Northern Nevada has a large number of very significant gold deposits. These are primarily related to the Carlin, Getchell, Battle Mountain – Cortez, and Northern Nevada Rift (NNR) trends. Two of these trends may be projected through or adjacent to the Mine area. The style of mineralization relating to these trends include the Epithermal Low Sulfidation mineralization along the NNR and the Carlin type mineralization.

In the Mine area, exploration and mining has been focused on three types of mineralization: - Mercury in laterally extensive near-surface replacement silica zones and more localized sinter deposits, principally in the middle tuff unit;
- Disseminated gold in the middle and lower tuff units, andesite, and the Ordovician Vinini Formation beneath the hot spring mercury deposits; and
- Deeper high-grade gold-silver quartz-adularia veins and fracture zones in the Vinini Formation.

Hot spring-related silicification produced surface sinters, silicified bodies beneath the sinters, and widespread, generally bedding parallel, silicification in volcanic and sedimentary rocks. Cinnabar is the only ore mineral in these silica-rich deposits, and it is intergrown with chalcedony and opal (Wallace, 2003). All of the principal areas of known gold mineralization in the Mine area are associated with these silicified areas, occurring beneath them in underlying Tertiary or Paleozoic rocks at various depths below the principal silicified horizons.

Gold in the Mine orebody and in gold-bearing zones beneath the nearby Velvet and Clementine mercury mines is disseminated in the middle and lower tuff units and in an andesite. The lower tuff unit beneath the andesite is the main host rock at Hollister for disseminated mineralization. The deposit also occurs at a major facies interface, where lobes of basaltic andesite that extend from the southwest interbed with tuffs to the east, tapering north-easterly in the area of the deposit, and likely exerting a combination of both permeability-porosity control and chemical control on the position of mineralization (Rhys, 2014). Within the open pits, structural control to mineralization is secondary to stratigraphic control.

The open pit at the Mine exposes both oxidized and supergene parts of the orebody. Most of the tuffaceous rocks have been oxidized, and primary sulfides are preserved in parts of the andesite, the lower tuff unit, and in the Vinini Formation. Drill-core samples from the sulfide zone show that electrum and pyrite are the primary ore minerals, and adularia, quartz, and chalcedony are the major gangue minerals (Deng, 1991). Quartz veins in the Miocene rocks are uncommon and usually small.

Veins at the Mine contain the bulk of the high grade (greater than 0.3 opt Au) mineralization known within the property. The principal veins, Clementine and Gwenivere, comprise semi continuous vein systems with internal ore shoots, and local echelon steps or splays. The veins in mineralized areas typically range from 0.5 to 2.0 feet in width, but locally can be more than 5.0 feet in width. The veins are almost entirely hosted below the TertiaryOrdovician contact. Veins trend west northwest with steep dips and define a vein system over a strike length of more than 2,000 feet. Overall Ag:Au ratios vary within the vein system, generally increase downward, as is common in many epithermal districts.

The Hollister veins are narrow but long in strike length with ore in shallow plunging oreshoots which developed between 300 and 500 feet below the Ordovician-Miocene unconformity. The veins are dominantly extensional in character and show no significant displacement of the host stratigraphy across them. Miocene extensional west-northwest trending faults in the Vinini Formation are the primary control on vein mineralization.

Vein style is typical of low sulphidation epithermal veins and is closely comparable in mineralogy and texture to vein systems developed at the Midas deposit to the north, as well as other Miocene aged epithermal vein system in the NNR (Rhys, 2014). Metallic species in the veins include pyrite, marcasite, Ag-selenides and minor base metal sulphides (including chalcopyrite, sphalerite). Ag bearing phases, as determined by petrographic studies by Larson (1998) and Peppard (2002), are dominated by naumannite, with lesser acanthite, aguilarite, tetrahedrite, jalpaite, stromeyerite, cerargyrite, cervellite, native silver and uytenbogaardtite in decreasing abundance. Au is present mainly in electrum (Larson, 1998). Late Fe-Mn carbonate, barite and fluorite are locally present, the former two often in late voids and vugs.

The Clementine vein, the most continuous and longest vein, was the source of approximately half of the gold produced during underground mining to-date. The Gwenivere vein system contributed approximately 35% to gold production. Other veins that mainly surround the eastern Gwenivere vein, and parts of the Gloria vein in western portions of the mine, contributed the remaining portions of the Au-Ag mineralization. Oreshoots in the principal veins are developed for up to 600 feet below the Tertiary-Ordovician contact principally within the Paleozoic sequence. The Gwenivere Vein has been defined for roughly 700 feet along strike, 300 feet down dip, and is typically 0.5 to 2.0 feet wide, with local segments reaching 5.0 feet wide. The Clementine Vein has been defined roughly 500 feet along strike, 600 feet down dip, and is typically 0.5 to 2.0 feet wide. The Gloria Vein has been defined roughly 200 feet along strike, 100 feet down dip, and is typically 0.5 to 2.0 feet wide. Veins generally have a moderate degree of continuity, with the best grades occurring in ore shoots.

Access to the mining areas will be from haulage drifts; up to 15 feet wide and between 15 feet to 17 feet high. Drift gradients vary from ±15% to reach the desired elevation.

Mining methods used are longhole stoping, cut and fill stoping, or shrink 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.

Longhole stoping has the highest degree of mechanization of the three expected mining methods at the Mine and is the lowest cost method. This method requires the greatest amount of waste development and can be mined to a minimum width of three feet. The potential for unplanned wall dilution with this method is the greatest.

Stope development drifting is planned at nominal five feet wide and 10 feet high to accommodate the production drill and load haul dump (LHD) scoops. Levels are located at 38 foot vertical intervals to control dilution and may be increased as experience is gained in mining the Gloria veins. Stope widths are designed to a 3.5-feet minimum mining width or the vein width plus 1 foot of dilution.

To prepare an area for longhole stoping, access for the mobile equipment must be developed to each level. Mine utilities for communication, water, electrical power, and compressed air must also be provided through the access development. The minimum pillar height achievable with this method is 28 feet and is limited by the stability of the intervening pillar between levels and planned mining width. Mining will progress upwards from the lowest level of the stope block. Drilling and blasting will be carried out from the drift above the active stope while the broken mineralized material will be removed from the bottom drift. The loader used for excavation is equipped with line of sight remote control to allow the removal of all blasted rock without exposing the operator to the open stope and the potential risk of ground falls.

Once a stope is emptied and laser scan surveyed, cemented rock fill (CRF) which consists of mine waste, cement, and selectively, fly ash, will be mixed on the surface and transported underground in the same trucks used to haul blasted rock to the surface. CRF will be 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 300 to 500 pounds per square inch (psi) will be achieved by blending a mixture containing up to 4% 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 8% cement binder.

Cut and fill stoping was the preferred method used by previous owners of Mine. 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 housing an ore chute, a manway with ladders, and a small hoist for supplying the stope with necessary supplies. Cut dimensions are nominal 3.5 feet in width and 6.0 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, 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. As many of the previously developed stopes have already begun with the timbered raise accesses, the Mine can easily continue the cut-and-fill stopes until the remaining ore is extracted. The authors personally inspected existing timber raises and found them to be in excellent condition.

In the rare instances where not backfilling one stope will not affect any adjacent ore, it is possible to employ shrinkage stoping. Shrinkage stopes will likely be accessed from the same timbered raises as the cut-and-fill stopes, but rather than slushing all the ore back to the ore chute during the development of each cut, a minimal amount of swell will be slushed off and the broken ore will be used as a platform to breast down the next cut. Once every cut of the shrinkage stope is developed, the entire stope is mucked out from drawpoints on the access level, leaving a void that is not backfilled.

Klondex will process all of the ore from the Mine at the Midas mill located 18 miles from the mine, saving significant money in ore transport costs.

This flow sheet has no allowances for the processing of preg-robbing ore. Recent test work indicates blending Mine ore with non-preg robbing ore and a blinding agent will improve the gold recovery from 84.5% to above 90%. Based on the test work, the Mine component of the mill blend will be limited to a maximum of 40%.

To maximize the Mine component of the total mill tonnage, it was decided to retrofit the Midas mill to include a CIL circuit. This will allow the Midas mill to process 100% Hollister ore instead of limiting the tonnage to only 40% of the total feed. The mill flow sheet changes include;
- Retrofitting 4 leach tanks with carbon retention screens;
- The addition of carbon advance pumps in each of the 4 tanks; and
- The addition of a carbon safety screen and loaded carbon screen.

Using the adjusted flow sheet, incoming ore will be ground and thickened according to the existing flow sheet. The slurry will be leached in the four leach tanks with the carbon present and pumped directly to cyanide detoxification, bypassing the CCD and zinc precipitation circuits. Loaded carbon will be pumped from the first CIL tank to storage bags and transported to the Aurora mill for stripping and regeneration. The regenerated carbon will be shipped back to the Midas mill for reuse. The leach circuit is capable of either leaching pregrobbing ore (using the four refurbished tanks) or leaching non-preg-robbing ore by removing the carbon in the CIL tanks and using all eight existing leach tanks with the CCD circuit.