The Fire Creek deposit is considered to be a low-sulfidation, epithermal deposit.

Low-sulfidation epithermal systems are also referred to as quartz ± calcite ± adularia ± illite or adularia-sericite epithermal systems. These nomenclatures refer to the oxidation state of the ore fluid sulfur component, gangue mineralogy and hydrothermal fluid pH, respectively. Ore-fluids in a low-sulfidation hydrothermal system are reduced, have a near-neutral pH and are dominated by deeply-circulated meteoric water. These deposits form in the shallow crust, 0.5 to 1.5 miles at temperatures of greater than 300°C in subaerial volcanic settings. Steeply-dipping, open-space veins are common. Quartz is the principal gangue mineral and can be accompanied by chalcedony, adularia, illite, pyrite, calcite, and rhodochrosite. Boiling is the dominant metal deposition mechanism and commonly results in vein textures including crustiform-colloform bands and platy calcite and/or quartz-after-calcite pseudomorphs. Ore metals are usually Au-Ag, Ag-Au or Ag-Pb-Zn and, contrary to the ore-fluid source, metals in NNR-related epithermal deposits are sourced from mantle-derived basaltic magmas (Kamenov et al., 2007).

Zoned hydrothermal alteration comprises widespread and deep propylitization that grades upwards to clay, carbonate and zeolite formation. Proximal alteration comprises quartz, adularia, and pyrite. High-level advanced argillic alteration characterized by clay-carbonate-pyrite or kaolinite-alunite-opal ± pyrite alteration can be present above the ore-grade zone and is the result of steam-heated, acidic, ascending fluids generated during boiling.

Electrun is primarily present in its native state along discrete layers within veins. Native electrum can occur as large clots or bands, dendritic growths, and fine-grained disseminations. Other less common habits include encapsulations in quartz, pyrite replacements and coatings on pyrite or arsenopyrite (Thompson, 2014). Silver occurs encapsulated in quartz and locally in naumannite or ruby silver encapsulations in quartz (Thompson, 2014). Dark grey ginguro bands of an unidentified silver-bearing mineral are present along vein banding as well. The silver gold ratio is approximately one to one.

Mining methods include several different techniques such as end slice stoping with delayed backfill, also referred to as longhole stoping, and drift and fill stoping. The final choice of mining method will depend upon the geometry of the stope block, proximity to main access ramps, ventilation and escape routes, the relative strength or weakness of the mineralized material and adjacent wall rock, and finally the value or grade of the mineralized material. The choice of mining method will not be made until after the stope delineation and definition drilling is completed. Each method will be discussed briefly in the following paragraphs.

End Slice Stoping:

End slice, or longhole, stoping has the highest degree of mechanization of the three expected mining methods at the Project, is the lowest cost method and generally provides the lowest total cost per ounce. End slice stoping requires the greatest amount of waste development and can be mined to a minimum width of four feet. The potential for unplanned wall dilution with this method is the greatest.

The amount of mineralization that can be removed prior to backfilling will be constrained by the strength of the gangue material and jointing present immediately adjacent to the stope. Backfill, consisting of either waste rock or cemented rock fill, will be transported from the surface using the same haulage equipment used to remove mineralized material and waste rock from the mine. Where possible, waste rock will be retained within the mine and placed directly into a stope requiring backfill. The stope will be backfilled from the drift used for drilling and blasting.

Cemented rock fill (CRF), which consists of screened mine waste, fly ash, and cement is mixed on the surface and transported underground in the same trucks used to haul blasted rock to the surface. CRF is placed to create an artificial pillar where additional mining is planned adjacent to or underneath the stope being filled. Normal backfill unconfined compressive strengths (UCS) of 400 to 600 pounds per square inch (psi) are achieved by blending a mixture containing up to four percent cement and fly ash. When mining is anticipated to occur below the backfilled stope, the UCS of the fill will be increased up to 1,000 psi by adding up to eight percent cementitious binder.

This method can be employed where the wall rock is too weak for end slice stoping, the vein dip is less than 50° or where there is variable vein geometry. Drift-and-fill stoping is the highest cost mining method of the two considered.

Drift and Fill Stoping:

A drift and fill stope is initiated by driving a waste crosscut from the access ramp to the vein. The cross cut is driven at a negative gradient up to minus 15% in order to reach the lowest elevation of the stope. Drifting along the vein strike progresses in both directions from the cross cut. Drift dimensions are a minimum of six feet in width and 10 feet high. The width can be increased to accommodate wider sections of the vein.

Once the end of the stope is reached, the drift is backfilled with CRF if there is unmined ore below or with unconsolidated waste backfill (GOB) if mining below is not planned. Once filled, breasting down the waste above the back of the cross cut begins at a gradient sufficient that the sill of the crosscut is now at the same elevation as the back of the preceding drift. This process will be repeated until all the vein within reach from the cross cut has been mined out, and mining will proceed from the next level above.

Cut-and-Fill Stoping:

Cut-and-fill stoping is an option where mineralization extends above the uppermost waste development accesses. A cut and fill stope is initiated by driving a waste crosscut from the access ramp to the vein. The access is then prepared for a timbered raise to advance upward on the vein. The raise consists of segmented compartments which house an ore chute, a manway with ladders, and a small hoist for supplying the stope with necessary supplies. Cut dimensions are a nominal fur feet in width and ten feet high. The width can be increased to accommodate wider sections of the vein. As the cuts are developed, the ore is slushed back to the timbered raise and loaded into trucks at the bottom of the ore chute. Cellular grout is pumped up the raise for backfill prior to breasting down the next cut.

One major advantage of the cut-and-fill method is the reduced need for waste development to access every vertical sublevel. Instead, the ladderways can be driven up to 300 feet vertically without additional level accesses. One major drawback, however, is the cost of cellular fill and timber, as well as slower ore production compared to longhole stoping. The Company has employed cut and fill stoping via timbered raises at its other properties and has developed safe and efficient procedures that can be utilized here as well.

Backstoping:

An alternative to cut and fill stoping, in areas where mineralization extends above the uppermost waste development access, is backstoping. Backstoping eliminates the requirement for a timbered raise to be driven up from the level. It is safer and more productive than cut and fill.

After accessing the vein via a cross-cut, a sill drift is driven in the vein, up to 200 feet long. Blast holes are then drilled up into the mineralized vein, usually on an angle and charged from the bottom. The stope material is then blasted down into the void created by the drift and removed with a remotely operated loader. The height of the backstope is limited by ground conditions and consistency of vein dip angle – and not likely to exceed 60 feet. Some initially mining by backstope is underway.

Cut and fill stoping is more selective than backstoping. Additionally, mining a backstope eliminates the possibility of developing further sublevels above the stope without more waste development. Backstopes are also difficult to fill.

A local contractor transports mineralized material from Fire Creek to the Midas Mill on public roadways which is a distance of approximately 131 miles. Mineralized material from several Klondex mines is processed through the crushing circuit. The mill has two 500-ton fine ore bins located between the secondary crusher and the ball mill and one bin is dedicated to each mine. Head samples are taken on each reclaim conveyor at regular intervals, and tonnage measured by a belt scale prior to comingling the mineralization streams.

The Midas Mill was constructed in 1997 and has a nameplate capacity of 1,200 tpd. The mill uses conventional grind-leach technology with Counter Current Decantation (CCD) followed by Merrill Crowe precipitation. A CIL circuit was added in 2017. Doré refining is finalized by Asahi refineries in Salt Lake City, Utah. Midas has performed toll milling periodically since 2008.

Underground mineralized material is extracted and delivered from Fire Creek and the Midas Mine to the run of mine (ROM) pad where it is placed on short term ROM mineralized material stockpiles. Typical mineralized material classifications are: low-grade less than 0.3 opt Au or less than six opt silver; high-grade (0.3 to 0.5 opt gold or six to 20 opt silver); and ultra-high-grade (more than 0.5 opt gold or more than 20 opt silver). Separate stockpiles are maintained for each mine. Underground mineralized material is hand-picked on the pad for scrap wire mesh and rock bolts before being fed to the crusher.

Mineralized material is crushed in two stages through a 30-inch by 40-inch primary jaw crusher and 53-inch secondary cone crusher. Both jaw and secondary crusher products are fed to a six feet by 20 feet Nordberg double deck vibrating screen fitted with two-inch top deck and one-half inch bottom deck screen panels to produce a 95% passing one-half inch product. Magnetic material is removed from the crusher screen feed by a continuous self-cleaning belt magnet to protect the cone crusher from damage. Screen undersize is conveyed to one of two 500 ton fine mineralized material bins.

Crushed and screened material is transported from the fine material bins by individual belt feeders into the 10.5 feet by 15 feet rubber lined Nordberg ball mill. The ball mill is charged with a blend of three-inch and two-inch grinding balls to maintain an operating power draw of 800 horse power (HP). Mill discharge pulp is pumped to a nest of four ten-inch Krebs cyclones (three duty and one standby) for classification. Cyclone overflow, at 80% passing 200 mesh, reports to the trash screen.

Cyclone underflow reports to a two millimeters (mm) aperture scalping screen, with the screen undersize being distributed by three-way splitter to the ball mill, verti-mill, and gravity circuit. Lead nitrate solution is added to the ball mill feed chute to enhance silver leach kinetics.

A split of the screened cyclone underflow reports to the 250 HP verti-mill for open circuit grinding with the verti-mill discharge overflowing back to the primary ball mill discharge pump box. The verti-mill is charged with one inch grinding balls. A split of the screened cyclone underflow also reports to the 20-inch Knelson concentrator for gravity gold recovery. The Knelson operates on a 30-minute cycle providing concentrate for cyanidation in the CS500 Acacia Leach Reactor which conducts three 750 to 1,000 kg batch leaches each week. Pregnant solution from the leach reactor reports to the CCD circuit pregnant solution tank.

Cyclone overflow is screened to remove any plastic debris before reporting to a 42.5 feet diameter pre-leach thickener. Thickener underflow at 50% solids is pumped to the leach circuit consisting of eight 28 feet by 30 feet air sparged leach tanks, providing a leach residence time of approximately 90 hours at 600 tons per hour (tph) feed rate. The pH in the first leach tank is maintained at 10.4 to 11.0 through the addition of hydrated lime, produced from the on-site slaking of pebble lime. Sodium cyanide concentration in the second leach tank is maintained at 1.25 grams per liter (gpl).

The leach circuit discharge is pumped to the first of five 42.5 feet diameter CCD thickeners, where the pulp is counter-current washed with barren Merrill Crowe liquor at a wash ratio of approximately 3.2:1. CCD thickener underflow at each stage is maintained at between 50 and 54% solids to maximize wash efficiency.

Pregnant CCD solution at a pH of 11.0 and 400 gallons per minute (gpm) flow rate is fed to one of two disc filters operating in duty/standby mode utilizing diatomaceous earth for clarification. The clarified pregnant solution is then pumped to a packed bed vacuum de-aeration tower, prior to the addition of zinc dust and lead nitrate to precipitate precious metals from solution. The Merrill Crowe solution is then pumped to one of two plate and frame filter presses for sludge recovery. The sludge is collected from a filter press weekly and smelted to produce 5,500 ounce silver and gold doré bars.

Tailings pulp from the last CCD thickener is pumped to the Inco SO2/Air circuit for cyanide destruction. Cyanide destruction is performed in a single 20 feet by 20 feet agitated, air sparged tank providing approximately one hour reaction time. Ammonium bi-sulphite, lime, and copper sulphate as a catalyst are added to the tank on a ratio control basis to achieve target weak acid dissociable (WAD) cyanide permitted levels. Routine picric acid analyses are used by operating personnel to maintain WAD cyanide in the INCO cyanide destruction tank discharge pulp at target levels.

Following cyanide destruction, the plant tailings pulp is thickened before discharged to one of two lined tailings storage facility (TSF) for consolidation and water recovery. Clarified decant pond solution is evaporated or returned to the mill process water tank for reuse in the plant.