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By Kr. J. Dedelyanova

Selective flocculation of iron minerals could improve the economics of high intensity magnetic separation as concomitant process. When the minerals are finely disseminated is necessary a finer grinding to be applied which generates a lot of slimes. The HIMS is not enough efficient for particles below 20 microns ? territory where the selective flocculation is already well known industrial practice, so the combination of these two methods of beneficiation can bring a higher grade at high recovery to be achieved. The dynamics and selectivity of the precipitation process is studied in batch tests with flocculant starch and dispersants sodium silicate and sodium hexametaphosphate. The influence of preliminary conditioning with nitrogen acid at pH 2 before flocculation is also investigated. Flowsheet including these two methods is tested and the distribution of iron is shown.

Jan 1, 2014

By Patricio Berrios, David Barriga, Peter Amelunxen, Gerson Sandoval, Roger Amelunxen

Laboratory test equipment and methods have undergone relatively little change since their introduction in the 1950’s. The Denver sub-A cell is still by far the most commonly used laboratory batch cell, but bottom-driven and instrumented cells are beginning to appear. Test procedures have also changed little over the past 60 years, and most engineers continue to rely on empirical models for predicted plant scale performance from laboratory scale tests. This approach is problematic, because changes in the froth phase characteristics in the lab hamper the reliability and reproducibility of the empirical extrapolation to plant scale. In a previous work, the authors reported on the impact of froth recovery at the laboratory cell, showing that variability in the froth recovery can lead to significant error in the modeled kinetic rate constant unless it is properly accounted for in the scale-up model. In this work, we review the role of hydraulic entrainment in the lab cell, and how variability in particle specific gravity, shape factor, and froth residence time can impact the entrainment recovery.

Jan 1, 2014

By Gurubasayya Swamy

This paper presents the consequences of modernization and expansion of process plant by implementing SAG and Ball Mill and Carbon-in-pulp technology at the Hutti Gold Mine. From the inception of the process plant, it has undergone many changes for the improvement. In the beginning, only gravity recovery of gold was carried out by using Pebble Mills and Strake Tables. Later amalgamation, Cyanidation and Zinc precipitation and finally Carbon-in-pulp gold recovery system was implemented. Grinding plant was modernized by implementing SAG and Ball Mill and Knelson concentrator for free gold recovery as a result, the production rate, rate of recovery, reduced man power, Modernization and expansion of process plant has also yielded in safe and smooth running of the plant. Keywords: sag and ball mill, cip, modernization and expansion, productivity, gold recovery

Sep 1, 2012

The ever increasing demand for energy and the progressive depletion of crude oil resources, have renewed interest in oil shale as an alternative fuel resource. Weathered oil shale samples were assembled from El-Nakheil phosphate mine in Quseir area, Eastern desert, Egypt. The oil shale samples were prepared and analyzed by XRD (X-ray diffraction), XRF (X-ray fluorescence spectrometry), and optical microscopic techniques. Rock-Eval pyrolysis and total organic carbon were used to study shale hydrocarbon potential. A leaching procedure using inorganic and organic acids was used to selectively dissolve the oil shale associated ash forming minerals. XRD results showed that calcite is the dominant carbonate mineral, whereas primary silicate minerals are quartz and montmorillonite. SEM (Scanning Electron Microscopy) observations indicated that the organic matter is dispersed throughout the rock matrix. The acid soluble fraction upon leaching was determined to be 37.59 and 42.74 wt% by using 5% HCl and 10% HCl, respectively. The TOC was increased from 20.78 to 29.88 and 34.09 wt% by using 5% HCl and 10% HCl, respectively. Keywords: el nakheil oil shale, characterization, xrd, xrf, sem, elemental analysis, proximate analysis, ultimate analysis, thin section, rock eval pyrolysis, acid leaching

Sep 1, 2012

By A. V. Bogdanovich

In diamond ores processing, final waste tailings size, rejected to dump, is one of the most important parameters. On one side, it is necessary to minimize losses of diamonds, whose size exceeds 1 mm, on the other side ? to provide for the plant high throughput and sufficiently low capital and energy costs, which is in greater part connected with disintegration process. Extensive laboratory, pilot-scale and commercial-scale dressability studies of different types of diamond ores have been performed at JSC ?Mekhanobr Engineering? in recent years. With that, special consideration was given to determination of the required ores? disintegration size that would provide for a high recovery of diamonds. In diamond ores processing, in order to exclude breaking of large diamonds, stage disintegration is used with regrinding of light fractions and heavy fraction X-ray fluorescent separation tailings. On account of rather weak bonds of diamonds with enclosing minerals, in crushing and grinding of diamond ores, selective disintegration takes place, with aggregates? breakup along plane surfaces of least bonding, which facilitates liberation of diamonds and their transition to free state. Heavy components losses with different sizes of final dump tailings were studied in processing of large samples. It was established, that in order to almost completely exclude losses of heavy components over 1 mm size, light fractions had to be ground to 2.5-3 mm size, while X-ray fluorescent separation heavy tailings of -15+3 mm size do not contain diamonds larger 1 mm in aggregates, and their regrinding does not increase total recovery of diamonds. On the basis of the conducted studies, relationships regarding heavy components and diamonds losses with respect to diamond ores final disintegration size were obtained, providing for more rational organization of the ores disintegration and processing flow sheet. Keywords: diamond, waste tailings size, disintegration process, recovery of diamonds

Sep 1, 2012

By Andrzej Wieniewski, Cezary Regula, Bozena Skorupska, Krystian Cichy

"In this paper, experiences from Polish non-ferrous metals ores enrichment plants in the copper and blende concentrates quality improvement are discussed. The blende concentrate being produced in ZGH Boleslaw S.A., with Zn content of about 54,5% is characterized by the relatively high magnesium content reaching a level of 1% resulting from the dolomite presence, while for the zinc production by electrowinning a concentrate in which magnesium content should not exceed 0,3% is required. The implemented new technology for the quality improvement of blende concentrate is based on the leaching of zinc concentrate with sulfuric acid aimed at dissolving of the contained dolomite, and then flotation of the leached concentrate. It allows to obtain a concentrate with magnesium content below 0,1% at the simultaneously increased zinc content to about 57,5%. In the paper, laboratory tests results on the development of optimal flotation flowsheet for the gypsum removal and the industrial tests results concerning the process optimization of this flotation are discussed. In turn, the copper concentrate being produced in Polish copper ore enrichment plants KGHM PM S.A., is characterized by the increased organic carbon content - above 8%. It creates problems with the smelting of this concentrate because of difficulties with the heat removal derived from the combustion of organic carbon in the flash smelting process. The implementation of segregating flotation to the industrial practice allows the effective production of two concentrates with the differentiated chemical compositions - a concentrate with increased copper content and reduced organic carbon content for the flash smelting process, and a concentrate with lower copper content and increased organic carbon content for the shaft smelting process. In the paper, conditions of the segregating flotation and the achieved results are described."

Jan 1, 2014

By Mingxia Liu, Qi Liu, Tao Jiang

"In addition to physical parameters such as small mass and high specific surface area of fine and ultrafine minerals, chemical parameters such as the adsorption affinity of collectors on the minerals also plays an important role in fine particle flotation. In this study, we have examined collectors with two different ionized groups, carboxylate and hydroxamate, on the flotation of fine niobium oxide (pyrochlore). To facilitate comparison, two hydrocarbon radicals were tested: octyl and oleoyl, which led to four collectors that were studied: octyl hydroxamic acid (OHA), sodium octanoate (C8OONa), oleoyl hydroxamic acid (C18HA), and sodium oleate (NaOl). Small-scale flotation and isothermal microcalorimetric titration measurements were performed to determine the flotation recoveries of different sized pyrochlore and the enthalpic behavior after adsorption of the collectors on pyrochlore. It was found that the flotation performance of the carboxylate-based collectors was generally inferior to hydroxamate-based collectors. Sodium octanoate could not float pyrochlore irrespective of the particle size of the pyrochlore tested. With a longer hydrocarbon chain, sodium oleate was effective to float the relatively coarse pyrochlore (>20 µm) but its collection efficiency dropped sharply for particles smaller than 20 µm. On the other hand, the two hydroxamate-based collectors, octyl hydroxamic acid and oleoyl hydroxamic acid, showed good collecting ability to both the coarse and fine pyrochlore. The isothermal microcalorimetric titration studies indicated a consistent correlation of chain length with floatability of collectors for relatively coarse pyrochlore, as well as a positive relationship between the adsorption affinity and fine pyrochlore flotation recoveries. The key finding of this study was that efficient fine particle flotation could be achieved when strongly chemisorbing collectors were used.INTRODUCTION The first detailed study of the effect of particle size on flotation was made by Gaudin (1931) with sulphide ores. In 1942, Gaudin again demonstrated that (Gaudin et al., 1942) fine particles possess different flotation properties from larger particles. Since then the effect of particle size on froth flotation has been an important research topic by many researchers, e.g., Gaudin (1957), Miettinen et al. (2010), Warren (1976) Jameson (2010), summarizing the behavior of fine particles in terms of their low particle momentum, leading to low rate of collision with bubbles which reduces the chance of adhesion. The suggested remediation methods are mainly based on two approaches: increasing particle size or decreasing bubble size (Warren, 1976). Sand-fine splitting followed by separate flotation treatment of the sand and fines, using deep flotation cells, using starvation quantities of collectors, and using long flotation times, are also strategies to address the problems of fine particle flotation."

Jan 1, 2018

By B. Z. Wu

The ore dressing of tin-polymetallic lean ores is a difficult problem and the research emphasis in domestic nonferrous metals mining field. Through research and tackle key problem,and in order to supply reference in dealing with the same types ores, this paper expatiates the exploiture and appli-cation of six item new technology and technics in ore dressing of tin-polymetallic lean ores. It include the technics of multi-step compounding disposing tailings, grinding step by step, grinding and processing circularly technology, high thickness without circle flotation technology, dispersedly flotation tin step by step technology, exceed fine lean plumbum stibium zinc flocculation flotation technology, the synthetically recovery of waste water after ore dressing technology.

Jan 1, 2014

By Zofia Swierczek

This paper reports the results of a detailed mineralogical study of typical, low-grade Australian nickel laterite run-of-mine ores. Successful development of an effective beneficiation process for upgrading such low grade laterite ores relies on good knowledge of the laterite mineralogy and the value metal(s)? distribution characteristics. Both QEMSCAN and quantitative X-ray diffraction techniques were used to conduct the mineralogical analysis of three types of laterite ores, and the results were compared with those obtained by chemical analysis. Some of the major challenges faced included accounting for the mismatch between the nickel values measured by QEMSCAN and those recorded by chemical analysis. This was primarily due to the relatively low concentration of nickel in the ores and its wide dissemination among several major host minerals in complex association. Careful analysis showed that the limit of nickel detection in iron oxide host mineral grains under normal operating conditions was around 4%. Keywords: nickel laterites, mineralogy, characterisation, QXRD, QEMSCAN

Sep 1, 2012

By Mingfei He, Yude Gao

"The leaching kinetics was studied during leaching of indium-bearing zinc oxide to recover indium. The effects of temperature, sulfuric acidity, particle size, stirring rate on dissolution rate of indium were investigated. Results showed that the leaching process was controlled by chemical reaction at the condition of low sulfuric acid concentration. While the acidity increased to 1.8 mol/L, the leaching process was transformed into an internal diffusion controlling model . The corresponding value of reaction orders under low acidity and high acidity were 1.194, 0.156, respectively. Further calculation results indicated that the leaching reaction activation energy was only 10.97kJ/mol under the condition of high acid. The leaching rate constant was inversely proportional to the square of particle size, while the effect of stirring rate on leaching process was insignificant. The leaching process can be enhanced by increasing the temperature and acid concentration, as well as the decrease of particle size. And a final kinetics eqution was proposed. KEYWORDS Indium, extration, leaching kinetics, zinc oxide INTRODUCTION Indium[1,2] plays an important role in many aspects of living and industrial production owning to its special physical and chemical properties. The reserve of indium is rare on the earth, and most of indium is distributed in some other minerals[3]. Some researchers [4,5] have proposed that indium mainly substitutes for the lattice of other minerals by isomorphous. Therefore, the extraction of indium should take into account of the production technologies of those minerals to comprehensively recover the valuable metals."

Jan 1, 2018

By Ramon V. V de Araújo, Roberto B. E Trindade, Pablo R. Fica-Piras

"Lead has been reported to affect blood vessels and kidneys. It also produces adverse effects in the mental and physical development of children and nervous system of adults. It is also considered as a possible cause of cancer. Because of all these threats to the human life, the reduction of its environmental occurrence, by means of recycling, is an important additional advantage to the usual ones derived of the procedures of the secondary metallurgy: lead can be recycled several times maintaining its characteristics equivalent to the primary one.Furthermore, this way of producing the metal avoids the environmental impacts of primary lead melting. Also, as an additional advantage, it triggers the metallurgical process with a raw material in what we observe higher lead content than in the natural ores. Some of the potential sources for these purposes are: ashes, slags, electrode residues, anodic muds, spent electrolytes, solid and liquid wastes, industrial or automotive catalysts, worn-out or obsolete equipment components and/or final consumption products.About 40% of world’s lead production comes from automotive batteries, mainly composed of four parts: plastic box, electrolyte (diluted sulphuric acid), contacts and connectors of Pd-Sb and plates (PbSO4, PbO2 and PbO). The plates represent 75% of the total mass. Metallic parts can be melted at moderate temperatures and then refined. The remaining fraction, constituted by the non-metallic portions, can be recovered by hydro and electro-metallurgical routes. These routes are preferred instead of pyrometallurgical ones due to the ultimate emissions of lead particulates and sulphur dioxide.The principal constituent of the residual paste of batteries is the lead sulphate. In this process, the paste is treated in two stages: in the first, lixiviating with Na2CO3 (desulphurization of the paste), obtaining PbCO3 and Na2SO4; in the second, the lead carbonate is lixiviated with HNO3, producing Pb(NO3)2, soluble, that can be converted to metallic lead by electrolysis. The reaction of PbSO4 with Na2CO3 in aqueous solutions is fast: the global rate is limited by the diffusion of the carbonate ions inside the reaction medium."

Jan 1, 2003

By B. K. Mohapatra

Small manganese ore bodies in south Odisha, India occurs both as stratiform and stratabound ore bodies. These two ore bodies show contrasting mineralogical, petrographical and chemical characteristics. Stratiform type of Mn-ores consists of manganese, iron and silicate- phases in approximately same proportion. In contrast, stratabound ore bodies contain a high proportion of manganese oxides, silicates and phosphorus phases. Phosphorus is found within apatite in the stratabound ore bodies while in the stratiform ore it occurs as an adsorbed state with iron phase (goethite). Both ore bodies are low-grade (17-23% Mn) therefore do not have a market value. However, the stratabound ores are enriched in phosphorus (P2O5: >1%) compared to the stratiform ores. Both the ore types were processed through gravity, magnetic and roast reduction followed by magnetic separation techniques. The strata bound ore could be upgraded to 33% Mn from a feed of 23% by combination of dry and wet magnetic separation methods. But the phosphorus content also suits up (P2O5: 1.2%) in the processed product making it unsuitable for further use. The stratiform ores do not respond to either gravity or magnetic separation because of close association of manganese with iron phases. However, this could be beneficiated through roast reduction followed by magnetic separation. An ore with 17% Mn content could be upgraded to 40% Mn with a 42% yield and 78% recovery. The goethite, having phosphorus, converts to magnetite and is retained within the magnetic fraction, thereby reducing the concentration of phosphorus in the non-magnetic fraction, which becomes the product. This report demonstrates the influence of ore petrology in determining the up-gradational potential of manganese ore in the southern Odisha, India. Keywords: manganese ore, physical beneficiation, phosphorus rich ore

Sep 1, 2012

By Ivan Bonchev, Elisaveta Djurova, Evgeni Visariev, Viktor Stoilov

"Two types of slag cooling is used world wide - cooling in ladles and cooling in pits. The flotation comparison of the type cooling regimes was done at Aurubis Bulgaria. In Copper smelter Aurubis Bulgaria (Pirdop), the residual copper in the slags from Flash furnace and Converters is recovered by flotation. The slag flotation plant processes blend of three parts FSF and one part converter slag. The slag processing comprises: pit cooling, crushing, two stages grinding (autogenous and ball) and flotation. The copper content in flotation concentrate is over 20 % and in flotation tail is under 0.5 %. In order to improve the copper recovery, different cooling rate of the liquid slags should be achieved. Three cooling experiments with flash smelting furnace (FSF) slag have been done: two of them in pits with different volume and one cooling experiment in metallic ladle. The ambient air temperature during the cooling tests was from -4 to +4 OC. Laboratory flotation tests in open cycle were done with an aim to compare the slags behaviors cooled at different conditions. In this paper the effect of different cooling on the flotation process will be presented. The copper recovery and distribution by sizes in the tails, copper content in the rougher concentrate, and the concentrate yield are discussed."

Jan 1, 2014

By Kwan Ho Kim

In this study, free fall tests are conducted with the objective of understanding the breakage properties of waste concrete for the guide of the direction of the better comminution process. Waste concrete of 75mm×53mm is tested at various drop heights with or without heat pretreatment. Test results indicate that a block of waste concrete does not break by free fall, but eventually breaks by repeated free falls. The number of free falls required for complete separation of large aggregates from the cement mortar considerably reduced when the samples were heated at 400?. Therefore, a great energy saving can be obtained when heating the waste concrete prior to crushing. The energy saving was found to be dependent on the height of free fall but in all cases, was still large enough to offset the additional energy required for heating. Also, it was found that the heat pre-treatment gives an additional advantage in that it produced cleaner recycled aggregates. This may result from easier detachment of cement mortar from the aggregates due to weakening of bonding forces by heating. By analyzing the size distribution of the crushed product, a free-fall breakage model is developed using the particle breakage ratio and the breakage pattern. The developed model can be used to predict and to analyze waste concrete processing circuit.

Jan 1, 2014

By Fedotov. K. V., A. Y. Senchenko, V. A. Chanturiya

"The paper describes an efficient process to treat the ore from the largest in Russia copper deposit – Udokan. The recommended process is universally applicable to Udokan ores different in composition and allows reaching high copper recoveries during production of high-grade cathode copper and sulphide concentrate with 45% copper content. There are several promising trends for the process optimization and in the case they are implemented, economics of the whole project could improve significantly. One of the most important optimization options implies reduction of specific energy consumption in the course of comminution operations. Another possible option to optimize the already selected process flowsheet is production of cathode copper from sulphide concentrate directly at a mining and metallurgical plant without sending the latter to other metallurgical works.INTRODUCTIONUdokan copper deposit with resources of more than 20 million tons of copper is the largest in Russia and one of the largest copper deposits in the world. The deposit is located in Kalar District of the Trans-Baikal Territory, 30 km away from Novaya Chara railway station of the Baikal-Amur Mainline.Scientists and engineers from various research institutions have studied Udokan copper ores for years. Along the way, some conceptual process flowsheets have been proposed and each of them has its own pros and cons. The key requirement to the processes being developed was their universal application for processing of all the ore types using a single flowsheet. Due to the hard mining and geological conditions and because of complex material composition of Udokan ores it is of great difficulty to design a process which would provide for maximum economic benefits. That is why the process optimization has been ongoing until now.According to the Russian standards, the main input document for design of a metallurgical plant is referred to as a Process Design Report that covers the following key data sections: analysis of metallurgical testing results, validation of the process, design criteria, process flow diagrams, the mass and metallurgical balances, process description, equipment list and other information necessary for the subsequent design."

Jan 1, 2018

By M. E. Taranenko, V. A. Krivonosov, S. A. Nemykin

"In concentrating production energy-consuming power installations are wide used, the most powerful of them are the wet autogenous grinding mill operating in closed circuit with a spiral classifier. The analysis of the current state of the automation of the grinding process shows that the reserves for maximizing productivity and minimizing electricity consumption have not yet been exhausted on the existing process equipment. Usage of the achievements of the modern management theory reveals an opportunity for further optimizing of the grinding process. The purpose of the research is to develop an automated control system of complex ""the mill-classifier"" to conduct the grinding process in the most operating efficiency, ensuring the maximization of productivity and minimization of specific electricity consumption. The main idea of this work is to study the specific features of the ore grinding process and creation on this basis of automated control system complex ""mill-classifier"" which allows to find and to maintain the optimal mode of grinding based on identification of the current object parameters, observance and regulating of the object state coordinates. Research methods. We used the methods of system analysis, mathematical statistics, modern automated control theory, theory of systems identification, optimal control, modelling of dynamic processes. Developed scientific provisions: - identification algorithm of the loading sub-system management object of complex ""mill - classifier"" allowing to obtain the unbiased estimators of its parameters according to the data of normal object exploitation; - observing device, which differs from the classic version of the observer Luenberger with possibility of estimating not only the state coordinates of the control object, but also the low-frequency perturbation component; - automated system stabilizing the active power of the grinding mode on the basis of a states controller with Luenberger observer; - structure of the software and hardware that implements the developed models, algorithms, methods of control, management and optimization in the framework of the automation of grinding."

Jan 1, 2018

By T. Khonthu

The purpose of this investigation was to evaluate the relationship between different milling procedures, viz. a ball mill and an IsaMill respectively, and the flotation performance of the different mill products. Two ores were used, viz. a base metal sulphide (BMS) and a platinum group mineral (PGM) UG2 ore. Flotation was used as a diagnostic to investigate whether the different milling procedures may have different effects of the floatability of the particles. The present study is part of a research programme investigating this question using a wide range of comminutiondevices and procedures. Similar particle size fractions from the different milling procedures were used in the flotation tests. The particles were characterized in terms of particle size distribution, morphology (using scanning electron microscopy and image analysis software) and rheology tests. The results have shown that in general when comparing similar feeds to flotation in terms of equivalent particle size distributions as indicated by the d80 value, the ball mill generally yielded better recoveries and grades in the case of the BMS and lower chromite recoveries in the case of the UG2 ore compared to the IsaMill. No significant differences were observed in the morphological or rheological characteristics of the particles subjected to the different milling procedures. Although the results appear to indicate that IsaMilling is less advantageous than ball milling in terms of the flotation performance for the two ores studied it should be emphasized that these results do not take into account the relative advantage which IsaMilling has over ball milling in terms of direct and indirect energy consumption and GHG emissions nor, most importantly, does it consider the platinum group element recoveries in the case of the UG2 which may have increased significantly as a result of the finer grind which the IsaMill produces after 4 passes compared to the ball mill after 64 min. Such results which will be reported is due course may show that the greater liberation expected after IsaMilling of the PGMs outweighs any disadvantages indicated in this study.

Sep 1, 2012

By M. Barsky, E. Barsky

It is generally accepted that the optimal flow velocity for gravity classification is the final particle settling velocity or, which is the same, the particle spin velocity. However, we have found differences between the settling velocities and the spin velocities of the very same particles. Here we present the results of experiments in a cascading gravitational classifier comprising eleven stages. The feed material was introduced at each stage successively, from the first to the eleventh. The optimal flow velocity was determined for each case. From these experiments it follows that the value of the optimal flow velocity depends to a greater extent on the point of introduction of the feed (i.e., on the design of the device) than it does on the interaction between the solid phase and the flow.

Jan 1, 2003

By Z. Kwela, A. Manhique, W. W. Focke, W. van der Merwe

"The main features of the de Wet improved alkali fusion process for the manufacture of AZST are: Fewer process are steps necessary; radioactivity is removed by leaching rather than by precipitation; fewer waste streams are generated; there is less waste to discard and alkali reagent is recovered in the form of a sodium silicate solution that is essentially free of radioactivity.Here we report on our efforts maximize the alkali recovery in the sodium silicate product stream: Milled zircon, d50 ˜ 9 µm, was fused with caustic soda pearls in open reaction vessels at temperatures between 650°C and 850°C. Fusion times of 1, 2, 4, 24 and 336 hours (with intermittent milling) were investigated. The fusion products were characterized by infrared spectroscopy, X-ray diffraction and by wet chemical analysis.For prolonged fusion times the fusion products approached equilibrium compositions. However, the phases Na2ZrSiO5 and Na4Zr2Si3O12 observed at 850°C did not form at 650°C. Since these compounds do not hydrolyze in water they limit the recovery of alkali in the sodium silicate product stream. For 2-hour fusions a broad spectrum of non-equilibrium reaction-products, that are rich in sodium silicates, is formed."

Jan 1, 2003

By Z. B. Pan

The innovative point of this new technology can be described that under normal pressure and 70°C, using a new combined activating and oxidizing agent, the minerals bearing arsenic and organic carbon are oxidized effectively. This technology can be applied in commercial scale. The high carbon and arsenic bearing gold ore is taken from Gansu, where hundreds of tons gold deposit. The ore has 0.51% of arsenic and 1.75% of carbon (1.07% of organic carbon). The gold occurs mainly in native with particle size 0.02~0.6 mm and a few are enwrapped by pyrite and arsenopyrite. The traditional cyanide leaching process gains a very low gold leaching rate with only 5%. Through systemic experimental study, the technological flowsheet of flotation process for concentrating gold, then pre-treating the gold concentrate followed by cyanide leaching the gold concentrate are used. With 4.92 g/t gold of head sample, the gold concentrate with 46.52g/t gold grade and 96.53% gold recovery is gained by flotation process. Followed by pre-treatment for 6 hours, the gold concentrate is leached with cyanide after filtration and washing. The total gold leaching rate is up to 95.44%.

Jan 1, 2014