Project Geology
In the Project area, the greenstone–granite architecture is partially aligned and disrupted along a linear, east–west-trending belt that defines the position of the Sunday Lake Deformation Zone.

Supracrustal rocks within the Project area consist of a thick sequence of mafic to ultramafic lithologies, which are predominantly volcanic in origin, and are part of the Deloro assemblage. They occur within regional synclinal-anticlinal fold structures traced for over 30 km across the Project and are in structural contact to the south with the younger sediments of the Caopatina assemblage. These rocks are bounded to the north and west by the Opatica basement gneissic rocks. To the east and south, the Deloro assemblage is intruded by several large, weakly foliated granodioritic to tonalitic intrusions which are intersected by numerous local felsic to mafic dykes, sills and younger regional Proterozoic diabase dykes.

Mineralization
There are two recognized episodes of gold mineralization at the Detour Lake and West Detour deposits. The first episode consists of a wide and generally auriferous sulphide-poor quartz vein stockwork formed in the hanging wall of the Sunday Lake Deformation Zone. The sulphidepoor quartz vein stockworks observed in the hanging wall have sub-vertical north or south dips and are parallel to a series of east-west trending high strain zones. These veins form a weak stockwork and are boudinaged and/or folded.

The second episode is a stage of gold mineralization overprinting the early auriferous stockwork, principally in the hanging wall of the Sunday Lake Deformation Zone, with a higher sulphide content. The sulphide-rich gold mineralization predominantly fills structural sites in deformed quartz veins, fractures and veins crosscutting the foliation fabric but also in pillow breccias and selvages. The distribution of sulphide-rich mineralization is strongly controlled by the geometry of kinematic orientation (i.e., pyrite and pyrrhotite concentrations have a shallow westerly plunge similar to the plunge of the main flexure zone in the Sunday Lake Deformation Zone at an angle of about 40° (in the area of the former Campbell pit), shallowing to approximately 10° further to the west).

Deposit Model
The Detour Lake and West Detour deposits are considered to be examples of orogenic greenstone-hosted hydrothermal lode gold deposits.

Greenstone-hosted hydrothermal lode gold deposits are typical of the Abitibi Greenstone Belt, and in particular the gold deposits found along the Destor–Porcupine Fault Zone from Timmins, Ontario through to Destor, Québec. These deposit types are found in greenstone belts around the world and are responsible for a large proportion of past world gold production, including most of the Canadian gold production.

The majority of Archean orogenic greenstone-hosted lode gold deposits occur within volcano–plutonic domains, which are typically distributed along crustal-scale fault zones occurring along or in close proximity to terrane or sub-province boundaries (Card et al, 1989). Elongate belts of metavolcanic and some metasedimentary rocks containing subsidiary amounts of ultramafic to felsic intrusive rocks typically dominate these domains. The intrusive rocks will have typically been emplaced in multiple pulses throughout the geologic evolution of the area. Metamorphism within the belts is generally greenschist to lower amphibolite facies. The structure of the gold districts is characterized by the presence of multiple generations of structural fabrics indicating the presence of several periods of deformation.

The deposit model for the 58, 58N, and 75 Zones is not firmly established. These mineralized lenses share characteristics with both syenite-associated oxidized intrusionrelated 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 associated with hydrothermal lode gold deposits.

Syenite-associated intrusion-related gold deposits are distally related to major fault zones, and in association with preserved slivers of Temiskaming-like conglomerate rocks. The deposits consist of stockworks of gold-rich veins and zoned alteration and can be found within or at the margins of composite intrusive stocks, satellite dykes and sills, and along secondary faults and lithological contacts away from the intrusions. Ore bodies in these different locations are interpreted to represent proximal to distal components of large magmatic-hydrothermal systems centered on, and possibly genetically sourced from, composite intrusive stocks. These intrusions, and associated mineralization, are contemporaneous with the deposition of Temiskaming-like rocks and usually post-date major D1 regional deformation.

The Detour Lake Mine uses conventional truck-shovel open pit mining. The mine is operated using an Owner-operator mining equipment and labour strategy. Excluding the muskeg, overburden/till top layer, all material must be blasted. Pioneering drilling and blasting is required in the overburden/rock contact. Additionally, during winter months free digging of overburden material is not possible due to frost.

Design Constraints
The design of each one of the three final pits (Detour Lake, West Detour and North Pit) is guided by the respective optimized shells. The final pit design incorporates the geotechnical parameters prescribed by Golder (Golder 2019, 2020).

The Detour Lake pit design incorporates a double ramp access for most of the LOM. The final ramp and principal access will be located in the north wall. The West Detour and North Pit were designed using a single ramp access.

All ramps are designed to accommodate the safe operation of 795CAT super class trucks. Typical road width is 33 m with a total length of 40 m to allow for safety berm and drainage. Ramp widths are adjusted at the bottom of the pit to a one-way lane. Other design parameters include a ramp slope (10%), minimum curve radius (39 m) and construction constraints (super elevation, crowning).

Overburden and Pre-Stripping
Bedrock at both Detour Lake and West Detour is overlain by overburden composed of layers of muskeg, till, gravel, and sand. These overburden layers may be as thick as 40 m. At the start of each mining phase, this material must be stripped prior to the drilling and blasting of rock material. Depending on the material quality, weather and moisture content, this overburden material is typically ‘free-digging’ and does not require drilling and blasting. A portion of the till material (subject to quality control specifications) is stockpiled and used for tailings management area (TMA) construction purposes. The remaining overburden is stockpiled either in the main overburden stockpile or within designated areas of the WRSFs. A portion of this material will be rehandled for the reclamation of various waste rock stockpiles and/or TMA cells. Whenever possible, this overburden is directly placed for the progressive reclamation of ultimate WRSF slopes.

Mining Operations
The Detour Lake Mine uses conventional truck-shovel open pit mining. Excluding the muskeg, overburden/till top layer, all material must be blasted. Pioneering drilling and blasting is required in the overburden/rock contact. Additionally, during winter months free digging of overburden material is not possible due to frost. The mine operation also requires the management of old underground workings.

Underground workings records are considered of good quality. The mine has been operating in historical underground working areas for the past years; the reconciliation of historical recordings and field observation of these workings is very good. The assumption that stopes were backfilled using sand/rock filled has also been confirmed. There are procedures in place to ensure safe operations in these areas. The procedures define zones classified by the risk of ground instability.

The Detour Lake Mine has used a 12 m bench-height since operations commenced. A 6 m sub-bench was used to mine the areas requiring pioneering to improve operational conditions related to boulders and pinnacles of bedrock.

Starting in mid-2020 the mine transited to a 14.5 m bench height for areas to be primarily mined by rope shovels and to 7.25 m benches in areas to be mined using hydraulic shovels. The revised mine design has led to improvements in shovel productivities.

The mine operates 24-hr per day year-round on 12 hr shifts for all operational crews. Operational teams work on a seven-day in/seven-day out rotation. The mine has implemented a successful hot-seating process for its main production equipment (shovel, trucks and drills). Management/supervisory and support personnel work at site on different schedules.

Ore Mining and Stockpiles
Whenever possible, mined ore is delivered directly to the primary crusher in order to avoid unnecessary rehandling. When the mined ore tonnage exceeds the operating capacity of the crusher, the ore is placed on one of the ROM stockpiles for later mill feed.

The ROM stockpiles are located in close proximity to the primary crusher to allow for efficient feeding into the plant. Ore in ROM stockpiles is segregated into piles consisting of similar grade material to allow for controlled feeding as required. The ROM stockpiles provide additional buffering capacity to ensure that crusher feed is maximized while also limiting excessive tramming of loading equipment between ore and waste areas in the pit.

Crushing Circuit
The primary crushing system is 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. The primary crusher has a capacity of 5,000 t/hr at a 138 mm close side setting. Assuming an 80% availability, the primary crusher has a capacity of 96,000 t/d. 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.

Crusher Improvement Initiatives
The primary crusher is a key area where initiatives are planned to generate a finer product that will translate into a higher plant throughput.

The impact of fragmentation on throughput has been well documented and has a very high impact on throughput.

It has been established that each 50 mm reduction in the P80 equals about a 300 t/hr increase in plant throughput. In 2019, the Detour Lake Mine started a drill-and-blast optimization project that has resulted in higher throughput rates and better operational conditions. The initiative is ongoing as the mine aims to further optimize the process.

The choke feeding of the primary crusher is another key element to increase the plant throughput milling rates. It has been well established that the longer a crusher stays full, the finer the product will be (choke feeding).

The primary crusher choke feeding can be improved. The avenues to do this are to improve dispatching of trucks, better fragmentation to avoid bridging of the crusher, an excavator dedicated to the primary crusher re-feed, and improved rock breaker availability and performance.

One of the most important improvements associated with increases in the milling rate will be the addition of screens over the secondary crushers. The first consideration for efficient secondary crushing is to have the correct feed preparation. A secondary cone crusher achieves its best work when fines smaller than the desired crusher product are removed from the feed. This was not done during the initial design, and as a result, the secondary crushers receive direct feed from the primary crusher with all the fines still in the feed.

Grinding Circuit
The SAG mill operates in closed circuit with a pebble crusher while the ball mill operates in closed circuit with hydrocyclones.

The design circulating load from the cyclones to the ball mill is 250% of the SAG mill new feed. The pebble crushing circuit with a design discharge P80 of 13 mm processes the equivalent of up to 25% of the fresh feed.

Around 15% of the cyclone feed material is sent to the gravity recovery circuit. 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 99%. The pregnant solution is pumped to a tank in the gold room followed by electrowinning in a dedicated cell.

Curved Pulp Lifters
As part of the initiatives to increase the milling rates, the SAG mill pulp lifters will be converted to curved from their current radial configuration. Evidence of severe backflow has been seen on the pulp lifters indicating the pulp lifters are not empty when they go back inside the charge.

Re-Feed System After the Secondary Crushers
Every time a crusher goes down, whether the crusher is primary, secondary, pebble, the plant throughput is impacted. The goal with the re-feed system is to generate 750 kt of -2.5” feed annually, which is the equivalent size of the secondary crusher discharge. This material would be re-fed onto the SAG mill feed conveyor during of crusher downtime events.

Pebble Crusher Variable Speed Drives
By its nature, the rate of pebbles a SAG mill generates is not uniform. The rate is influenced by the ore feed size, mill speed and load inside the mill. The result is that the power of the pebble crusher cannot be maximized due to the variation in feed to the pebble crushers. The crusher is gapped so it can receive any surge in pebbles. The average power is therefore maintained around 450 kW while the motor is rated for 750 kW. The pebble crusher is rarely choke fed.

By installing a variable speed drive (VFD) on the pebble crusher motors, the speed can be increased and decreased to maintain a set-point of 550 kW for power draw. This will enable the crusher to be choke fed and reduce the particle size coming out of the crusher. With a reduced particle size, the SAG mill throughput will increase.

Ore Blending
Ore blending is an avenue that the processing plant will use to optimize milling rates. With the open pit supplying more ore than what the mill can process for the next five years, talc ore can be stockpiled and processed opportunistically. That means that when the ore from the pit is soft, talc ore can be stockpiled and reclaimed when the ore will be harder.

Increase Plant Operating Time to 93%
The operating time for the processing plant is calculated when the ore is on the belt feeding the SAG mills. Operating time is typically 1–2% lower than the mechanical/electrical availability. Although plant operating time has improved since start-up, the design rate of 92% has not yet been achieved over an annual basis, instead averaging around 88–89%.

The process plant is based on a robust metallurgical flowsheet designed for optimum recovery with minimum operating costs. The flowsheet is based upon unit operations that are well proven in industry.

The processing plant was designed to process ore at an average throughput of 55,000 t/d or 20 Mt/a, equivalent to milling rates of 2,500 tpoh with operating time of 92% in a 24 hour day. While operating time has lagged at around 89%, the milling has far exceeded the design rate averaging 2,952 t/hr in 2020 for 23.06 Mt milled. On at least 70 occasions in 2020, a 74 kt milled per day rate was achieved. Table 17-1 shows the production rates from 2017–2020.

With the upturn of current performance of the processing plant, ongoing optimization efforts and some new capital initiatives, the LOM plan assumes that the plant throughput will increase from 23 Mt/a in 2020 to 28.0 Mt/a in 2025 and thereafter. The annual plant throughput of 28.0 Mt/a is planned to be achieved by increasing the milling rate to 3,436 t/hr and improving operating time to 93%.

Leach and Carbon-in-PuIp
The gold recovery circuit selected was a “leach circuit 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. Four new leach tanks are planned to be built to support the increase in throughput to 28 Mt/a. The CIP design selected uses a carrousel type system with an average design retention time of 80 minutes in this circuit. Upgrades to the pumpcells will support the 28 Mt/a throughput rate. The VSD and motors for the CIP tails pumps will be upgraded to 500 hp from 350 hp. The current adsorbtion profile and modeling support the planned increase in throughput rate.

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. The acid wash stage is not currently used and will be converted into a plastic removal vessel. The build up of calcium is not sufficient to warrant the use of the acid wash vessel.

The stripping circuit is designed to handle up to 10 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 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.

The current elution capacity is estimated to be 2.5 strips of 10 t per day. In 2020, 1.75 strips per day were completed. There is sufficient capacity to cover the increased throughput for the LOM.

Thickening
The feed to the leach circuit is maintained at a density of 50–55% solids by weight using one high rate thickener. The overflow of the pre-leach thickener is recycled to the process water tank. The pre-detox thickener is used to thicken the leach tails slurry to 55–60% solids and the overflow is recycled to the process water circuit.

The thickener capacity is over 90 kt/d and can support the increase throughput without any modifications. Current flocculent consumption is low at 10 g/t.