The Coringa Gold Project is underlain by granitic intrusions of the Maloquinha group and rhyolites of the Iriri group (Salustiano Formation). The granites are granular, medium-grained, and consist of pink feldspar and quartz. The rhyolites are fine to medium-grained, porphyritic, and strongly magnetic. Sanidine and quartz phenocrysts occur in a fine-grained matrix of sanidine-quartz. Minor amounts of biotite also occur in the matrix which has been altered to chlorite.

There are two dominant structural trends on the Coringa Gold Project property:
• The 310° structures are interpreted as strike-slip faults with probably a dextral (right lateral) sense of displacement.
• Structures trending at 345° are interpreted as R-shears.

Mineralized veins at the Coringa Gold Project are associated with the R-shears. The dip of the veins ranges from 750 to the east to vertical, but they occasionally dip steeply westward (e.g., Galena Vein).

Mineralization at the CoringaGold Projectis associated with a shear/vein system that has a strike length of over 7 km. The mineralized zones vary in thickness from <1 centimeter (cm) up to 14 meters. Several veins (i.e., Galena, Mãe de Leite, Meio, and Come Quieto) occur along the main mineralized corridor and others, such as Serra, Demetrio, and Valdette, form subparallel zones. The average thicknesses for the veins included in the estimate of mineral resources are: Serra 0.52 m, Galena & Mãe de Leite 0.59 m, Meio & Come Quieto 0.41 m, and Valdette 0.80m.

Gold mineralization is almost exclusively associated with quartz-sulphide veining. Pyrite is the main sulphide, but minor concentrations of chalcopyrite, galena, and sphalerite are common. A genetic study of mineralization indicated that pyrite-chalcopyrite (+/- quartz) mineralization occurred first, followed by gold, with galena and sphalerite introduced late. Gold is typically free (or within electrum) and occupies fractures within sulphide grains. It is usually very fine grained and visible gold is rare. Gold in electrum is closely associated with quartz and pyrite. The bulk of the gold has a preference for deposition in the quartz matrix/groundmass (48% locking affinity) and within pyrite (31%) occurring in either fractures or as inclusions, as well as in other sulphides, oxides, and, to a lesser extent and depending on tectonic conditions, in silicates.

The mineralized veins exposed on the Coringa Gold Project are similar to those found in Orogenic gold deposits. These deposits formed over a 3 Ga time frame with peaks at 3.1 Ga, 2.7 to 2.5 Ga, 2.1 to 1.8 Ga, and 0.6 to 0.05 Ga corresponding to the episodic growth of juvenile continental crust. These deposits were formed during the Achean eon of the Precambian and are commonly referred to as Archean lode gold deposits. A large percentage of the world’s gold resource is associated with these periods.

In the Coringa gold deposit, shear zones of anomalously high strain are clearly seen and are mappable units (Global Resource Engineering, 2012). Gold deposition occurs within the quartz veins which were emplaced in the secondary extensional structures associated with the primary shear zones. These shear zones (linear units) occur in generally predictable orientations and are located in certain preferred settings, that is perpendicular to the maximum tension direction.

Ore zones are lenticular, tabular or irregular shaped bodies composed of veins, breccias zones, and/or stockwork systems. Veins transect lithological contacts and are not restricted to a specific rock type. Veins can be classified as replacement, extension, breccias, and fracture type veins. There is also a vertical zonation of the gold deposit, which reflects a change in deformation style, from brittle to brittle-ductile.

The underground mining method at Coringa will involve driving vertical raises from drifts driven through the vein, and subsequently breasting along the strike of the vein on each side of the vertical raise. The primary equipment used for production will be jackleg and stoper drills. This method was chosen due to its ability to effectively mine the narrow vein deposit while minimizing dilution. Although this mining method does utilize a large number of drillers which leads to a higher labor cost, it does eliminate the larger initial capital investment needed when using more advanced mining equipment. This proven mining method is currently in use at the geographically and geologically similar Palito mine, which is also owned by Serabi.

Vein systems have been broken into separate mining areas. This separation was based upon a trade-off analysis between a separate portal and ramp versus drifting over to the adjacent continuous stope area within each vein system. Each mining area will require its own portal, ramp, ventilation, dewatering, and underground utilities. The Serra vein system requires only one portal due to the continuous grade distribution within the planned stopes.

The production sequence was determined by mining the higher grade resource areas first.

Stopes are 30 meters vertical along dip and 32 meters horizontal along strike. Stopes are mined from a vertical raise driven from drifts. Each stope will contain eight stope raises. Access to these vertical raises is accomplished by a installing a series of ladders and timber stulls from the bottom of the production drift upwards. Two drifts will be driven through the vein in each stope. Drifts will be spaced at 15 meters vertically. Development of the vertical raises is accomplished with stoper drills with the subsequent horizontal breasting by jackleg drills. For each stope, a sill pillar approximately 1.5 meters thick spanning the length of the stope will be left in place, resulting in an extraction ratio of 95%. GRE has not examined the option of pillar robbing at the end of the mine life to recover the additional 5% of the resource. Within a stope, the lower drift and ore passes are developed before main production begins on the upper drift. This permits main production from multiple stopes through tramming of mineralized rock on the main production drift to ore passes spaced every 2 stope lengths.

It is important to note that the stope design for the PEA used a uniform rectangular stope shape and grid pattern. This approach included additional dilution along the periphery of each designed stope area. It also excluded areas above the cutoff grade. Actual production planning for operations will be more detailed that this simplified approach. Once complete, the detailed operational mine plan will both minimize the amount of dilution and include adjacent areas above the cutoff grade with a minimal amount of additional development.

The development sequence was designed to ensure that all required development for a level is completed before the stopes on that level are scheduled for production. This was done in order to avoid access and equipment conflicts between production and development on a level simultaneously.

Ramps will be driven at a 12% decline, this means that approximately 252 linear meters of ramp are driven for each 30 meter level. A 10% allotment was included for muck bays, cutouts and passing areas to alleviate potential vehicle congestion. Ramps will be spaced far enough from stopes to not cause geotechnical issues and will be driven in the waste rock of the footwall. Ramp crosscuts will connect the ramp decline to the footwall of the vein. One ramp crosscut will be developed per level of each mining sub area.

Two methods are available for advancing drifts. The first method consists of full face blasting where the vein is diluted to the full volume or rock required to advance the drift. The second method of resue mining involves driving the drifts in two separate blasts. The first blast removes the vein. After the broken vein is mucked out, the second blast removes the waste rock inside the drift heading. GRE estimated that resue mining will be used approximately 50% of the time when driving ore drifts, and that the full face blasting method will be used 100% of the time when driving waste drifts. These assumptions are based on Serabi’s mining experience at their Palito Mine.

The relative advantage of the full face drifting method is that the production rate is faster. The advantage of the resue mining technique is reduced dilution and higher grade material for the process plant; however, the advance rate is cut in half by adding a second blast.

Muck bay development was accounted for as an additional 10% of the total drift length per level. Muck bays will also be utilized as definition drilling stations; however, additional drilling stations may be developed as needed.

It is assumed that secondary escapeways and vent raises can be driven at the same time since they are both located at the ends of the mining blocks. Both the escapeway and vent raises will be driven outside of the mineralized vein in the hanging wall. In some cases vent raises will also be used as secondary escape ways. Escapeways will have ladders with landings spaced as required by applicable regulations.

Ore chutes will be spaced approximately every 60 meters along levels where they connect vertically to the level below. Ore chutes will be driven outside of the vein in waste rock of the footwall.

The processing facility for the Coringa Gold Project is a conventional gold cyanidation plant. It has been designed to treat 645 tpd (212,000 tpa) of ore containing 7.5 gpt gold with minor silver over a 11-year period. Annual gold production at full capacity will average 38,000 ozs. The gold doré product will be shipped to a refinery for further processing.

The process plant will be a combination of new and refurbished equipment, tanks and structures. Asimilar sized crush/grind/gravity/leach gold ore processfacility, located in Brazil, was purchased and relocated to the Coringa Gold Projectsite for re-use of the suitable equipment and materials.

Metallurgical test results of representative material from the Coringa Gold Project deposits were utilized to develop the final process flowsheet and plant design criteria.

The process plant incorporates the following standard unit process operations:
• ROM ore stockpile area and reclaim hopper
• Primary and secondary crushing with screening
• Ore stockpile with reclaim feeders
• Parallel single-stage ball milling in closed-circuit with cyclones
• Knelson gravity concentrator and Acacia IL reactor for concentrate leaching
• Cyclone overflow to trash removal screen
• A lime/pre-aeration mix tank prior and cyanide leach tanks
• CIP tanks equipped with carbon screens and a safety screen
• Cyanide destruction tank using SO2/Air process
• Tailings thickening and filtration followed by dry stacking of the filtered tailings
• Carbon wash/elution/strip/regeneration circuit with a 1.5 t carbon capacity
• Reagent storage, mixing, and distribution systems
• Electrowinning cells for Acacia and stripped carbon solutions to recover gold and silver
• Smelting and doré handling and security systems

The ROM ore is stockpiled and then reclaimed by front-end loader. The loader dumps the ore into a hopper equipped with a vibrating feeder that discharges into an 800 mm by 600 mm primary jawcrusher.

The jaw crusher product discharges onto a conveyor that feeds a 4 m long by 1.5 m wide double-deck vibrating screen. The oversize from the top and middle decks are combined and fed via a conveyor to two 3-foot diameter cone crushers, operating in parallel. The crushed material from the secondary crushers is recirculated via conveyor back to the vibrating screen. The final crushed product (undersize from the screen bottom deck), at an average particle size of 80% passing 12 mm, discharges onto a conveyor that feeds the fine ore storage bin.

Crushed ore is reclaimed from the fine ore bin via a belt feeder and conveyors that feed two ball mills operating in parallel - a 4.3 m long by 3.15 m diameter ball mill equipped with a 720 kW motor, and a 3.6m long by 2.7m diameter ball mill equipped with a 355 kW motor.

The ball mill grinding operates in closed circuit with a cyclone pack which classify the ground ore to a final particle size of 80% passing (P80) 75 microns. The cyclone overflow at a target 40% solids by weight, is passed over a trash screen and then is directed to the carbon-in-pulp (CIP) circuit. The cyclone underflow is split and returned back to the ball mills.

A bleed of the mill discharge slurry is pumped to a centrifugal gravity concentratorforfree gold and silver recovery. The concentrator tails are returned to the mill discharge hopper. The gravity concentrates flow to an Acacia intensive leach reactor (ILR). Acacia leach tails are returned to the grinding circuit and the ILR leach solutions are pumped to a tank feeding a dedicated electrowinning cell.

The CIP circuit consists of one pre-aeration tank, two leach tanks and seven adsorption tanks all in series. The pre-aeration tank receives slurry from the grinding circuit for aeration and pH adjustment to about 10.5 using hydrated lime. The slurry then flows to the leaching tanks where cyanide is added and the gold and silver are extracted into solution.

After leaching, the slurry flows downstream from tank to tank through a series ofseven adsorption tanks. Each adsorption tank, containing granular activated carbon, is equipped with a static intertank screen. The leached gold and silver are adsorbed by the activated carbon present within the tanks.

The metal-loaded carbon is transferred countercurrent to the slurry flow from the last tank forward. The highest metal loaded carbon resides in the first CIP tank. From the first tank the carbon is transferred to a vibrating screen for washing, then directed to the desorption column for further washing and metal stripping. The screen underflow returns to the first CIP tank. Total leaching residence time is 27 hours and total adsorption residence time is 19.5 hours at the designed flows and slurry density.

In the desorption column, the carbon is first acid washed with a weak solution of hydrochloric acid and then rinsed in a caustic soda solution. A Zadra stripping system is employed. In practice, a solution containing about 1% sodium hydroxide and 0.1% sodium cyanide at about 280 deg-F and 65 PSIG, is circulated through a pressure vessel filled with loaded carbon at a flow rate of 2.0 bed volumes per hour. The time required for pressure stripping is generally from 10 to 14 hours.

This strip (pregnant) solution is pumped through a dedicated electrowinning cell where gold and silver are deposited on cathodes. The cathodes are periodically removed from the cells and washed to recover the gold/silver sludge. The precious metal sludge is dried, mixed with flux reagents and then smelted to produce a doré product. The doré bar is then sampled and shipped offsite for final refining. The barren electrowinning solution is then recycled to the leaching circuit.

After stripping, carbon is washed with water and transferred to the regeneration kiln. The carbon is heattreated in the kiln and then returned to the last adsorption tank after screening for particle size control.

The slurry from last CIP tank, after passing through the carbon safety screen, is pumped to the cyanide neutralization circuit that utilizes the SO2/Air process to destroy cyanide in the tailing slurry. Copper sulphate and SMBS are added to the aerated mix tank to destroy the cyanide.

The detoxified slurry is pumped to a 15m diameter hi-rate thickener where the slurry is thickened to 60% solids. The thickened solids are pumped via surge tanks to a filter press where they are filtered to remove residual process solution. The process solution is recycled back to the process and the filtered tailings, containing approximately 15% moisture, are conveyed to a load-out shed. The filtered tailings are loaded and transported, using a front-end loader and truck, to a remote dry stacking tailings storage facility.

Any make-up water requirements will be provided from local sources of fresh water including mine water and site runoff collection.