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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 I. A. Agapov, I. S. Anisimov, N. V. Rylov, A. M. Sagitova

"The smelter slags represented silicate crystalline rock with medium to coarse-grained structure with spinifex skeletal network of olivine in hedenbergite matrix. Accessory magnetite, pyrrhotite and copper sulfides chalcopyrite, bornite, covellite, chalcocite and digenite occurred in dissemination in silicate matrix. The sample was studied by optical and electron microscopy with support of quantitative XRD. Modal mineralogical composition included olivine, pyroxene, magnetite, chalcopyrite, quartz, pyrrhotite. Copper deportment showed copper distributed in chalcopyrite (40 %), cuprite (22 %), covellite (21 %), bornite, chalcocite, digenite (5 %), Cu-bearing pyrrhotite (4 %) and spinel (2%). Liberation of copper minerals in the material ground to P80=71 µm was good with over 90 % sulfides in free grains and floatable binaries. The results showed theoretical possibility of 65 % copper recovery by flotation. Conclusion based on mineralogical data was precisely confirmed with flotation tests studies that proved to recover around 60 % of copper to commercial grade concentrate. INTRODUCTION Annual world copper matte smelter slags production is estimated to be millions of tones and contain thousands tones of unrecovered Cu, Zn and significant amounts of Ag and Au. Mineralogical study of one large processing sample of material of historical piles of Cu matte smelter slag was carried out to find possibility to recover copper and define theoretical recovery by flotation. The study of the slag material was aimed to determine possibility of upgrading the material on existing copper flotation plant to develop additional mineral resources. The objectives for mineralogical investigation were to target for best liberation fineness, determine theoretical recovery and possible concentrate quality for sulfide flotation tests. MATERIALS AND METHODS Large representative processing sample was collected by mixing of 120 t of slag material representatively excavated from historical piles of Cu matte smelter slag. The sample was used to define the processing technology to recover copper and gold from the material."

Jan 1, 2018

By P. Altinkaya, O. Celep, H. Deveci, E. Y. Yazici

"Copper-bearing gold ores have proved difficult to treat in conventional cyanide leaching process owing to prohibitively high consumption of cyanide by readily soluble copper minerals interfering with gold extraction. Cyanide leaching of gold from copper-rich ores can be improved by intensive cyanidation or application of physical and chemical pre-treatment methods (e.g., ultra-fine grinding, roasting and leaching). In this study, nitric acid leaching of a copper-bearing pyritic concentrate (0.3% Cu, 62 g/t Au and 110 g/t Ag) as pre-treatment prior to cyanidation was investigated. The mineralogical analysis of the concentrate has revealed that it is composed of mainly quartz (25%), pyrite (32.7%), sphalerite (5.37 %), galena (2.12 wt %) and anglesite (1.13 %). Arsenopyrite, chalcopyrite, covellite and bournonite were also detected as minor phases present. Previous studies carried out by the authors showed that direct cyanidation (1.5 g/L NaCN) of the as-received concentrate (d80: 80 µm) produced low gold (45% Au) and silver (46% Ag) extractions. Nitric acid pretreatment of the as-received concentrate was performed for 5 hours at 1-5 M HNO3 and 80oC. In these tests, 65% Cu/Pb and 94% Zn were leached. The nitric acid pre-treatment improved the extraction of gold and silver in subsequent cyanide leaching (1.5 g/L NaCN) to 82% Au and 87.5% Ag with a concomitant decrease (up to 54%) in cyanide consumption. It can be inferred from these findings that the removal of copper from the concentrate and the liberation of the occluded gold by nitric acid leaching as pretreatments can be appropriately employed for the treatment of the copper-bearing pyritic concentrate with high gold/silver content. INTRODUCTION Refractory ores are often characterized by their low amenability to cyanide leaching resulting low (often <80%) Au/Ag extractions (Bhappu, 1990; Celep et al., 2009; Marsden and House, 2006). One of the most common causes of refractoriness is the physical or chemical encapsulation of submicroscopic or fine-grained gold in sulphide minerals such as arsenopyrite and pyrite (La Brooy et al., 1994; Vaughan, 2004). To render refractory gold ores/concentrates amenable to cyanide leaching, the decomposition of sulphide minerals within a pre-treatment process to expose the encapsulated gold grains to the action of cyanide is required. In this regard, roasting, pressure oxidation and biooxidation are commercially used processes for pretreatment of arsenopyrite/pyrite gold ores/concentrates ahead of cyanide leaching (La Brooy et al., 1994). Presence of cyanide-soluble copper minerals in the ore leads to low gold recoveries and high cyanide consumptions under conventional cyanide leaching conditions. High cyanide concentrations (CN:Cu>4) should be employed to achieve high gold recoveries from copper-rich gold ores/concentrates. This may require recovery of cyanide from effluents. In this regard, SART Process (sulphidisation-acidification-recycling-thickening) employed industrially for recovery of cyanide and copper as a by-product. This process simply relies on the addition of sulphide (as Na2S, NaSH or H2S) to the leach solution prior to the acidification to pH 4-5. This results in the precipitation of copper as Cu2S and occurrence of free cyanide which is recycled to the leaching circuit after S/L separation (Dai et al., 2012)."

Jan 1, 2018

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 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 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 Mohammad Noaparast, Hamed Haghi

"In this research work, it was attempted to reduce the iron contents (Fe2O3) in relatively low grade silica sample which was obtained from Shenin silica mine, using magnetic separation. The prepared sample contained 93.75%SiO2, 0.44%Fe2O3, 2.78%Al2O3, and 0.08%TiO2 and the Fe2O3 had to be reduced to less than 0.08%. Mineralogical studies and XRD analysis indicated the presence of Hematite, Pyrite, and Goethite as the major iron impurity phases. Mineralogical studies of polished and thin sections demonstrated that the degree of freedom of silica is more than 80% in -300+212 ?m size fraction. After primary stages including crushing, milling, scrubbing and de-sliming, various methods such as low intensity magnetic separation (LIMS), high gradient magnetic separation (WHGMS) and dry high intensity magnetic separation (DHIMS) were applied to find out that which process would be appropriate for the elimination of iron impurities from this sample. Different experiments were carried out on the product of wet rod mill and scrubber. However the magnetic field intensity of WHGMS was increased down to 2Tesla, and there was not any acceptable result in reduction of Fe2O3 content lower than 0.08%. Thereafter the product of wet rod mill and scrubber were dried and reported to DHIMS. In these experiments, maximum magnetic field intensity was 1.4Tesla. In this case very significant results were achieved, and Fe2O3 content of final product were reduced to less than 0.05% which meets glass manufacturing standards."

Jan 1, 2014

By E. Barsky

The composition of multi-component materials is proved to be unambiguously determined by the composition uncertainty parameter (composition entropy). At the same time, every component, depending on its content, introduces a certain value into this parameter. Quality criterion of multi-component separation is determined by the total change of this parameter for all products obtained as a result of the separation process. This parameter can unambiguously estimate both multi-product separation and complicated schemes of modern beneficiation plants. It can be used to estimate binary mixtures separation, too. During the optimization of multi-product separation, another problem arises, which is not peculiar to binary mixtures. In this case, the total efficacy depends on the order of separation boundaries, as well. In the present paper, a specific example of initial product separation into seven products by six boundaries is investigated. It is clear that with respect to a single boundary, two products only can be obtained. In this case, the separation optimum is shown to correspond to the separation of boundary-class particles equally into both products. We have also established that from the point of view of entropy, the maximal possible separation efficacy corresponds to equiprobable composition of the initial mixture, i.e., the fine product quantity in the material to be separated is equal to that of the coarse product. Three groups of experiments were carried out. In the first group, the boundaries were consecutively specified from the finest to the coarsest one. In the second group, the boundaries were set vice-versa, i.e., consecutively from the coarsest to the finest one. In the third group, no sequence was observed. At every separation stage, a boundary was specified in such a way that the separated material composition was as close as possible to equiprobable one relative to this boundary. Total efficacy was different in all the three cases. In third group of experiments, its value is the highest, which confirms the advanced supposition. The obtained results are phenomenological, ? they can be applied to any processes of solid, liquid and gaseous materials separation.

Jan 1, 2014

By C. Dorfling, C. A. Snyders

South Africa has extensive mineral processing and extractive metallurgy capacity and infrastructure for metal production from natural resources. Decreasing ore grades, increasing labor costs, and more stringent legislation have made economic metal production from primary resources challenging. Opportunities exist for metal production from secondary resources such as mining wastes and end-of-life consumer goods. The objectives of this study were to evaluate the current status of metal production from secondary resources in South Africa and to identify potential opportunities for further development. Two main classes of secondary resources, namely mining waste and electronic waste, were considered. Primary metal producers are dominating metal production from mining wastes; several of the gold, platinum and chrome producers operate low risk and low cost tailings retreatment facilities that contribute significantly to their overall production. Further opportunities exist in, for example, the treatment of more than 450 estimated gold tailings dumps in Gauteng, as well as metal recovery from slags and furnace dust. The majority of companies involved with electronic waste recycling in South Africa focus on collecting and disassembling; only a few perform preliminary physical processing, with the recovered fractions then being sent overseas for metal recovery. Projections are that the rate of waste generation will increase significantly and that treatment will become critical to comply with more stringent environmental legislation. Projects related to process development, techno-economic evaluations and environmental impact studies for metal recovery from batteries, printed circuit boards, and fluorescent lamps are underway. South Africa is ideally positioned to further explore opportunities for synergistic relations between metal recyclers and primary metal producers. Processing of electronic waste for production of technology metals, for which limited natural resources exist in South Africa, would be a key enabler for local manufacturing industries and expansion of the value chain.

Jan 1, 2018

By I. Baltov

The Chelopech Mine, located in Central Bulgaria, is one of Europe?s largest underground copper- gold mines, currently mines and processes approximately one million tonnes of ore per year. Copper concentrates containing gold and silver with significant arsenic content are produced and sold on the world markets. The operation was acquired by Dundee Precious Metals in 2003. A modernization program for the Mine and Plant has since been in progress to increase both production and mechanical reliability, and also to optimize metallurgical recoveries. Since then, mine output and plant throughput have doubled, plant availabilities and utilization have increased significantly, and overall metallurgical performance for copper and gold improved. The paper presents the systematic approach for obtaining and analyzing data, communicating it, and making suggestions for changes to the technological parameters and flowsheet in order to maximize metal recoveries. These studies have correlated the ore characteristics (particle size distributions, metal assays, and Cu and Au mineralogical and phase analyses) with the actual plant performance (kinetics, recoveries, and throughput). Cu recovery by size has been calculated from the mineralogical analyses of feed, concentrate and tails, with the most significant losses being identified in terms of fraction and phase recoveries. This has provided the recommendations (some of them tested in plant) for changes in the grinding and flotation circuits.

Jan 1, 2014

By Chenyang Zhang, Zhigang Yin, Wei Sun, Jingxiang Zou, Jianyong He, Yuehua Hu, Zhijie Xu

"In this work, a novel compound 4-Amino-5-mercapto-1,2,4-triazole was first synthesized and its selective adsorption mechanism on the surface of chalcopyrite was comprehensively investigated using UV-vis spectra, zeta potential, Fourier Transform Infrared Spectroscopy(FTIR), X-ray Photoelectron Spectrscpy Measurements(XPS), and Time of Flight Secondary Ion Mass Spectrometry (ToF-SIMS) and first principle calculations. The experimental and computational results consistently demonstrated that AMT would chemisorb onto the chalcopyrite surface by the formation of a five-membered chelate ring. The first principle periodic calculations further indicated that AMT would prefer to adsorb onto Cu rather than Fe due to the more negative adsorption energy of AMT on Cu on the chalcopyrite (001) surface, which was further confirmed by the coordination bonding energies of AMT-Cu and AMT-Fe based on the simplified cluster models at a higher accuracy level (UB3LYP/Def2TZVP). The bench scale results indicated that the selective index improved significantly when using AMT as a potential chalcopyrite depressant in Cu-Mo flotation separation.INTRODUCTIONFroth flotation is one of the most important processes to separate valuable minerals from gangue by taking advantage of differences in the physical chemistry of their surface properties(Laitinen et al., 2014), particularly in their surface hydrophobicity. Hydrophobicity differences between valuable and gangue minerals could be regulated by adding different chemical additives (surfactants)(Kor, Korczyk, Addai-Mensah, Krasowska, & Beattie, 2014), such as collector and depressant. It could be conducted economically to selectively separate and enrich the complicated and/or poor ores, by adopting appropriate surfactants. For example, the beneficiation of molybdenite usually relies on a two-stage process to realize flotation separation of molybdenite from other unwanted sulphide ores(Wie & Fuerstenau, 1974). In the first step, a bulk molybdenite concentrate is produced by utilizing kerosene/ diesel oil and pine oil/alcohols as collector and frother, respectively. Second, sodium hydrosulphide, sodium cyanide, ammonium sulphide, or Nokes reagent are adopted as depressants in the separation stage. However, these depressants are usually toxicant and cannot meet the requirements of environmental protection. A large number of efforts have been taken to develop alternative environmentally-friendly depressants (Ansari & Pawlik, 2007; J. H. Chen, Lan, & Liao, 2013; Li et al., 2015; Yin, Sun, Hu, Zhai, & Guan, 2017; Yin, Sun, Hu, Zhang, Guan, Liu, et al., 2017; Yin et al., 2016). Although some novel and environmental-friendly depressants have been reported to achieve good performance in Cu-Mo flotation separation at a laboratory scale, there are still some challenging problems, including toxicity, high price or poor performance in practice, which severely limit their practical applications in Cu-Mo flotation separation on a commercial scale."

Jan 1, 2018

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

The present study investigates the nature of gold mineralogical distribution and association in a Cu-Au flotation circuit (1.6 g/t Au, feed head grade). The approach used in this study involves a combination of bulk gold assay and quantitative X-ray diffraction (QXRD), on a size-by-size basis, followed by QEMSCAN analysis on the -75+53 µm, -53+38 µm and -38+20 µm size fractions. The QXRD and QEMSCAN analysis showed that gold mineralogy is complex and fine-grained in the flotation feed, particularly in the -38+20 µm size fraction. Gold and pyrite concentration follow each other closely, suggesting gold-pyrite association. Quartz as the major gangue mineral shows a bi-modal distribution with respect to particle size. The identified gold particles were completely locked (~94 %) in pyrite as fine inclusions and 6 % locked as complex aggregates (-75+53 µm size fraction) whereas in the -38+20 µm size fraction, the identified gold grains occur as fine inclusions (~1-3 µm) within pyrite and silicates. However, in the intermediate size fraction (-53+38 µm), 75 % of the identified gold grains (10-22 µm) were over 90 % liberated and 25 % associated and locked in pyrite and complex aggregates (~2-4 µm). The identified gold grains in the -53+38 µm size fraction occurred as native and electrum with varying silver concentration between 8 and 20 % by weight. In the -38+20 µm size fraction, a gold telluride was also identified. Based on the mineralogical analysis of the feed, possible flotation strategies are discussed. Optimizing the flotation of pyrite is suggested as the primary Au recovery strategy. Keywords: flotation, pyrite, gold, liberation, QXRD, QEMSCAN

Sep 1, 2012

By P. M. Solozhenkin

On the territory Sakha (Yakutia) are located largest in Russian Federation reserves of antimony ores, Sarylakhsk and Sentochansk deposits, amaunt 10% of world stocks of antimony. A processing gold antimony ores Sarylakhsk deposit is carried out on gravitation-flotation circuit to receive gold gravitation concentrate and gold-antimony flotation concentrate with the contents of antimony over 60%. The gold-bearing concentrate is treated on place hydrometallurgicaly to a high-standard product, accepted by gold bank cash department. At the ore dressing plant reached the highest recovery in world in a concentrate of antimony 90%~91.6%, and gold 85.7%~86.0% as a result of scientific research works. Studied the mechanism of interaction cations of metals Pb2+, Cu2+, Cd2+, Hg2+, Ag+ with a surface antimonite presented on the basis of modern physical-chemical methods: nuclear quadrupol resonance, x-ray electron spectroscopy, electron spin resonance. The industrial tests of sodium dimethyldithiocarbate were carried out which confirmed high recovery of Sb and As. The leaching Sb2S3 by bacteria Desulfovibro desulfuricans and transformation antimonite in antimony trioxide by Tiobacillus Ferrooxidans improved following cyanidation . The extraction of antimony in a solution 96.0%~98.0% under optimal conditions. The common total extraction in the appropriate commodity products for gold reached 98%, for Sb 96.5%. At a combination of bioprocessing and cyanidation 85.5% of extraction Au was achieved. The new solvents for antimonite were found. Now by the Close Joint-Stock Co ?Sarylach-Antimony? industrial plant is constructed for processing Au-Sb concentrates to receive antimony trioxide and gold from various antimony deposits of Sakha (Yakutia).

Jan 1, 2014

By A. Ghadyani

This investigation was undertaken to assess the possibility of ash and sulfur removal by microbial action of high sulfur-content coal. Thiobacillus ferrooxidans and Thiobacillus thiooxidans have been used for biodesulfurization of coal. Feed samples were provided from Tabas coal mine located in Yazd province, Iran. The sample contained 28.80% ash and 2.30% sulfur. Flotation tests were implemented to optimize the effective parameters in these cells which were constant during bio-flotation tests. According to the results, obtained from various tests using Taguchi orthogonal array design (L27), the values of optimized parameters were pH at 7, 1000 g/t gas oil as a collector, 10% solid content, 15 min conditioning time, 100 g/t MIBC as a frother and 800 rev/min rotor speed. Bio-flotation tests are carried out by means of laboratory double discharge cell and mesophilic bacteria as depressant reagent using above mentioned optimized parameters. The bio-flotation experiments were conducted based on a Taguchi orthogonal array design (L9) with four variables (bacteria type, bacteria count, rotor speed of mixer and bio-conditioning time). The collected data was analyzed by using DX7.0, Minitab14.0 and WinRobust1.02 softwares for evaluation the effect of each parameter on the optimization criteria. Considering the highest combustible recovery and the lowest ash and sulfur content, the value of optimized bacteria type, bacteria count, rotor speed of mixer and bio-conditioning time were Thiobacillus ferrooxidans, 5×107 cells/g(coal), 100 rev/min and 20 min respectively. Under this condition, ash and sulfur content and combustible recovery were predicted 12.39%, 2.08% and 82.03% respectively.

Sep 1, 2012

By Jonas Addai-Mensah

Despite the fact that nickel (Ni) laterites accounts for about 60 % of the world?s nickel mineralization, commercial processing of complex, low grade Ni laterite ores by conventional beneficiation (e.g. magnetic, electrostatic and flotation separation) still presents intractable challenges, particularly in heap leaching. Invariably, low grade laterite ores require aggressive hydrometallurgical techniques (e.g., highly acidic lixiviant leaching at ambient or elevated temperatures) for Ni and cobalt (Co) extraction. As a precursor to heap leaching, agglomeration of the multi-mineral phase feed particles into robust and porous granules of 4-10 m permeable bed, is essential. In the present work, several basic and applied studies of the agglomeration behavior of real Ni laterite ores and selected oxides and clay minerals (hematite, quartz smectite and kaolinite) which constitute the predominant host gangue phases of typical laterite ores are performed. Isothermal, batch agglomeration tests involving sulphuric acid solution as a binder were carried out to produce coarse, narrow-sized (5 ? 40 mm) granules which were well characterized. The influential role of feed ore characteristics (e.g., mineralogy/chemistry and primary particle size distribution) and binder composition (30 vs. 44% w/w H2SO4) and dosage (15-25 wt.%) in agglomeration behavior and granule properties were examined through several characterization techniques including granule size, compressive strength, bed percolation and re-wetting stability and 3-dimensional micro-structure analyses. Links between feed mineralogy/chemistry and size, binder formulation, agglomeration behavior, curing conditions and final granule properties and structural characteristics are established. The implications of the findings to Ni laterite plant agglomeration for enhanced heap leaching are discussed. Keywords: nickel laterite, complex low-grade ores, agglomeration, heap leaching, micro-structure

Sep 1, 2012

As the high-grade copper deposits in the world become depleted, there is an increasing focus on developing lower-grade ores which contain problematic impurities such as arsenic. Indeed, many high-arsenic ore bodies remain undeveloped because of their high arsenic levels. Separation of arsenic from copper at the beneficiation stage is difficult using the current processing techniques and invariably copper concentrates produced from these ores contain high arsenic levels, sometimes so high as to make them unacceptable to copper smelters. New, sustainable approaches are needed to allow development of high-arsenic copper ores while at the same time minimizing arsenic dispersion into the biosphere. The most promising route for separating arsenic from copper at the flotation stage centers on the use of pulp potential control. In recent years CSIRO has demonstrated at laboratory scale, for a number of high-arsenic copper ores, that crude flotation separation between arsenic-bearing minerals and copper-bearing minerals can be made. The differences in specific minerals responses to changes in pulp potential and pulp pH are exploited for this separation. Such separations nominally produce a low-arsenic high-copper concentrate and a low-copper high-arsenic concentrate. The former can be smelted without penalty, while the latter requires further treatment to capture and stabilize the arsenic and to recover the residual copper. Consideration has now been given to treatment options for the high-arsenic low-copper product. A pyrometallurgical option, involving a low temperature roast to selectively fume off the arsenic and leave a high-copper calcine, coupled with an arsenic stabilization step, has been demonstrated in the laboratory, and a preliminary techno-economic evaluation shows the process to be economically favorable. These results were recently published. CSIRO has now investigated an alternate hydrometallurgical option, involving atmospheric alkaline sulphide leaching of the high-arsenic product, where the arsenic is selectively dissolved leaving a copper-rich leach residue. The leaching testwork was carried at on the same high-arsenic low-copper material used in the roasting studies, enabling a techno-economic comparison of the alternate roasting and leaching treatment options to be made. The alkaline leaching work is briefly described and the results of the process option comparison presented and discussed. Keywords: arsenic, copper, economics, enargite, ore flotation, tennantite INTRODUCTION

Sep 1, 2012

By Yuhua Wang, Haitao Zheng, Guangli Zhu, Yuehao Zhao, Fushun Yu, Dongfang Lu

"Spodumene, feldspar and quartz possess similar surface properties, which make the difficult to separate spodumene from other silicate minerals by flotation. The differences of crystal structure, surface property and floatability of three minerals were studied by various technical means. The results show that Li-O and Al-O bonds in spodumene structure are not strong and easy to be broken. However, Al-O and Si-O bonds in feldspar and Si-O bond in quartz are very strong and not easy to be broken. High contents of Al and Li on spodumene surface were found by XPS measurements, which is benefit for the adsorption of anionic collector. The surface energies of main cleavage planes of spodumene (110), feldspar (001) and quartz (001) are 1.88 J/m2, 2.97J/m2 and 2.68J/m2, respectively. Hence, the exposed surfaces of feldspar and quartz are not saturated and easy to adsorb OH- ions from solution and cause the lower IEP than spodumene. The contact angles of three minerals in water are lower than 16º, suggesting that the hydrophilicity of three minerals is high, and the order of hydrophilicity is quartz > feldspar >spodumene. Correspondingly, contact angles and flotation recoveries of spodumene are bigger than that of feldspar and quartz in presence of a combined anionic and cationic collector named YOA.INTRODUCTION Flotation technology plays most important role for the recovery of spodumene from pegmatite ore. Generally, spodumene with fresh surface is easy to be floated using soap and oleic acid as collector, and the best flotation pH range is 4.0 to 9.0. However, the flotability of spodumene will become worse when its surface was polluted by slime which produced by the weathering of spodumene and other silicates (Dresler et al. 1998; Moon and Fuerstenau 2003; Menéndez et al. 2004; Yu et al. 2015). In addition, the species of Ca2+, Fe3+, Mg2+ and other metal ions in pulp can not only activate the flotation of spodumene, but also the activation of gangue minerals, and then the flotation selectivity of spodumene from gangue minerals became wores (Wang and Yu 2007; Yu et al. 2014). In order to eliminate the negative effects of slime and metal ions on the flotation of spoddumene, the desliming and long time scrubbing with the addition of NaOH and Na2CO3 were adopted. Usually, NaOH shows the activation for spodumene flotation, and the recovery of spodumene increases with the increase of NaOH dosage. Na2CO3 can eliminate the non selective activation of Ca2+, Fe3+, Mg2+ ions in the pulp, and the concentrate grade of spodumene increases with the increase of Na2CO3 dosage. Anionic collectors such as chelants, hydroxamic acid, oxidized paraffin soap, naphthenic acid and oleic acid were widely used for the flotation of spodumene (Dresler et al. 1998; Moon and Fuerstenau 2003; Menéndez et al. 2004; Yu et al. 2015)."

Jan 1, 2018

By Gautam Chandra Mondal

An attempt has been made to carry out qualitative assessment of mine water discharges from the different mines of West Bokaro Coalfields of Hazaribag district of Jharkhand, India. A total of 33 mine water samples were analyzed for pH, electrical conductivity (EC), major cations (Ca, Mg, Na, K), major anions (HCO3, SO4, Cl, NO3), and heavy metals (Fe, Mn, Cu, Zn, Pb, Cr, Ni, Co, As). The hydrochemical data has been evaluated in terms of major ion chemistry, solute acquisition processes and suitability of mine water for domestic, industrial and irrigation uses. The pH of the analyzed mine waters varies from 5.27?8.43 indicating acidic to alkaline nature of the water. The electrical conductivity (EC) values varied from 276 µScm-1 to 1598 µScm-1 for measured samples and spatial differences between the EC values reflects the wide variation in lithology, surface activities and hydrological regime prevailing in the region. The anion chemistry of the analyzed water samples shows the dominance of SO4 and HCO3 with a minor contribution from Cl, NO3 and F. The observed high values of sulphate in the mine water of the area may be attributed to the oxidative weathering of pyrites. On average chloride is contributing 5% of the total anionic balance (TZ-). Fluoride and nitrate have low concentrations and on average contributing <2% the total anions. The cation chemistry indicates that 79% of the samples are Ca>Mg>Na>K, while the 6% belong to Ca>Na>Mg>K and 1% of Ca>Mg>K>Na. Weathering of rock forming minerals and cation exchange processes normally control the levels of these cations. Concentrations of some heavy metals (Fe, Mn, Ni, Cr) are found to be well above the prescribed permissible levels recommended for the drinking water. The water quality assessment for drinking purposes indicates that most of the mine water having high TDS, total hardness (TH) and sulphate concentrations. This suggests that mine water of the study area is not suitable for direct use in drinking and domestic purposes and needs a proper treatment before its utilization. The assessment of water for irrigation uses show that the water is good to permissible quality and can be used for irrigation uses. However, high values of salinity, %Na, SAR, RSC and Mg-hazard at certain sites restrict its suitability for agriculture uses. Keywords: mine water; ion chemistry; quality assessment; West Bokaro coalfields

Sep 1, 2012

By S. Kelebek, A. Khan, J. Caceres, Beckett. D.

"Pyrrhotite in nickel-copper sulphide ores is nickeliferrous and intimately associated with pentlandite. Due to its low nickel content, pyrrhotite rejection is an important step in processing of these ores. In our laboratory, various reagents were investigated for their effect on pyrrhotite depression. Among these, a combined use of triethylenetetramine and sodium metabisulphite was found highly promising.The work described in this paper has two parts. The firs part is a summary of results obtained on process samples from the pyrrhotite rejection circuit at Falconbridge’s Strathcona Mill. The second part involves simulation of the circuit performance based on the reagent conditions used in the experimental work and the concept of split factors.Two sets of reagent conditions were evaluated. The first one involved baseline conditions, simulating the current reagent performance. The second one simulated the impact these new reagents can have on the circuit performance.The simulation results obtained under baseline conditions indicated a good agreement with the current circuit performance. The simulation results suggest that at practically the same pentlandite recovery and tailing grades, the circuit concentrate nickel grade can be doubled through effective pyrrhotite rejection by these new reagents."

Jan 1, 2003