The Tocantinzinho deposit is best classified as a granite-hosted, intrusion-related gold deposit. It has geological similarities to several other gold deposits of the Tapajos region. Host rocks include multiple texturally varied phases of dominantly granitic composition. Gold is disseminated, commonly shows consistent grades over wide intervals, and is generally accompanied by finely to coarsely disseminated sulphides (pyrite, locally chalcopyrite, galena, sphalerite). Isolated multi-ounce gold grades occur in intervals containing conspicuous quartz-galena sphalerite-chalcopyrite veins, and are not volumetrically significant. Most gold-enriched zones contain common thin (<1 cm) veinlets of quartz, chlorite, sulphide, ± calcite.

The Tocantinzinho deposit forms a sub-vertical, northwest-trending elongate body approximately 1,000 metres long by 150 metres wide. It has been drilled to approximately 350 metres depth and remains open below this depth. Within the mineralized granite gold grades are remarkably consistent with an average of 1.1 g/t.

Mineralized granites at Tocantinzinho are divided into two sub-units by alteration mineralogy and colour: smoky and salami granite. Grade distribution is similar in both units and therefore they have been grouped for resource estimation purposes. Both smoky and salami granites are composed dominantly of potassium-feldspar and large (mm- to cm-scale) distinctively amoeba-shaped quartz grains. The textural and compositional similarity in the two granitic host units leads to the interpretation that both represent the same protolith, but have undergone different alteration processes. The blebby, amoeba-like quartz textures and rare miarolitic cavities are interpreted to reflect a late-stage, volatile rich magma.

Smoky mineralized granite is strongly silicified with chlorite and trace to 1% sulfides. Color ranges from grayish to greenish due to the chlorite alteration. Sheeted milimetre-scale chlorite ± sulphide veinlets are common. Salami mineralized granites are distinctively bright red due to hematite dusting on potassium feldspars. Sheeted veinlets are common and similar in scale to those in the smoky granites but are generally filled with quartz ± sulphide. Contacts are diffused between smoky and salami granites and a complete gradation exists between the two units.

Hydrothermal alteration at Tocantinzinho includes a probable early potassic alteration event, a chlorite±sulphide phase, silicifi ation (both pervasive and related to quartz sulfide veining), and a late carbonate phase evidenced by fine calcite veinlets. The chlorite alteration consists of sheeted chloritesilica veinlets.

The mine was designed as a single open pit operation using 3 pit Phases mined over 11 years, including a pre-stripping year, with a peak mining rate of 26 million tonnes of material per year. The open pit is designed to a depth of 345.0 m using an overall slope angle that ranges from 35 - 43 degrees. The mining fleet was selected to excavate a 5.0 m bench using a truck and shovel fleet. The average cost of mining a tonne of material is estimated to be US$1.66/t based on a labour rate study and quotes for diesel, explosives and tires.

The mine will produce 4,400 ktpy of ore. The total ore to be mined is estimated as 49,047 kt with a LOM average grade of 1.25 Au g/t and an average waste: ore ratio of 3.27.

The process plant will have the capacity to process 4.4 million tons of ROM ore per year, with an average of 1.25 g/t of gold and metallurgical recovery of 90.1%, resulting in an average 159,000 ounces of gold per year. The process route will involve crushing, ball mill grinding, classification by hydrocyclones, gravimetry, flotation, intensive leaching and CIP (Carbon in Pulp) circuits.

The ore body has been classified into oxide and sulfide ores. The process of scrubbing the oxide ore was not clearly defined in the testwork. The solution of two distinct circuits for handling the ROM oxide and sulfide ores is based upon the mine plan identifying and delivering the ore to the correct dump hopper. The two distinct ROM handling circuits may result in a higher availability for the crushing and stockpiling circuit as a portion of the feed is sent directly to the ball mill discharge pumpbox.

The oxide ore fines broken down in the rotary scrubber are mixed with the ball mill discharge which is then sent for classification. A fraction of the classification underflow is batch concentrated by a centrifugal gravity concentrator and then sent to intensive leaching in an Acacia type reactor. The Acacia reactor normally has a washing step to remove slimy fines carried over in cyclone underflow. One or two continuous leach reactors could be justified to have continuous supply of pregnant solution for electrowinning out of one or two gravity concentrators (Knelson type).

Both the oxide and sulfide ores are amenable to cyanide leaching with a very good recovery.

The intensive leaching circuit will be a purchased package called an Acacia Dissolution Module which has facilities for leaching and separation of solid and liquid from the leached pulp.

The pregnant solution from the Acacia reactor will be transferred to the electrowinning cell. The barren solution from the electrowinning cell will be recirculated to the Acacia reactor via the barren solution tank.

The gold contained in the pregnant solution is deposited on the steel wool cathode within the electrowinning cell. The cathode are removed periodically and taken to the smelting furnace.

Once the cyclone overflow passes through the trash screen to remove any organic material the slurry is sampled for metallurgical analysis of the flotation feed. The slurry then reports to the flotation conditioning tank, which is equipped with an agitator. The conditioned slurry then flows to the five tank type Rougher flotation cells. The reagents required for this stage of flotation will be distributed into the conditioner tank, and the first two rougher flotation cells.

The Rougher concentrate will be conducted by gravity to the flotation cleaner phase, or alternately to the regrind feed pumpbox.

The concentrate from the cleaner flotation cells will flow by gravity to a pumpbox the concentrate, will be pumped to the regrind cyclone feed pumpbox.

The Cleaner tailings will flow by gravity to a pumpbox; the tailings will then be recirculated to the Rougher flotation cells.

The classification of the regrind slurry will be conducted in a battery of five 10” diameter hydrocyclones with four operating and one standby. Cyclone feed slurry will be pumped from the cyclone feed pumpbox to classification.

The cyclone underflow will flow by gravity to the Regrind ball mill. The regrind ball mill discharge will be equipped with a trommel screen. The regrind ball mill product, with a P80 of 0.1 mm, will flow through the trommel screen. The slurry reports to the regrind cyclone feed pumpbox and the spent grinding media and trash to a tote for disposal. This closes the cycle with a circulating load forecast of 300%.

The overflow of the regrind hydrocyclones with a P80 of 0.06 mm will flow by gravity to a thickener. The thickener overflow will discharge into a pumpbox, from which it will be distributed as process water. The thickener underflow pulp with 50% of solids will be pumped, to the leaching stage of the process.

The leach circuit feed passes through the trash screen with apertures of 0.6 mm to remove any organic materials. The trash screen underflow is sampled for metallurgical analysis and is then directed by gravity to the first, or if bypassed for maintenance, the second leach tank where reagents; sodium cyanide, leaching agent, milk of lime and pH adjuster are added. The leaching cycle will take place in six leach tanks, equipped with the agitators. Each tank will have air injected. The pulp flows by gravity from one tank to another, resulting in a total residence time of 48 hours within the leach circuit.

The leach product is sampled and flows to the CIP feed pumpbox. CIP feed pumps will transfer the slurry to the CIP circuit.

The CIP circuit consists of eight tanks, equipped with agitatorsand DSM type inter-stage screens with apertures of 0.6mm aimed at retaining carbon.

The pulp in the CIP circuit flows by gravity due to the elevation difference between each tank. The adsorption process is countercurrent to the pulp flow, i.e., regenerated carbon feeds the last tank after passing over the vibrating dewatering screen, and will subsequently be pumped by the carbon forwarding pump to the previous tank and so on until it reaches the first tank. The loaded carbon from the first CIP tank will be pumped by the carbon forwarding pump, to the loaded carbon dewatering screen.

The loaded carbon retained by the loaded carbon screen, will report to the loaded carbon tank after sampling. The loaded carbon is then is transported via an eductor to acid washing and regeneration, or directly to elution. From either of these, carbon is transported to the regeneration kiln.

The discharge from the last CIP tank will flow by gravity to the vibrating carbon safety screen. The retained carbon will be taken to the next stage of the process, elution and electrowinning. The carbon safety screen underflow will be sampled and flow by gravity to the CIP tailings thickener. Overflow from the thickener containing recovered sodium cyanide solution will be distributed to the leach and CIP circuits via pumpbox and pump. The thickener underflow will be pumped to the effluent treatment tank.

The effluent treatment tanks are equipped with agitators. Process reagents will be dosed into these tanks which will also be supplied with sparging air. The treated effluent will flow by gravity to pumpbox from which it will be pumped by pumps, one running and one standby, to three retention ponds with 4,700m3 capacity each. The treated effluent will be transferred via pumps to the final tailings pumpbox from which combined with the flotation tailings it will be pumped to the tailings dam.

A solution containing 0.2% sodium hydroxide and 1.0% sodium cyanide is prepared in the electrolyte tank. The solution is pumped bya pair of pumps, one running and one standby, through the heat exchanger where it will be heated to 95°C and taken to the elution column. The heated solution is used to promote desorption of loaded carbon from the CIP process which is fed into the top of the columns after dewatering by the vibrating screen.

After approximately 60 hours of operation of the elution column the electrolyte solution containing gold flows through heat exchanger for cooling and on to the electrolyte distributor. The distributor will split the flow to feed the three electrowinning cells or alternately to the electrolyte tank where it will be recirculated to the elution columns. The desorbed carbon will be transported to acid washing and carbon regeneration via an eductor. The gold from the loaded solution fed into the three electrowinning cells will be deposited on the cathodes made up of steel wool. The cathodes are removed periodically and taken to the smelting furnace. The solution now depleted by electrolysis will be discharged into the electrolyte tank.