primary gyratory crusher (60 x 113 inches). The crusher selection was based on a feed top size of 1,200 mm and a product P80 of 165 mm with an availability of 60% at 55,000 tpd. The live capacity of the feed and discharge hoppers of the gyratory crusher was designed for two truck loads each, assuming a nominal payload of 300 t.The crushed ore storage pile was designed with a live capacity corresponding to approximately 12 hours of crushing or 30,000 t, and an overall capacity (live plus dead) of 100,000 t.

Ore is reclaimed from the stockpile through two reclaim tunnels, one for each grinding line. Four apron feeders, two in each reclaim tunnel, discharge the crushed ore onto a belt conveyor that feeds a secondary cone crusher operated in open circuit. The secondary crusher is fed with the gyratory product with a P80 of 165 mm, and gives a product with a P80 of 50 mm. The secondary crusher product is conveyed directly to the SAG mill. The secondary crusher is equipped with a bypass chute to maintain high process plant availability. During maintenance of the secondary crusher, the bypass is put into place to feed the SAG mill directly from the stockpile.

The SAG mill operates in closed circuit with a pebble crusher while the ball mill operates in closed circuit with hydrocyclones. The total power required to grind ore to final ball mill product is fairly high at 23.3 kWh/t (25 kWh/t installed), reflecting the hardness of the ore.

Gravity
Around 34% of the cyclone feed material is sent to the gravity recovery circuit based on GRG test work completed on the metallurgical samples. From the design criteria, the calculated tonnage going to the gravity circuits of the two lines is approximately 2,100 tpoh, from which the equipment was sized. Six gravity concentrators with 48 inch bowls were selected (three per grinding line supporting a strategy of two units in operation at all time). A batch intensive cyanidation system is used to process the gravity concentrate. The extraction performance of gold from the gravity concentrate by the intensive cyanidation system is designed at 99%. The pregnant solution is pumped to a tank in the gold room followed by electrowinning in a dedicated cell.

Leach and CIP
The gold recovery circuit selected was a “leach followed by a CIP” circuit. The designed retention time for leaching is 29 hours. The leach feed size was designed at a P80 of 95 µm.

The CIP design selected uses a carrousel type system with an average retention time of 80 minutes in this circuit. The carbon handling systems were designed to handle 20 t of carbon per strip. All the carbon transfers are done using recessed impeller pumps to minimize carbon abrasion.

Acid Wash, Stripping, Electrowinning, and Refining
The stripping system uses a modified version (i.e. on-line electrowinning and no pregnant solution tank) of the high-pressure Zadra process to recover the gold from the loaded carbon.

The circuit incorporates an acid wash stage and is designed to handle up to 10 t of carbon per day.

Based on the upgrading ratio expected in the CIP circuit at the provided operating conditions, one strip every two days per CIP line is required (2 CIP lines), which corresponds to one strip per day in total. The stripping circuit is designed to handle up to 20 t of carbon per stripping cycle (approximately 8-10 hours); 100% of the carbon is regenerated in two regeneration kilns designed to handle up to 20 t of carbon per day.

The electrowinning is done “in-line” with the stripping circuit. The flow of pregnant solution is split between three rows of two electrowinning cells. One dedicated electrowinning cell handles the pregnant solution from the high intensity leach circuit (ILR tank).

The refining equipment is designed to handle the gold from the stripping circuit and from the gravity recovery system. The electrowinning sludge is filtered, dried and mixed with fluxes, before being smelted using induction furnaces.