Garrcon Project — 2014 Technical Report

This technical report reproduces processing sections from a 2011 preliminary economic assessment, which has not been updated to show the 2012 Garrcon mineral resource update.

Report context

The recovery methods section (Section 17) of the March 3, 2014 technical report (A.C.A. Howe International Limited Report No. 975) has been extracted from Howe's 2011 NI 43-101 technical report and PEA (Hannon et al., 2011). The reader is cautioned that the 2011 Garrcon PEA is based on Howe's 2011 Garrcon mineral resource estimate, which has now been replaced by the 2012 Garrcon mineral resource update presented in this 2014 report. The 2011 Garrcon PEA reported in the following subsections has not been updated to show the changes in the Garrcon mineral resource.

Processing route

Testwork basis

Available information on test work is limited to two small samples. There are no operating results available relating to the recoverability of the gold at the Garrcon deposit. The SGS work allows for a standard gold processing method to be tested.

Proposed processing philosophy

The operating philosophy is to get the gold out of the rock as soon in the process as possible. To accomplish this, a standard gravity-cyanide circuit is proposed for rock with greater than 0.3 g/tonne gold. Rock with a gold content between 0.15 g/tonne and 0.3 g/tonne would be sent to a heap leach facility.

Proposed concentrator design

The proposed concentrator for this study is based on an annual mineable resource throughput of 4.1 Mt, or 11,300 tpd at a 93% plant availability, for the production of a gold product (gravity product, float product or gold bars, depending on the options chosen). The processing plant will operate 24 hours/day, 365 days/year. No assumptions are made on the grade of the material as this has an insignificant impact on capital and operating costs within the range of values analyzed. The tests done to date suggest that this could be on the order of 0.9 g/tonne. The company plans (subject to column test confirmation) to process all material between a grade of 0.15 g/T and 0.3 g/T in the heap leach system and material with a grade greater than or equal to 0.3 g/T will be processed in the gravity/vat leaching process.

In this circuit, it is assumed that sulphide separation requires a finer grind than 2000 micrometers and that essentially half of the crusher product will be ground. It also assumes that the sulphide separation is performed at a size of approximately 75 micrometers. A gravity separation circuit is included, as spirals; however, this may have to be replaced by a flotation circuit. In addition, no waste stream is produced at this point. The sulphide stream is assumed to represent 12% of the initial flow, or 1,360 tpd. This stream is reground and vat leached. The gravity, or flotation, underflow/tailings, is then vat leached.

The processing scenario presented in this report includes various run-of-mine stockpiles, crushing, fine and coarse ore stockpiles, heap leaching, the loading and unloading of gold onto the carbon, carbon regeneration, cyanide recycling and destruction, electrowinning and gold bar production. This circuit is shown as a block diagram in Figure 17-1. All processes concerning the tailings have not been included in the processing analysis.

Crushing circuit

The crushing section assumes a feed material similar to the blast fragmentation size analysis, provided by Northern Gold. The following assumptions are made in the configuration and costing of this circuit: 12-inch top rock size; 100 metres of conveyors required to feed the crushing plant for the run-of-mine stockpiles; and the suitability of vertical shaft impact crushers.

Stockpile costs are taken as $100,000. This cost is ball-park only as it depends on the location and the size that will be required; Howe is estimating a 2-month supply for the mill, amounting to about 680,000 tonnes and requiring an area of about 5 ha. The coarse ore storage will consist of run-of-mine stockpiles that will be established and maintained in three categories. The main stockpile will be feed for the mill, assumed to have an average grade of 0.9 g/tonne. The second stockpile will be a run-of-mine stockpile for heap leach operations and will consist of rock grading between 0.15 and 0.30 g/tonne. The heap leach stockpile will be available to supplement mill feed if required. The third stockpile will be low grade stockpile whose grade is as yet unknown.

VSI costs were substituted by Single rotor impact crusher at 400 t/hr, or a 32-inch by 44-inch model. Furthermore, it is assumed as there are two stages of this crusher so that two will be required. The use of VSI must be further investigated and compared to additional stages alternative crushers in terms of effectiveness, capital cost and operating costs. It is assumed that 11,300 tpd of 0.3 g/T, or greater, ore will be processed through the mill. The 8,000 tpd (9 months/year) of ore, ranging between 0.15 and 0.3 g/T, will be processed through the heap leach.

Gravity-cyanide circuit design basis

The process circuit includes the following assumptions: liberation of crystals within the rock can be achieved at a size of 75 micrometers; sulphides, the gold bearing silica and carbonates can be separated using gravity separators; gold associated with the sulphides is exposed either on surface or in cracks, at a size of 75 micrometers, thus can be leached; and the proper percent solids for the leach must be decreased from the hydrocyclone overflow by using thickeners.

The consumable costs are estimates only as the pumps have not been priced or sized, the exact mill size or grinding media consumption rate has not been determined and the chemical consumption rates have not been determined. The consumables include steel media, activated carbon and cyanide.

Basis of design assumptions include: that a grind size can be determined that will both liberate the pyrite and expose the gold sufficiently for cyanidation (a two-stage grind may be considered if these two sizes are considerably different, and this size can be determined using representative mineralogical grain size studies and cyanidation tests); a pyrite separation stage is to be performed if gold is found within the pyrite as small particles capable of being leached on exposure to cyanide solution as opposed to being encapsulated within the sulphide crystal lattice itself; that the size of pyrite liberation is appropriate for gravitational separation techniques (it may be necessary to use flotation as opposed to gravitational separation, and the selection of the type of gravity separation is also dependent upon this size); that the regrind of the pyrite be performed to reduce the pyrite to a size where the particles of gold are exposed; and the silica/carbonate stream does not need to be reground and sufficient gold recovery is achieved at 120 micrometers.

Grinding

This circuit is used to reduce the particle size from the crushed ore to that required for cyanidation or flotation. The grinding is assumed to be performed using ball mills in closed circuit. The resulting particle size is assumed to be on the order of 120 micrometers. The consumption of grinding media, or steel balls, has been estimated at approximately 1.5 kg per tonne of ore, or 17 tonnes per day. This will change depending on the grind size chosen and ore specific characteristics. This does not include any of the capital or consumption of potential regrind circuits.

Regrind

A regrind circuit reduces the particle size of the gravity circuit concentrate (pyrite) to expose the gold within the pyrite. It has been assumed that this circuit will process 12% of the ore; the costs of this regrind will be that of a hydrocyclone acting as a thickener and closing the circuit and one ball mill. The costs of alternatives (Knelson/Mozley, jigs or flotation) have not been approximated at this time.

Sulphide concentration

This circuit assumes that the pyrite in the rock will be removed and treated separately from the silicates and carbonates and that the largest grind size possible, to liberate the pyrite, is used in the grinding circuit. The type of separation used to remove the pyrite depends on the particle size. For larger sizes, gravity separation can be used. Two scenarios have been analyzed: (i) spirals and (ii) Knelson and Mozley type separators. Jigs should be considered for larger particle sizes and flotation for smaller sizes.

In scenario (i) the comminution circuit product is pumped at 25% solids (w/w) to a number of spirals, probably a series of three, that produce a heavy and light component. It is assumed that the heavy component will be composed of 12% of the feed. It is unlikely that this circuit will achieve the recovery of the test work, so the recovery has been set at approximately 30% of the gold. It is assumed that the spirals have a capacity of 8 tonnes per hour; however, this is just an estimate that depends on the type of spiral, the percent solids, the size of the gold and other particles and other factors. At 11,300 tonnes per day, 470 tonnes need to be produced per hour, or about 60 spirals in parallel. Assuming a 50% mass recovery in each, 30 will be needed in the second stage and 15 in the third for a total of 105 spirals. Two pumps and an appropriate splitter could feed the first stage, and one pump and splitter feed the second and third stages.

Silicate/carbonate leach

In this circuit the feed is assumed to be taken directly from either the grinding circuit cyclone overflow or the gravity circuit underflow to a thickener where the percent solids is adjusted from about 25% to 45%. Depending on the particle size, it may be possible to use screens or cyclones to replace the thickener. The thickener has not been designed and its eventual size will depend on the grind size and rheology of the resulting slurry. Generally, this is assumed to be in the range of 0.3 to 1.3 m² per tonne of ore. The feed thickener is only changing the percent solids from 25% to 45% thus will be on the smaller side of these values. A value of 0.3 m²/tonne, or 3000 m², is used in this estimate. In the 11,300 tonne per day scenario this results in a thickener 62 metres (200 feet) in diameter. This is probably an over estimate of the size required.

The tanks themselves are based on a requirement of a 24-hour residence time, based on 8 tanks in series, in two parallel lines. In the 11,300 tonne per day scenario a volume of approximately 1000 m³ will be required per tank. Air lift will be required on each tank; this can also be done using oxygen sparging to increase the cyanidation rate. The solids then flow to a series of 16 leaching tanks, two parallel lines of 8, to give a total residence time of about 24 hours. Double the number of tanks is assumed for the 48-hour retention time scenario.

Cyanidation consumables

The consumables have been estimated for only the whole rock cyanidation scenario. The cyanide consumption may change depending on the iron content. All other consumables are anticipated to remain approximately the same. Lime has been included in this section despite the fact that it is added in the grinding circuit as it is required when cyanidation is performed. The consumables for the refinery have also been included in this section because without cyanidation refining would probably not be performed on site. No cost estimates have been made on transportation or inventory costs. No cost has been included for kiln fuel.

Cyanide consumption is based on a consumption of 0.9 kg/tonne. Heap cyanidation consumables are included in the heap section.

Heap leaching

It may be possible to use a heap leach to process the rock containing insufficient gold for the conventional leaching processes. The advantage of the heap leach is usually cost; however, this would come with a reduced recovery. The recovery is assumed to be 65% in this costing (a number supplied by the company and not proven by metallurgical testing). This scenario assumes the leached rock will contain between 0.15 and 0.30 g/T gold. The rock would be removed from the crushing circuit between the tertiary and quaternary crushing stages. This circuit assumes that the fragmentation from intense blasting along with crushing will develop the necessary cracks to enable cyanide solutions to contact the gold (the optimal size for leaching has yet to be determined). The cyanide consumption in the heap has been assumed to be 0.35 kg/tonne of ore. Lime has been included in this section; however the amount of lime that will be required is not yet known as there are known carbonates in the rock itself.

Key reported parameters

Technical qualifications

The report presents flowsheet assumptions and design information for the Garrcon Project; unreported results are not inferred.