Berezitovy Project — 2012 Technical Report

This article describes the processing route evaluated and implemented for the Berezitovy gold project, based on the July 2012 technical report.

Report context

The Berezitovy Project technical report, dated July 2012, documents processing methods for the project in Russia. The report covers flowsheet options studied from testwork, a description of the operating plant constructed between May 2006 and December 2007, process control, consumables, water balance, production data, plant expansion plans, and the assay laboratory. The plant was operating at approximately 1.4Mtpa in September 2011, with planned increases to 1.8Mtpa in 2012 and 2.0Mtpa by 2013.

Processing route

Flowsheet options studied

Based on testwork results, five potential processing options were studied:

• Option 1 – Cyanide leaching followed by selective flotation of zinc and lead from leach tailings.
• Option 2 – Gravity separation and cyanide leaching followed by selective flotation of lead and zinc from leach tails.
• Option 3 – Gravity separation followed by cyanide leaching of gravity tails, with gravity separation producing a coarse gold and a saleable gold-lead middlings concentrate, and zinc flotation of leach tails.
• Option 4 – Gravity recovery of coarse gold, production of a lead-rich gravity middlings processed in a separate cyanide circuit, cyanide leaching of gravity tails, followed by flotation of zinc from leach tailings.
• Option 5 – Gravity recovery of gold, flotation of gravity tails to produce a bulk concentrate subjected to cyanide leaching, with recovery of lead and zinc from leach tails using selective flotation.

It was concluded that the highest gold recovery would be provided by Option 2, at up to 90%. This option also included selective recoveries of lead and zinc of 61.7% and 58.2% respectively, which could be processed by a pyrometallurgical plant.

Options 3 and 4 gave acceptable gold recoveries of between 87.0% and 87.3% but were rejected due to the potentially complicated flowsheet and because gold recoveries to doré product were low.

Option 1 was eventually preferred over Option 2 due to the simpler process sheet and the elimination of gravity gold concentrate refining, although Option 1 gave the lower gold recovery of 86.5%. Testing at Lakefield indicated very little increase in recovery using gravity ahead of cyanidation.

Plant construction and expansion history

The plant was constructed between May 2006 and December 2007. Many major equipment items, including the jaw crusher, the SAG mill, and the original ball mill, were sourced second hand, which impacted significantly on operating efficiency.

In October 2009, the plant was treating ore at a rate of 1.1Mtpa. Tailings filtration capacity was being increased and an additional ball mill installed, planned to enable the processing rate to increase to 1.5Mtpa in 2010 and 2.0Mtpa in 2011.

In September 2011, two new Russian disc filters and a new ball mill were fully operational, and the plant was treating ore at approximately 1.4Mtpa.

Ore stockpiles

Ore is removed from the pit and delivered to one of two stockpiles: an on-balance/medium grade ore stockpile containing 1,398,020t at an average grade of 1.71g/t Au, and an off-balance low-grade ore stockpile containing 2,519,140t at an average grade of 0.74g/t Au.

Ore is delivered to the main plant feed hopper using a front end loader. The hopper is fitted with a 1.1 x 0.8m grizzly, and oversize material is broken using a hydraulic hammer drill. Ore is extracted from the hopper using an apron feeder and transferred to a jaw crusher, which reduces material to less than 200mm. The crushing rate is variable depending on the ore being treated but typically averages 250tph. The crushing units are shut down for maintenance work 6 days a month.

The primary crushed ore is conveyed to a crushed ore stockpile containing approximately 49,800t of ore at an average grade of 1.98g/t Au. The stockpile suffers from segregation of coarser material, which tends to gravitate to the bottom. Ore is reclaimed from the crushed ore stockpile using three feeders, then conveyed to the grinding circuit.

Grinding

Ore is ground in a SAG mill–ball mill circuit to a size of 80% passing 74µm. The SAG mill is 6.7m x 2.4m, fitted with a fixed speed 1,800kW motor. Approximately 250tph of material is fed to the SAG mill, where it is ground at 75% solids. The SAG mill was originally fitted with a 12.5mm discharge grate, which restricted throughput; the grate size has been increased to 18mm.

A major problem of the SAG mill is the production of approximately 20t of critical size material every day. Previously, this material was recycled back to the crushed ore stockpile. In late 2011, a KSD-600 crushing unit was installed within the grinding circuit to process this critical size material.

The 8.5 x 4.4m ball mill is powered by a 2,240kW motor. Previously, problems with the ball mill meant the optimum ball loading could not be reached and the target grind size was not achieved, resulting in a reported recovery loss of 2%. Numerous repairs and a recommendation from Irgredmet to implement a 23% ball loading rectified this problem.

The milling rates of the original circuit ranged from 185 to 220tph. During the second half of 2010, a new 4.5 x 6m ball mill was installed in parallel with the existing circuit. In September 2011, it had increased milling rates to an average of 250tph. The modifications made to the grinding circuit appear to have resolved the issue of under grinding, with 80-82% now passing 74µm.

The mills are out of operation for an average of 6 days a month for maintenance work, accounting for a plant utilisation factor of 0.81.

The addition of cyanide to the grinding circuit causes problems with sampling, as the mill water contains solubilised gold, necessitating analysis of both solid and liquid phases to obtain an accurate gold assay.

The cyclone overflows are screened using a combination of Delkor screens to remove wood chips and vegetation.

In October 2009, a small centrifugal gold recovery unit (0.8m diameter) was installed within the grinding circuit for trial testing. A gravity table was also tested around the same period. Both pieces of equipment proved uneconomical and were subsequently removed from the circuit.

Leaching

The ground pulp, at roughly 24% solids, passes to a 24m diameter thickener. The thickener overflow is returned to the grinding circuit, and the underflow, at 52-53% solids, is pumped to the leach tanks.

There are eight leach tanks, each with a capacity of 430m³. The tanks are air sparged, but the levels of dissolved oxygen are not determined. Cyanide is added to tanks 1 and 5, and the pH is maintained at 11.5.

Carbon adsorption takes place in six 330m³ adsorption tanks. Previously, 28t of Malaysian carbon was used in the circuit. This has recently been substituted for new Thai carbon, the optimised loading of which into the column is still under review. An initial loading of 84t yielded unsatisfactory tailings grades. Approximately once a week, the carbon loading is reduced by between 5-6t to measure the effect on the grade of the tailings. If still deemed unsatisfactory, the carbon loading is reduced by a further 5-6t until an optimised loading level is reached. In September 2011, there were 54t of Thai carbon in the adsorption tanks.

The loaded carbon, typically assaying 1.5kg/t Au, is screened and passes to desorption. Carbon reactivation is achieved partially in a kiln and partially via an acid-wash process.

Desorption and electrowinning

Gold is recovered from the carbon using a conventional desorption-electrowinning and smelting process.

There are three desorption circuits: one capable of processing 6m³ of carbon, while the other two are capable of processing 4m³ of carbon each. The units are of Chinese manufacture and are capable of stripping carbon down to 120ppm Au.

Electrowinning takes place in four 4m³ electrowinning cells, with one back-up cell should maintenance be required on one of the four operating cells. The cathode sludge, which assayed 177kg/t Au, is smelted on site to produce doré.

The smelter is operated approximately twice weekly, dependent on the rate of material passing through the plant. There was originally only one furnace, but the delivery of a new Chinese furnace and the manufacture of another furnace on-site, both in 2010, increased the complement to three.

The smelter is involved in the monthly production of approximately 700kg of doré, typically assaying 32-34% Au and 32-34% Ag, as well as quantities of Zn, Pb, and Cu.

Filtration of leach tailings

The Berezitovy plant uses dry deposition for tailings disposal. There are currently two dry tailings dams, with a combined capacity of 3.3Mm³. The leached tailings are pumped to a filter station which originally consisted of three disc filters of Chinese manufacture, each with a 200m² filtration area. The filters proved unreliable due to poor construction materials and mechanical failure, and it was reported that nearly all working parts had to be replaced. The old Chinese filter cloth has recently been replaced with better quality Canadian filter cloth, allowing each filter to run at a rate of 60-80tph.

In October 2009, two additional disc filters of Russian manufacture were in the process of being installed. These are now fully operational, each with a 100m² filtration area and capable of processing 40-60tph. It was reported that Canadian filter cloth was due to replace the current filter cloth in the two Russian disc filters.

In the second half of 2010, modifications and repairs within the filtration plant allowed for increased capacity of tailings filtration to cope with the current processing rate. However, even with the modifications made, the filtration station would be incapable of dealing with all plant tailings if the processing rate was increased to 2.0Mtpa as planned. Plans to modify existing, or install additional, filters to deal with increased production rates were still in preliminary stages.

The initial problems of filtering the leach tailings resulted in large quantities of leach tails being sent directly to the wet tailings dam without prior filtering. In 2008, it was reported that some 60% of plant tailings was sent to this dam, although this figure fell to 30% in 2009. The modified filtration station is operating such that no tailings have had to be sent direct to the wet tailings dam in 2010 or 2011. Any pulp that was pumped into the wet tailings dam was directed back to the filtration station once modifications had been completed.

If the level of the wet tailings dam was to exceed its capacity of 500,000m³, the excess would be diverted to an emergency tailings pond. This has not occurred at any point during the operating life of the plant.

Process control

The plant is controlled using a SCADA system supplied by Festo (German manufacture). The system monitors the SAG mill feed rate, together with cyclone feed and leach tank 1 pulp densities. The levels of pH and cyanide in the leach circuit are also monitored and controlled.

Plant sampling

The Technical Control Department (OTK) is responsible for sampling of the process plant, undertaken according to detailed procedures stating the sampling methodology and sampling interval for each processing stream. As well as chemical analysis, process streams may be tested for pH, particle size distribution (PSD), cyanide concentration, and specific gravity.

Key sampling points include:

• Run of mine ore sampled manually from conveyor 2 (the conveyor transporting crushed ore from the stockpile to the SAG mill) once every hour.
• A weightometer on the SAG mill feed, calibrated once per month by the instrumentation department using known weights and chains, and once yearly by the Amur Centre for Standards Accreditation.
• Sampling of the plant feed is complicated by partial dissolution of gold in the grinding process; the sample is filtered and both solids and liquid phases are assayed for gold.
• Automatic samplers on the first leach tank and on the leach tailings pulp as it enters the filtration plant, with cuts made every 20 minutes.
• Feed to the first adsorption column sampled every 2 hours.
• Stripped carbon sampled after each desorption cycle, approximately every 12 hours; pregnant solution passing to the electrowinning circuit sampled every hour; flow measurements around the desorption circuit made hourly.

Consumables

The 2011 data for major plant consumables are given in Table 17.1.

The consumption of all items is up on previous years, most notably grinding media at 3,400g/t compared with 2,030g/t in 2009. This is attributed to the installation of a second ball mill in parallel with the original, accounting for an additional 1.2kg of grinding balls per tonne of ore processed.

Carbon consumption is also higher than in 2009, up to 150g/t from 118g/t, due to the higher quantity of carbon entering the adsorption tanks.

The grinding media is sourced from Russia and the sodium cyanide from Korea. The carbon was previously obtained from a Malaysian supplier but is now sourced from Thailand. The rates of consumables are typical for a moderately hard, non-refractory ore.

Water balance

A water and mass balance for the processing plant was detailed for the milling and thickening

Key reported parameters

Technical qualifications

The report describes process design and water-balance information; unreported operating values are not inferred.