The Soledad Mountain gold-silver deposit is best interpreted as a volcanic rock-hosted, low sulfidation, epithermal vein system of the low base metal type. Individual major veins formed by episodic deposition of quartz, adularia, sericite, calcite, and sparse sulfide minerals in open faults and fractures, coeval with adjacent sheeted and quartz-vein stockwork zones and quartz ±adularia ±sericite alteration of nearby wall-rocks. Other examples of this deposit class include districts such as Oatman (Arizona), Bullfrog (Nevada), Bodie (California), and Tayoltita (Mexico). Soledad Mountain contains an unusually large number of individual veins within a relatively small area by comparison to the examples cited. Post mineral faulting at Soledad Mountain has extensively sheared and brecciated the veins, most likely due to mid-Miocene to present-day wrench-fault tectonism between the nearby San Andreas and Garlock fault systems.

Gold and silver mineralization at Soledad Mountain occurs in a swarm of mainly northweststriking, subparallel to anastomosing, low-sulfidation, epithermal quartz veins that formed in faults and fractures within the Miocene rhyolitic volcanic units. Over 20 gold silver veins and related vein splits have been identified and modeled as part of the project resources. Veins generally strike N40°W and dip at moderate to high angles to the northeast and to the southwest, and occur in parallel and, locally, en echelon patterns over a total strike-length of 7,000 ft and a total width of 4,500 ft.

Mineralization consists of finegrained pyrite, covellite, chalcocite, tetrahedrite acanthite, native silver, pyrargyrite, polybasite, native gold and electrum within discrete quartz veins, veinlets, veinlet stockworks, and irregular zones of silicification. Gangue minerals include quartz, potassium feldspar (adularia), ferruginous kaolinitic clay, sericite, hematite, magnetite, goethite, and limonite.

The veins formed by intense alteration of volcanic rocks and by deposition of quartz and sericiterich material in fault and fracture zones. The alteration and veins are generally low in sulfur, with total sulfide content generally being 1% or less. Vein “zones” consist of one or more central veins surrounded by either a stockwork or parallel zones of sheeted, narrow quartz veins. The effect is to have a core vein of 1 ft to 20 ft in width (with gold grades being generally greater than 0.1 oz/ton), surrounded by lower grade mineralization in the adjacent quartz-vein stockwork and sheeted vein zones. The widths of the stockwork and sheeted vein zones vary from 5 ft to 150 ft. The boundary between mineralized and non-mineralized material must be determined by assay.

Native gold and electrum are generally associated with siliceous gangue and occur as particles with diameters ranging from less than 10 µm to as much as 150 µm or more. Electrum contains about 25% silver. Gold grades greater than 0.1 oz/ton appear to occur where veins exhibit multiple generations of quartz, adularia and sericite. Sheeted veins and stockwork veinlets decrease in grade laterally outward from the core veins. Silver to gold ratios vary from 1:1 in shallow portions of veins in the south half of the deposit to greater than 35:1 at deeper levels (600 Level) in the north half of the deposit. Silver to gold ratios increase generally with depth, averaging about 10:1 at the surface of the Golden Queen vein, to about 35:1 at the 600 Level in the same vein. There is also a general horizontal zonation, from relatively silver rich in the northeastern vein systems (e.g., Queen Esther – Independence), to gold-rich structures in the southwest (e.g., the Sheeted Vein system). The district silver-to-gold ratio average ranges from 15:1 to 18:1.

Alteration within mineralized zones consists of fracture-controlled and disseminated fine grained silica, adularia, sericite, and minor pyrite. Intense quartz-feldspar-sericite alteration reportedly occurs in zones from about 10 ft to over 150 ft wide. Volcanic rocks are weakly silicified and argillically altered between and adjacent to zones of strong silicification. Weakly silicified and argillically altered rocks grade laterally into weakly to strongly propyllitized + illitized volcanic rocks. Propylitic alteration is best developed in the quartz latite flows.

Important vein systems, from the northeast to southwest, are the Black, Reymert, Karma-Ajax, Independent, Queen Esther, Silver Queen, No. 1 Footwall, Golden Queen - Starlight, Soledad, Alphason, Gypsy, Echo, Hope, Elephant, Bobtail, Excelsior, and McLaughlin. Post-mineral offset of about 300 ft on the east-dipping, apparently listric, Main Fault has displaced the Starlight vein in the footwall, from its upper continuation known as the Golden Queen vein in the hanging wall. Portions of the Soledad and No. 1 Footwall veins are also displaced by offset on the Main Fault. Veins northeast of the Golden Queen vein dip from 40° to 70° northeast. Veins southwest of the Golden Queen Vein dip about 70° southwest.

A zone of “Flat Ore” is present between the Starlight and Silver Queen Vein, in the hanging wall of the Main Fault. Flat Ore is a complex zone of veins and stockwork mineralization that is from 100 ft to 125 ft thick and nearly horizontal that at least in part consists of blocks of the mineralized zones cut by the Main Fault. Individual, parallel and en-echelon vein systems are present over a total strike length of 7,000 ft trending northwest, and a total width of 4,500 ft. Veins and vein zones are from 5 ft to 150 ft in thickness, 325 ft to 3,000 ft long, and from 300 ft to 1,000 ft in extent along dip. The horizontal distance between individual veins is from 50 ft to greater than 400 ft.

The operation will be an open pit operation. Wheel loaders, a hydraulic excavator and haul trucks with a capacity of 100 ton will be used as the primary mining equipment. Smaller equipment will be used for pioneering access roads, and mining narrower benches, and final ore extraction at the bottom of the various mining phases. Support equipment such as a grader, a water truck and tracked dozers and a wheel dozer will be used for road and bench maintenance, dust control and work in the waste rock disposal areas.

Inter-ramp slope angles for the majority of pit walls are designed at 55 degrees for walls up to 500 feet (152 meters) in height. Pit walls higher than this have been designed for an overall slope of 45 degrees by incorporating step-outs and increased catch bench widths. During the 2014 review of the design guidelines, it was recommended that the pit wall angles in zones of adverse structure or lower quality rock mass associated with the central portion of the West Pit be designed with a lower overall slope and a value of 40 degrees was selected for this area only.

Run-of-mine ore will be delivered to the crushing-screening plant located south of the Phase 1 heap leach pad. The crushing-screening plant consists of a three-stage crush and is sized to process 5.1 million tons (4.6 million tonnes) of ore per year.

The crushing-screening plant includes a primary and secondary cone crusher, primary screen, a HPGR as the key comminution device and the required ore chutes and conveyors.

The HPGR discharge will be conveyed to an agglomeration drum where cement and process solution are added, and then conveyed by overland conveyor and a series of grass-hopper conveyors to a stacker and placed on the heap leach pad.

Gold and silver will be recovered by dissolution in a dilute sodium cyanide solution and then by recovery in the Merrill-Crowe process, which includes the typical clarification, deaeration, precipitation with zinc dust, and filtration, drying (retorting), and smelting of the precious metal sludge into a doré product.

An assay laboratory is included on site and is sized to handle from 50 to 200 solid and 10 to 50 solution samples per day on one operating shift.

The heap leach pad is a multi-lift single-use pad. The design considered development in four stages to minimize initial capital costs and improve solution management. The ultimate pad capacity is approximately 51.6 million tons (46.8 million tonnes) and is designed for an ultimate height of 230 ft (70 m). Individual lifts have been designed for 33 ft (10 m) nominal in height.

An overflow pond has been designed based upon the water balance for the Project for a capacity of 28.5 million gallons (108,000 m3). This capacity is adequate to contain heap drain-down and runoff from storm events.

Once prepared, the heap surface will be irrigated with dilute cyanide solution by drip emitters, for a primary leaching cycle of 70 days. Additional underlying lifts will continue to leach to reach the ultimate recoveries for gold and silver. The leachate or pregnant solution and the recycle solution will be collected in a network of perforated pipes and will be directed to a pump box, and will then be pumped to the Merrill-Crowe plant.

Gold and silver are typically recovered by dissolution in a dilute sodium cyanide solution and then by precipitation with zinc or adsorption on activated carbon. The zinc precipitation process, referred to as the Merrill-Crowe process after its developers, is used to recover gold and silver when the silver to gold ratio is greater than 10:1. This ratio is expected to average 10:1 for the Project (range 5:1 to 15:1) and ratios greater than 30:1 were noted in test work. The MerrillCrowe process is well established and the process is highly efficient.

In the Merrill-Crowe process, suspended solids and dissolved oxygen must first be removed from the pregnant solution. Clarifying filters are used to remove the suspended solids to less than 1 ppm. A vacuum tower is used to remove dissolved oxygen from the clarified solution. Zinc dust is metered into the deaerated solution and combines with the gold and silver cyanide complexes in a rapid, cementation-type reaction and precipitated as micron-sized particles of metallic gold and silver.

After precipitation, the solution is pumped to plate and frame filter presses where the gold and silver precipitates are removed. These filter presses are located in the refinery and this is where all subsequent processing takes place. As with gold and silver, any mercury present in the pregnant solution is also precipitated. The precipitate is removed manually from the filter presses, placed in retort pans and these are loaded in a mercury retort. The precipitate is heated to the point where any mercury that is present will be converted to mercury vapors. The mercury retort operates under a vacuum. Water and mercury vapors are drawn through a water-cooled condenser and a mercury trap and any mercury is finally collected in bottles for disposal.

The condensed mercury retort off gas stream requires additional cleaning before being discharged to the atmosphere. A carbon adsorber that uses sulfur-impregnated carbon can typically remove 99.99% of residual mercury from the condenser gas stream.

The dried precipitate is mixed with selected fluxes, typically silica, borax and soda ash, sodium nitrate and melted in an induction furnace. When melted, these fluxes form a slag. The slag is cooled and crushed and occluded particles of gold and silver are recovered by gravity for further processing. The molten mix of gold and silver, i.e. the doré, is poured into a series of cascading molds. Doré is cooled, cleaned and shipped to a commercial refinery where gold and silver bullion are produced for final sale.

The barren solution is pumped to the barren solution tank and is returned to the heap.

For closure, cyanide concentrations in the leach solutions must be reduced to the weak acid dissociable (WAD) standard of 0.2 ppm (0.2 mg/L) and a pH ranging from 6.0 to 8.5. To address this, a staged rinse with fresh water will be run, and natural degradation processes will contribute to removing residual cyanide. Solutions from each cell of the heap and from the lysimeters and leak detection monitoring points will be sampled regularly and taken to the assay laboratory on site for analysis.