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Development of the Greenvale orebody, containing 40,000,000 tonnes of nickel ore, began in 1971 and production commenced in 1974. Annual capacity is 2,800,000 wet tonnes. The variations in ore grade and physical characteristics which occur throughout the ore- body had a major influence on the selection of mining methods, grade control procedures, and equipment. After removal of overburden by scrapers, ore is excavated by draglines sidecasting into windrows. Feed to the crushing plant is .reclaimed by front end loaders into rear dump trucks according to a grade selection pro- gramme. After crushing, ore is stockpiled into blended units of ore of required quality. Ore is stacked and reclaimed for train loading using techniques designed to give maximum mixing of ore characteristics through Bach unit. These procedures are repeated on discharge at the plant site at Yabulu. The railway, constructed by Queensland Nickel Pty. Ltd. to Q.G.R. standard, and operated by the Q.G.R., is 216 km long and includes eight major bridges and three tunnels. Two trains per day, one of four locomotives and 92 waggons, and one of three locomotives and 69 waggons, haul crushed ore to Yabulu. Train loading is semi-automatic. A small township and associated services have been constructed for employees at Green- Vale.

Jan 1, 1974

By A Atkins

Six recent case studies were selected based on the eight main causes of mine disasters (Table 1; Quinlan, 2014) and where credible information sources, such as Royal Commission Reports, Coronial Investigations and textbooks, were available. The six case studies are listed below. Details are available via this link or the conference PowerPoint accompanying this paper. 1. Beaconsfield, Tasmania – 2006 Fall of Ground/Seismic Event (Chandler, 2009; Melick, 2007; Wright, 2012). 2. Bingham Canyon, USA – 2013 Manefay Slide (Ross, 2017). 3. Bronzewing, WA – 2000 Hydraulic Fill Inrush (Hope, 2002). 4. Brumadinho and Samarco, Brazil – 2015 and 2019 Tailings Storage Facility (TSF) failures (Bowker, 2019; Darlington et al, 2019; Demajorovic, Lopes and Santiago, 2019; Morgenstern, 2016). 5. Northparkes, NSW – 1999 Cave Air-blast (Bailey, 2003; Hebblewhite, 2003). 6. Pike River, NZ – 2010 Methane Gas Explosion (Macfie, 2013; Panckhurst, Bell and Henry, 2012.

May 5, 2022

By D Kentwell

For deposits where a full assay suite of elements/oxides is available and where several variables are critical to the evaluation of the deposit, it is desirable for the estimated block values of all elements/oxides to sum to 100 per cent of the weight per cent. This is difficult to achieve in practice and the estimation methods used need to be designed to minimise bias of the key elements/oxides by ensuring that block estimates are consistent with the individual assays and that the estimated totals are as close as possible to 100 per cent.This paper presents a case study of two domains within a magnetite hematite deposit, one with a high level of drilling classified as Measured and one with a lower density of drilling classified as Indicated. Several estimation methods were tested to evaluate their ability to maintain consistency of the proportions of each element/oxide and sum to 100 per cent. The following methods were evaluated: ordinary kriging, co-kriging, inverse distance and min/max autocorrelation factors (MAF) all as implemented in the Geovariances Isatis software.The comparisons showed that the MAF method produced the most robust estimates in terms of block totals and that grades and tonnages at the reporting cut-off were not significantly different between methods. At an individual block scale all methods produce artefacts that are less than desirable but can only be remedied by additional drilling.CITATION:Kentwell, D, 2015. Comparison of min/max autocorrelation factors, kriging, co-kriging and inverse distance block totals – a magnetite hematite case study, in Proceedings Iron Ore 2015, pp 465–474 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 13, 2015

Thank you for the opportunity to speak at this important conference attended by aspiring mining professionals, some of whom will become the leaders of the mineral resources industry that forms the solid foundation of AustraliaÆs past and that will fuel the exciting growth of AustraliaÆs future.I want to briefly touch on the history of our business, then look at the opportunities that exist in the contract mining industry and finally comment on what the educators need to consider in their curriculum to meet the needs of our business. Byrnecut started out as an idea in early-1987 developed by a group of four mining people. At the time the industry was on the high end of a mining boom, the era of the Bonds, Connels, Skases and so on. Riding on this wave of euphoria we borrowed $600 000 from the National Australia Bank and, without having enough knowledge to lose sleep over the potential risks, started business in mid-1987. The euphoria quickly evaporated with the 1987 October stock market crash. The bank pulled our overdraft, equipment finance rates increased to over 20 per cent per annum and my house was mortgagedseveral times over. The business learning curve was extreme for a young mining engineer who had developed his trade in the relatively sheltered environment of a large public mining house, namely WMC. We were somewhat bitter that the banks had done a runner. We were a private company, backing the business with personal assets with our shoulders to the wheel, whilst the Bonds of the world had theirnoses in the trough ripping off mum and dad shareholders. All of this had a sobering effect, namely keep your feet on the ground, stick to the knitting, work hard and worship cash flow. Since that stressful but adrenaline pumping beginning the company now employs over 600 people and is now working for major mining houses at over a dozen locations in Australia, east Africa and WesternEurope and the National bank is again our banker. On this basis I consider myself qualified for the subject of this speech.

Jan 1, 2002

The objective of a mine planning exercise is to plan a safe mine that maximises net present value (NPV) for the project. Two primary elements are key to design of a mine plan optimised to maximise profit: selection of an optimised production rate and selection of an optimised equipment fleet. Both depend upon accurate estimations of capital and operating costs for the equipment and for the life cycle of the mining operation. Once a trial production rate is determined, a primary equipment fleet, ie excavators, haulers, and blasthole drills, can be determined based on material characteristics and haulage distances and gradients. An auxiliary fleet, dozers, road maintainers, service trucks, etc can be determined using less rigid productivity criteria. With an equipment fleet determined, operating and support personnel can be assigned based on shift requirements and equipment operating needs; and material and supply requirements can be estimated. Knowing the areal dimensions of the major equipment items, road widths and repair shop dimensions can be determined. Personnel numbers and supply requirements determine the square metre sizes of offices, change rooms, and storage facilities. Reliable sources of capital and operating unit prices and costs for all these items are referenced in the paper. Once preliminary capital and operating cost estimates are calculated by applying unit costs to equipment, personnel, material and supply needs, potential profit can be tested using DCF techniques. An optimised mine plan can then be determined by testing the potential profitability of multiple mine production rates. Because of the economies of scale, full capital and operating cost estimates must be completed for each production rate in order to produce meaningful comparisons.

Dec 6, 2010

By M Russ, J Gediga

This paper shows an economic feasible approach to implement greenhouse gas (GHG) reduction measures into the mining and metallurgy industry (MMI). The goal to improve energy consumption is directly linked to the reduction of GHG emissions in the MMI. A baseline scenario of the considered reference system and of the respective reference year has to be designed and analysed. In a second step an analysis of the same operation using the best available technology (BAT) processes is carried out to generate a basis for a benchmark system. The identified gaps will be reported and the potential GHG emission reductions will be put into relation to the investment, running cost and generated carbon credits of the new process technologies to be implemented. This economic feasibility calculation is necessary to realise a cost efficient GHG reduction roadmap decision-making for implementing new technology into the companyÆs business operations. The GHG reduction roadmap is developed using the marginal abatement curve approach.

Jan 1, 2009

Ground conditions at the Kencana mine, Indonesia, are generally poor and an underhand cut and fill mining method is used for extraction. Due to these poor ground conditions, all vertical development has been conducted using blind, top-down, sink and line excavation methods. Due to time constraints, raiseboring was considered the only viable excavation method for the installation of a return ventilation system for the K2 (Kencana extension) project. In order to achieve successful raiseboring in the poor ground conditions a number of strategies were used including: a detailed geotechnical study for optimal position of the shaft/s, splitting shaft excavation vertically, raiseboring of multiple smaller diameter shafts, extensive pregrouting, installation of drainage holes, and rapid installation of support following excavation.

Nov 25, 2010

Mineral exploration and development can have a significant impact on the environment and local communities, in some circumstances extending well beyond the mine site. Poor environmental performance and consequent social impacts by some companies have affected the reputation of the whole industry to the extent that the principal public concern about mining is its impact on the environment. This is consistent with public attitudes about other major natural resource sectors such as forestry and commercial fishing. Traditionally the community has relied on governments to set the benchmark for acceptable environmental and social performance through establishing and revising regulations governing the mining industry. This however assumes that regulations keep pace with community expectations. It also does not take into account those circumstances where governments can have conflicting roles as both watchdog and project partner. Government regulations and industry self-regulation are not in of themselves sufficient for the minerals sector to demonstrate that it has adequately addressed environmental and social issues associated with the sector. For the minerals sector to improve its reputation it must first improve its environmental and social performance. To do this requires the sector to operate to acceptable standards of environmental and social performance and for the sector to be able to credibly communicate this to its stakeholders such as investors, non-government organisations (NGOs), customers, staff, regulators and local communities. Independent third party certification of environmental and social performance (æcertificationÆ), already operational in forest and fisheries management, is proposed as a mechanism to enable the minerals sector to achieve this. World Wide Fund for Nature (WWF) and industry partners will be collaborating in a project to ascertain the applicability of æcertificationÆ to the minerals sector. A number of economic, social and environmental benefits are identified that could accrue should mining companies be able to differentiate themselves on the basis of environmental and social performance, an attribute evident in æcertificationÆ schemes. External audiences are seeking increasing levels of information and data on environmental and social performance of mining companies, be it local communities, investors, financial institutions or NGOs. Environmental and social reporting should be a key communications device that stakeholders can rely on as a reliable, credible source of information on environmental and social performance. WWFÆs contribution to the debate about environmental and social reporting has been to develop a process to assess the quality of environmental and social reports against reporting ingredients which WWF believe are important if a report is to be an accurate reflection of the reporting entitiesÆ environmental and social performance.

Jan 1, 2002

By D FitzGerald, S Jakica, R Paterson, J Silic

A new 2.5D airborne electromagnetic (AEM) algorithm has been developed by Intrepid Geophysics and Jovan Silic & Associates. The advantage of 2.5D (2D geology, 3D source) AEM inversion in 3D geological mapping applications and identification of conductive drilling targets, compared to the more commonly used conductivity-depth imaging (CDI) transforms or 1D inversions, are demonstrated using examples from banded iron formations (BIFs) and other geological settings. We show emerging AEM systems are capable of providing estimates of economic rock unit thicknesses, dips and fault definition at an accuracy that mitigates the need for pattern drilling.The 2.5D inversion application used in this work and described in Silic et al (2015) is a substantially changed version of ArjunAir (Wilson, Raiche and Sugeng, 2006), a product of CSIRO/AMIRA project P223F. The changes include a new forward model algorithm and a new inversion solver. The application enables the accurate simulation of 3D source excitation for full domain models inclusive of topography, non-conforming boundaries and very high resistivity contrasts. Solution is accurate for a geoelectrical cross-section, which is relatively constant along a strike length that exceeds the AEM system footprint.The major innovation includes a new inversion solver with adaptive regularisation, which allows the incorporation of a misfit to the reference model and the model smoothness function.Memory usage has been dramatically reduced and is estimated prior to execution. For speed the software has been parallelised using Intel message passing interface (MPI) and can be used on standard computing hardware or computing clusters. Data from survey lines with lengths exceeding 100 km can be inverted on high end laptop computers.We allow flexibility in the selection of components and in the estimation of noise.AEM inversion examples are shown for surveys from BIFs, base metals (volcanogenic massive sulfides) and geological mapping projects and the results are compared with known geology and drilling. We demonstrate the much improved mapping and target definition delivered by this inversion method when compared with the other more common transforms or inversion methods used on these projects.CITATION:Paterson, R, Silic, J, FitzGerald, D and Jakica, S, 2017. High accuracy .D airborne electromagnetic inversion method for banded iron formation and other geological settings, in Proceedings Iron Ore 2017, pp 319–330 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Jul 24, 2017

By B Godel, E R. Ramanaidou, M Verrall

Fortescue Metals Group Ltd Iron Bridge (IB) operation is a magnetite deposit located in the Pilbara region of Western Australia situated around 145 km south of Port Hedland. This Mesoarchean deposit belongs to the 3230 Ma Pincunah banded-iron formations (BIF), a member of the Kangaroo Caves Formation, Sulfur Springs Group (van Kranendonk, 2000).

Sep 18, 2023

By A E. Annels, I P. Hodkinson an, S P. Barr

Narrow vein systems represent an important source of gold, which have generally been exploited by small to medium sized companies. Their evaluation poses particular problems because of their grade and geological complexities. The determination of grade is arguably the most important activity in resource estimation, as it is generally grade and not tonnage that makes an orebody. This is particularly true of narrow vein-gold deposits whose resources are generally relatively small, though medium to high-grade (6 - 40+ g/t). They often display extreme grade anisotropy in all directions. Two grade populations are generally observed; 1) low, often subeconomic, background grades and 2) high-grades associated with discrete ore shoots. It is possible for gold to be distributed within the wallrocks as well as within the vein. Extremely high-grade values typify this style of mineralisation and require special consideration during grade estimation to determine if top cutting is valid or not. Estimation of both width and accumulation in 2D on vertical longitudinal projections or inclined planes have historically been used to evaluate narrow veins. Resource estimates based on traditional 2D methods is still widespread, but has in some cases been replaced by computer-based 3D modelling. Grade estimation methods are well-known, ranging from classical to geostatistical. Historically, classical methods such as inverse distance weighting have been used, though geostatistical indicator kriging methods are sometimes applicable. Geostatistical studies are driven by the production of reliable variograms, which in narrow gold veins is often problematic. It is generally difficult to define resources from drillhole intersections beyond the inferred and indicated categories with any confidence. In the majority of cases underground development is required to define measured resources and ore reserves. Trial mining/bulk sampling is often used to test estimated grades and also provides important data on ground conditions and the parameters relating to the proposed mining method required for ore reserve definition. Regular reconciliation should be undertaken between the estimated stope grades, grades indicated from stope/truck sampling and grades realised in the mill in order to determine whether correct evaluation strategies are being applied. The important feature of narrow vein-gold deposits is their strong geological controls; in response to this, geologists must understand the geology as best as possible.

Jan 1, 1999

MR JOHN S Foxall wrote: Mr Clayton in his paper on the above subject (Proceedings, N.S. No. 34) made an assertion which should not be allowed to remain unchallenged. On the top of page 10 appears the following statement;- "By viewing the box at adepression angle of 45°, all the dimensions are disclosed and the box may be projected on isometric co-ordinates" and lower down on the same page:- "An isometric plan, being virtually a projection on a plane inclined at 45°, the nearer the dip approaches this figure so is the outline of the stoping more truly shown." Those statements embody an inaccuracy which would be misleading were an isometric plan to be constructed to suit a special case, ie the case of a mine whose lode dipped at 45°, or approximately 45° and it would be found that in such a plan the outline of the stoping would not be truly shown. The reason for this is that the depression angle at which an object is viewed in order to...

Jan 1, 1920

Underground excavations are subjected to complex stress changes caused by nearby mining, but may also be affected by repeated dynamic loading from blasting and seismic events. In some cases, the combined, cumulative effects can lead to unforeseen damage, displacement and instability. Simulating dynamic loading of an underground excavation is a very complex task. The state of rock mass stress, damage and deformation must first be simulated sufficiently requiring small-scale structures to be explicitly modelled. Supporting elements must be represented with high similitude and then the dynamic event must be applied to the system in the case of blasts, or evolve as a consequence of rock mass instability for seismic events. A special case is for the effect of pit blasting on underground excavations - discontinuous rock phenomena at scales from mine, to bench, to pillar, to tunnel need to be captured and this represents one of the most computationally intensive tasks in mine simulation. Recent advances in computer power allow these tasks to be completed at a mine scale, incorporating discontinuities and geological detail matching the best practice for static Finite Element analyses. The results are a worthwhile contribution to mine decision-making, but require careful interpretation and thoughtful inputs - even more so than quasi-static analysis - as the cumulative effects of unfavourable modeling practices are greater. Some learnings from experiences with mine scale, discontinuum dynamic simulation are summarised, including a discussion of considerations of scale and similitude for this type of analysis.

Nov 25, 2010

In most mining operations the ore undergoes several processes such as drilling, blasting, loading, hauling, crushing, grinding and liberation to become the final salable product. Drilling and blasting is an important step in this process chain and itÆs results such as fragmentation, muckpile shape and looseness, dilution, damage and rock softening effect the efficiency of downstream processes. The value created per ton of broken ore is the difference between the price it commands when sold as the final product and the cost to produce it. Traditionally, the total process in the mining industry is classified into two groups as mining and milling. These are managed as separate cost centres inspite of the interdependency. Each process has a budget and production target and emphasis is usually on maximising production (tons) and minimising cost rather than the overall profitability of the whole business unit. The efficiency of each process is considered to be satisfactory as long as they are within budget and meet the production targets. The mine and mill managers usually try to optimise each process independently rather than the entire process. This paper discusses the potential pitfalls of decreasing the drilling and blasting cost per ton of broken rock without considering its impact on downstream processes. It introduces a holistic approach to blast optimisation by identifying and measuring the leverage that blast results have on different downstream processes and then optimising the blast design to achieve the results that maximise the overall profitability rather than just minimising the drilling and blasting costs. This paper demonstrates the benefits of such a holistic approach to blasting based on computer model simulations and field studies from metal and open cut coal mining.

Jan 1, 2001

By K Katov, S T. Ciutina, J D. Pyke

The development and optimisation of mill discharge trommels, and specifically their screen panels, can be a neglected area in the complex operation of a minerals processing plant. Trommels are often seen as a ‘set and forget’ component of the circuit, even when significant improvements in design are possible. A redesigned trommel and panel arrangement has simplified and added flexibility to the plant shutdown schedule at Kirkland Lake Gold’s Fosterville Gold Mine in Central Victoria.Prior to the trommel redesign, the shutdown frequency was determined by the expected life of trommel panels at 14 weeks. Planning downtime around this item led to higher maintenance costs and wastage in other areas of the plant. This was particularly disruptive with respect to mill lining change-outs, which were forced out of sequence and often changed prematurely. Fosterville Gold engaged Metso Australia Limited to propose a new trommel design that would better suit the site. The new trommel was engineered to balance simplicity with performance, and incorporated the latest in Modular trommel design, modelling and Finite Elemental Analysis (FEA). Modular Panels and Spiral Segments achieve significantly longer life, and allow in-situ panel change outs, flexibility in spiral and dam positioning and an improvement in classification performance. Trommel panel life has significantly increased from 14 to 71 weeks in the trommel’s highest wearing zone. The new design has eliminated some risks to safety by reducing manual labour and lifting requirements during shutdowns. There is greater flexibility with respect to shutdown requirements. The site is no longer at the mercy of trommel panel lifespan, which allows for better optimisation of wear components in the circuit. CITATION:Pyke, J D, Katov, K and Ciutina, S T, 2018. Modular SAG mill discharge trommel for maintenance flexibility and performance, in Proceedings 14th AusIMM Mill Operators' Conference 2018, pp 613–620 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Aug 29, 2018

By R Lastra, Y Choi, C Hamilton, P Kondos

Hydroseparation (http://www.cnt-mc.com) was used to concentrate sized fractions (from <38 ¦m to 150 ¦m) of various ore products ahead of carbon-in-pulp (CIP)-treatment. The resulting size-by-size concentrates were mounted as monolayer polished sections and subjected to automated unattended precious metal searches by image analysis (see Lastra, Wilson and Cabri, 1999). Systematic variations in Au-deportment permit quantification of Au reporting as discrete grains from fresh-feed through to pressure oxidation (POX)- and CIP-feed. Application of hydroseparation, image analysis and assays to characterising tailings (from <10 ¦m to 106 ¦m) from four mine sites is also described. Results are encouraging enough to predict that the combination of hydroseparation (HS) and automated image analysis could become a standard methodology for study of precious metal ores and metallurgical products.

Jan 1, 2008

By E Holloway

The optimisation of an open pit mine can be broadly split into two stages, the first being the ultimate pit optimisation (and subsequent stage designs), and the second being schedule optimisation. Macroeconomic uncertainty clearly impacts on both however interrelationships do also exist. Generally, the opportunity exists for the ultimate pit to be revised during each planning cycle, however when this reprieve is removed the uncertainty for the underlying business increases. As one example, the introduction of an in-pit waste dump can be seen to significantly limit the option to revise the ultimate pit. When in this situation there are two potential risks that present: (i) the long-term price used for the ultimate pit selection is optimistic, and too large of a pit is selected resulting in waste being mined which uncovers material not of a significant value to deliver a profit; (ii) the long-term price used for the ultimate pit selection is pessimistic, too small of an ultimate pit is selected, and ore of realisable economic value left behind. Both of these outcomes are risks and as such can be assessed and quantified. CITATION: Holloway, E, 2018. Incorporating economic variability into a strategic planning framework, in Proceedings Complex Orebodies Conference 2018, pp 88–90 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 21, 2018

By K A. Heasley

The LaModel program was initially developed in 1993, and it is used to model the stresses and displacements in thin tabular deposits such as coal seams. It uses the displacement-discontinuity variation of the boundary-element method with a simplified laminated overburden model and because of this formulation, it is able to quickly and effectively analyse large areas of single- or multiple-seam mines. Over the last couple of decades, LaModel has been used by many mine engineers and researchers for improving pillar designs and mine safety, and it has been continually updated to add functionality and improve its ease of use and accuracy. The present LaModel can calculate seam displacements and overburden, multiple-seam and total seam stresses, pillar safety factors, energy values, roof-bending stresses, fault effects and surface subsidence. It can automatically generate the mine and overburden grids from AutoCAD drawings, and it includes wizards for automatically generating Bieniawski strength coal pillars, easily calibrating the overburden properties to observed abutment stresses and calibrating the gob material to expected gob loading. This paper will present the formulation and features of the modern LaModel program and demonstrate its utility with a multiple-seam case study from the rugged topography in Southern Appalachia.CITATION:Heasley, K A, 2014. LaModel – a program for coalmine modelling, in Proceedings AusRock 2014: Third Australasian Ground Control in Mining Conference , pp 409–414 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Nov 5, 2014

By Y-B Kang, J H. Park, J Park

In order to understand the effect of oxygen injection into the molten metal on the valuable metals (VM) recovery and interfacial phenomena, several high-temperature experiments were conducted using a high-frequency induction furnace with a quartz tube as the reaction chamber at 1673 K. Copper-based waste printed circuit board (PCB) that contains some impurities was selected as the end-of-life resource. CaO-SiO2 based pre-fused slag and oxygen gas were used in present works. The composition change and microstructure of the samples were analysed. As oxygen was injected into the system, impurities were rapidly removed. After the removal of impurities, continuously injected oxygen lowered interfacial tension between liquid slag and liquid metal. Lowered interfacial tension induced Cu loss into the slag.

Aug 21, 2024

By G DeschOnes, M Fulton

The Eleonore property (James Bay district of Northern QuTbec), owned by Goldcorp, hosts gold within stockworks of quartz-tourmaline-arsenopyrite veins and veinlets contained within microcline (potassic alteration) and brown tourmaline replacement zones. The preliminary investigation to recover gold from drill core samples showed that flotation followed by ultra-fine grinding and cyanidation is the preferred option. Ultra-fine grinding is used to liberate gold finely disseminated in sulfides. The flotation concentrate used in the investigation had a P80 of 11 ¦ and contained 75.5 g/t Au, 5.0 g/t Ag, 10.3 per cent S, 0.07 per cent Cu and traces of antimony (0.03 per cent). The mineralogical characterisation showed the presence of 65 per cent gangue minerals, 23 per cent pyrrhotite (hexagonal), 9.6 per cent arsenopyrite, 2.2 per cent pyrite and 0.08 per cent chalcopyrite with trace amounts of galena (0.05 per cent) and sphalerite (0.03 per cent). Gold was in the form of native gold and electrum. A straight cyanidation conducted at pH 11.0, DO of 3 - 5 ppm, 2000 ppm NaCN, 20¦C and for 72 hours extracted 97.0 per cent Au and consumed 31.5 kg/t NaCN. The equilibrium was reached within 24 hours. Increasing the dissolved oxygen to 16 ppm reduced the consumption of cyanide to 22.0 kg/t with 96.7 per cent Au extraction. A 16-hour pretreatment with 2.0 kg/t lead nitrate, followed by a 24-hour cyanidation reduced the cyanide consumption to 6.0 kg/t with a gold extraction of 97.5 per cent. Increasing the lead nitrate addition to 6.0 kg/t reduced the cyanide consumption to 1.2 kg/t by passivating the surface of pyrrhotite. The gold extraction remained at 97.0 per cent. It was possible to decrease the concentration of cyanide to 800 ppm NaCN without compromising gold extraction. The associated cyanide consumption was 1.2 kg/t NaCN. The passivation of gold grains was an issue when the cyanide concentration was decreased to 550 ppm NaCN. The gold extraction decreased to 92.2 per cent. The average lime addition for all the tests was 7.4 kg/t and the average silver extraction was 80 per cent. Efficient cyanidation of an ultra-fine sulfide concentrate with only 800 ppm NaCN represents an advance in leaching of ultra-fine sulfide concentrates. The high addition of lead nitrate would certainly make cyanide destruction easier; consequently reducing the effluent treatment costs and making possible to meet the discharge regulations criteria.CITATION:DeschOnes, G and Fulton, M, 2013. Design of a leaching strategy to extract gold from Eleonore Mine ultra-fine sulfide concentrate, in Proceedings World Gold 2013 , pp 71-78 (The Australasian Institute of Mining and Metallurgy: Melbourne).

Sep 26, 2013