Search Documents

By A. V. Abshilawa, D. G. Talakhadze, R. I. Sturua

"Processing of the tailings accumulated over the lifetime of Madneuli Copper Concentrator is an opportunity to extract additional salable copper and gold products. This paper presents a pre-concentration process for the valuable components (copper and gold) of the accumulated tailings. The developed process makes it possible to obtain a copper pyrite concentrate with a yield of 14.05%, copper grade of 1.36%, and recovery of 83.22% containing 3.78 g/ton of gold at a recovery of 84.31%. The bulk concentrate is produced by a combined process bringing together selective flotation to produce salable copper concentrate and hydrometallurgical processing of the selective flotation tailings to recover gold. KEYWORDS: FLOTATION, ALBION PROCESS, LEACHING, COPPER CONCENTRATE, SELECTION Madneuli Copper Concentrator, Georgia, was put into operation in 1975. Over its lifetime, 40 mn tons of tailings have been accumulated at the tailings storage facility (TSF). Sampling of the TSF has indicated that the average copper and gold grade is 0.23% and 0.56-0.65 g/ton, respectively. Accordingly, the TSF can be classified as a technogenic reserve. This paper discusses pre-concentration of valuable components using flotation, followed by processing of the bulk concentrate in the Albion process. Mineralogical analysis of the accumulated tailings showed that the main ore mineral is pyrite (11.71%). Copper minerals are mainly represented by chalcopyrite (0.7%). The main nonmetallic mineral is quartz (67.3%); sericite, chlorites, etc. also occur. Pyrite is present both as liberated grains and intergrowth. The absolute majority of copper minerals are found as intergrowth with pyrite, less often with quartz. Ore minerals are slightly oxidized. Chemical analysis results are shown in Table 1. Particle size distribution shows that the size class -0.2 mm - 0.045 mm has the highest grade of 0.85 to 0.97 g/ton of gold. Size class +1.25 mm has the lowest grade. Based on the chemical and mineralogical studies and particle size distribution, an optimized process was developed for pre-concentration of the valuable components prior to hydrometallurgical processing. Several gravitation, direct selective, bulk, and bulk selective flotation processes were trialed. (1,2)"

Jan 1, 2018

By Sunil K. Tripathy

Particle characteristics such as size, density, liberation and magnetic susceptibility play vital role in magnetic separation technique. To understand the performance of magnetic separation process for chromite beneficiation industry, particle characteristics need to be analysed. However, to date, this has been an expensive and time consuming task, thus only minimal amounts of this type of data are available in the literature. Several new methods are being developed, that will produce the required mineralogical and liberation data to provide processing information rapidly and inexpensively. Among these, QEMSCAN is a unique and advanced characterisation equipment used for liberation analysis for different ore whereas S.G. Franz Isodynamic Separator for determining the magnetic susceptibility. In Indian chromite deposits, the content of iron bearing minerals in the ore is typical which controls the Cr:Fe ratio in the product while treating in gravity based beneficiation plant. In the present investigation, detailed characterisation of two distinct chromite ore samples from Sukinda chromite deposit, India, was investigated which includes advanced mineral characterisation using QEMSCANand magnetic susceptibility measurement by using Frantz Isodynamic separator, etc.. Presence of haematite and gibbsite along with chromite mineral were identified as major gangue minerals in the high grade ore from XRD pattern whereas Goethite, Quartz, Haematite, Kaolinite, Ringwoodite identified in the low grade ore. The mass magnetic susceptibility of low grade ore is very high 5.2 cm3/g compared to 2.7 cm3/g for high grade ore, due to the presence of abundant quantity of iron bearing minerals. Subsequently, these two types of ore were subjected for magnetic separation for improving Cr:Fe ratio by using different magnetic separators viz. induced roll magnetic separator, wet high intensity magnetic separator, rare earth roll magnetic separator, rare earth drum magnetic separator. The results of magnetic separation studies indicated that, Cr: Fe ratio of the chromite ores can be improved to 3.99 and 1.97 from the feed assaying 3.09 and 1.29 respectively. Studies on rare earth roll magnetic separator indicated that both the ores were upgraded to a maximum Cr:Fe ratio level at optimised condition. Keywords: characterization, QEMSCAN, magnetic separation, rare earth roll magnetic separator, industrial mineral, chromite

Sep 1, 2012

Taknar polymetal (Cu-Zn-Au-Ag-Pb) deposit and concentration flotation plant are located in the northeast of Iran. The flotation feed contains Chalcopyrite, sphalerite and pyrite with 4, 4 and 8 percent respectively. Cu and Zn contents are 1.26 and 3.5 present. Pyrite decreases the quality of base metal concentrates and increases the amount of sulfur compounds produced in the base metal extraction processes. Therefore, depression of pyrite is essential in the concentration of base metal sulfides by flotation. For flotation tests, 4 mixed collector systems, MIBC (20g/t), lime and tow depression systems (sodium cyanide (60g/t)-zinc sulphate and dextrin (300g/t)-zinc sulphate) were used as collector, frother, pH regulator and pyrite-sphalerite depressants in the chalcopyrite flotation stage, respectively, and SIPX, MIBC, lime and copper sulphate (as activator) were used in sphalerite flotation from first stage tailings. The results have shown that, sodium cyanide is a suitable depressant for pyrite and sphalerite in the first stage of flotation. Compared to dextrin as depressant, chalcopyrite recovery was decreased approximately 1.83%. In addition, when sodium cyanide was used as pyrite depressant, copper sulphate consumption raised approximately 90 g/t for activation of sphalerite. From Environment and HSE points of view, dextrin is natural, biodegradable, and non-toxic while cyanide is very high toxic. Keywords: sequential flotation, Cu-Zn massive sulphide, pyrite depressant, dextrin, cyanide

Sep 1, 2012

By N. A. Dementyeva

To improve gold recovery from preg-robbing carbonaceous ores a technology including gravity, flotation, recleaning of the flotation concentrate, intensive cyanidation of the gravity concentrate and carbon-in-leach (CIL) process for intensive cyanidation tailings and the recleaned concentrate was developed. The final flotation tailings are shipped to tailings disposal. Total gold recovery by the proposed technology from the ore with the head grade of 0.8 g/t was 83.9%. The gold loading onto the activated carbon from CIL process was 6.5 g/kg.

Jan 1, 2014

By A. Güney

Currently, recycling technologies use relatively homogeneous plastics because hand-sorting waste feed is too expensive. Many promising technologies for separating mixed thermoplastics are being investigated, but they are still uneconomical and unreliable. Burning plastics for recycling is not used because of poisonous gases released when their burning. PET and PVC cannot be separated using conventional gravity separation techniques because their specific gravities overlap; PET density changes from 1.33 g/cc to 1.37 g/cc and PVC density is between 1.32 g/cc and 1.37 g/cc. For this reason, plastic recycling needs new techniques. Among these techniques froth flotation, also used in mineral processing, can be useful for its cost and simplicity. In this method, a plastic is selected from mixture, tried to regain its natural property, hydrophobicity. In this study the experiments were applied by using plasticizer reagents to make the selected plastics hydrophobic and us-ing NaOH to render the other hydrophilic. Post-consumed PET and PVC are evaluated by concerning the effects of selected parameters (particle size, concentration of plasticizer reagent, amount of NaOH, conditioning time and the effect of conditioning temperature). At the end of the experiments, optimum parameters for selective flotation of PET from PVC was found as particle size of -4.76 + 3.36 mm, NaOH amount of 30 kg/t, conditioning temperature of 75°C, conditioning time of 30 minutes, reagent amount of 300 mg/L. Consequently, PET was re-covered with 99.7% purity and 57% efficiency.

Jan 1, 2014

By O. D. Khmelnitskaya

The research on selecting of an efficient technology of pyrite cinders complex treatment, sulfuric acid production wastes, have been made. The grades in the initial product were 2.04 g Au/t, 35.0 g Ag /t, 0.38%Cu, 1.1% Zn. Laboratory tests on cyanide technology, thiourea and sulphite leaching were carried out. Magnetic and flotation beneficiation and biohydrometallurgical technology were tested. After considering the results of the tests, the most economically efficient technology of pyrite cinders complex treatment with cyanide leaching of precious metals was determined. The technology involves water washing of soluble copper and zinc compounds, lime treatment and cyanide leaching of gold and silver, detoxification and utilization of mineral wastes. The precious metals are recovered from solutions by cementation onto zinc. Copper is recovered from rinse waters by SW/EW and then zinc is precipitated as hydroxide. All the solutions and organic agent are used in recirculation and therefore the technology is practically waste-free. Mineral wastes can be used as iron-containing materials for cement industry. The technology has been tested in pilot-plant scale under continuous operation with complete re-circulation of cyanide solutions. In this process recovery of 54% Au, 45% Ag, 30% Cu and 50% Zn has been achieved.

Jan 1, 2014

By G. Önal

Recovery of nickel and chromite from the tailings has been studied on the samples with 0.214 % Ni and 5.10 % Cr2O3 content and taken from Orhaneli Chromite Concentration Plant. Mineralogical examinations of the samples showed the existence of chromite, olivine, pyroxene, chrysotile, anti-gorite, chlorite, idingite, tremolite and calcite together with trace amounts of millerite, pentlandite, pyrite, pyrrhotite, hematite and limonite type iron oxide minerals in partially serpantinized dunite type particles. In order to recover chromite from Chromite Concentration Plant tailings Mozley Table and Knelson Separator are used and concentrates containing approximately 42% and 46% Cr2O3 are produced respectively with 40% recovery. Most of the nickel remained in the tailings. Therefore, acidic leaching tests are performed in order to dissolve nickel from these tailings. Effects of particle size, acid concentration, temperature, leaching time and solid/liquid ratio on the nickel leaching efficiency are investigated. The optimum leaching conditions are found as 250 g/L acid concentration, 1/5 solid/liquid ratio, 2 hours of leaching time and 60°C temperature, almost all nickel is dissolved at the optimum conditions and acid consumption is about 841 kg/t. Repeated leaching tests are carried out in order to have enough Ni content in the leach liquor for the further treatment of solutions. After repeated leaching tests, leach liquor contains 3 kg/m3 Ni, 48 kg/m3 Fe, 2.2 kg/m3 Si, 23 g/m3 Cr and 590 kg/m3 H2SO4. Ni is recovered from that solution by precipitating with H2S followed by HNO3 dissolution and precipitating Ni as hydroxide with 63% Ni content. Leach cakes are subjected to the gravity separation (Mozley table) to recover the remained chromite and a chromite concentrate with 47% Cr2O 3 content is produced with 53% recovery. As a result of the abovementioned tests, a conceptual process flow sheet is proposed for the treatment of these chromite concentration plant tailings.

Jan 1, 2014

By Yizhong Yuan, Dao Li, Nannan Xue, Tao Liu, Yimin Zhang

A new multi-stage counter-current circulation leaching process for extracting vanadium from stone coal was proposed, and the main factors on multi-stage counter-current circulation acid leaching process were investigated. The results show that the vanadium recovery could reach 86% under the conditions of sulfuric acid concentration of 4 M, liquid to solid ratio of 2 mL/g, temperature of 95 °C, reaction time of 60 min, mean particle size of 125 µm and circulation acid leaching stage of three. The losses of vanadium in acid leaching solution appeared during the four-stage acid leaching process. The strong adsorption performance of new formed KAl(SO4)2(H2O)12 crystal was main reason.

Jan 1, 2014

By Yuri Yemelianov, Stanislav Balikov, Alexander Epiforov, Andrey Boldyrev, Andrey Bogorodsky

"Test work was carried out using copper gold-bearing flotation concentrate with 6.85%Cu, 14.30%Fe, 20.70%S and 2.34%As. Gold and silver grade in the concentrate is 54.6 g/t and 92 g/t, respectively. Gold and copper are valuable components in the concentrate. Gold is partially associated with pyrite and fahlores. Copper in the concentrate can be found as tennantite (Cu12As4S13). The 71% of gold from the flotation concentrate is recoverable by cyanidation with simultaneous cyanide consumption in the amount of 96 kg/t. The major reason of refractoriness is that gold is fine encapsulated into sulfides and the presence of cyanicide-copper. The concentrate ground as much as 95% minus 10 µm was treated using low-temperature POX at the temperature of 100±5°C, L:S ratio of 5:1 and the process time of 1-5 hours. Residual sulfide sulfur in the cake is 2.28% which is 89% sulfide oxidation. The highest copper recovery to POX solution is 91%. It was found that gold recovery to cyanide solution is 85-94% during alkaline pre-treatment of POX cakes followed by cyanidation. In this case, cyanide and lime consumption is 75-90 kg/t and 120 kg/t, respectively. In order to recover precious metals from POX cakes leaching using sodium sulfite was carried out. Test work was conducted at varied reagent concentration, L:S ratio and leaching time. Gold and silver recovery achieved was 85-95% and 59-63%, respectively."

Jan 1, 2014

By A. A. Lavrinenko

There are discussed the possibilities of using of flotation tailings of platinum-copper-nickel sulfide ores of the Norilsk region for the preparation of backfill mixture instead of a specially produced rock and binder materials to reduce the cost of stowing operations and impact on the environment. It is shown that due to the presence in the tailings of flotation up to 10% of sulfide sulphur, which reduces the strength of the fill materials, and more than 1 g/t PGM it is necessary the preliminary extraction of the sulfides in the product for further processing that give an opportunity to reduce the losses of valuable components. It is shown that by the flotation of the current tailings it is possible to extract 91% of sulfide sulphur, 60% of palladium and 30% platinum in 17% sulfide concentrate containing 2.1 g/t palladium, and to reduce the sulphur content in the tailings up to 0.7%. Due to the easy oxidation of pyrrhotite, there is recommended fractional delivery of the collector with a total consumption 200 g/t. There are discussed results of the flotation study with an additional collector of PGM (sodium diizobutyditiophosfinate), which one show the possibility to increase extraction of platinum on 45%, as well as extraction of sulphur and palladium. For the flotation of sulfides and PGM from the current tailings there is recommended a combination of collectors for primary flotation and only butyl xanthate using in control operations. The possibility of using sludge fraction of the current tailings as a component for the preparation of the combined cementing materials after exposure of sulfides is analyzed. It is shown that the efficiency of the flotation of sulfides from this fraction can be enhanced using selective flocculants. There are discussed the advantages of implementation of completely hydrophobic butadiene-styrene copolymer in comparison with partially hydrophobic polymers. There is discussed an efficiency of the mechanochemical treatment of old tailings in turbo mill and sulphuric acid to remove oxide films from the surface of pyrrhotite before its flotation. Due to the pretreatment operation there is achieved increasing of sulphur extraction and after the introduction to the flotation circuit three control operations there is produced the fill material with sulphur content 2.5%. Keywords: tailings, copper-nickel ore, pyrrhotite, butyl xanthate, sodium diisobutyl dithiophosphinate, selective flocculation, butadiene-styrene copolymer, mechanochemical treatment

Sep 1, 2012

By Tamara F. Kondrateva, Evgenia E. Savari, Galina V. Sedelnikova, Rauf Y. Aslanukov, RAS (INMI), Grigori I. Karavaiko

"The investigation is dedicated to study of biohydrometallurgical processing of refractory gold-arsenic concentrate with the large content of pyrrhotite. Sulfide concentrate of chemical constituents: 75g/t Au, 4,7g/t Ag, 6,3% As, 29,3% S, 2,3% Sb, 0,5% ? (organic) and of mineral composition, %: 11,8 arsenopyrite, 39 pyrrhotite, 12 pyrite, 2,9 antimonite and the rest – aluminosilicates, carbonates and quartz, was studied. The recovery of gold from concentrate by cyanidation is 51,2%. The rest of gold is embedded as finegraded segregation in sulfides.Sulfide oxidation process study is described. Bacterial oxidation was executed on a pilot laboratory unit using combined autotrophic bacterium: A. thiooxidans , A. ferrooxidans, Leptospirillum ferrooxidans and sulfobacillus thermosulfidooxidation, pH=1,5-2,2, t=28-32?C, oxidation time 120 hours. Pyrrhotite is the least stable sulfide mineral with electrode potential 0,45V versus arsenopyrite (0,55 V) and pyrite (0,55-0,6 V) and is oxidated primarily. Pyrrhotite oxidation is followed by large amount of elemental sulphur, which has a bad influence on cyanidation.It’s established that the rate of arsenopyrite (for the first 48 hours) from concentrate containing 39% pyrrhotite is 2-3 times as less as the rate of arsenopyrite bio-oxidation from concentrate without pyrrhotite. The arsenopyrite bio-oxidation is accelerated only after the elemental sulfur was oxidized by bacterium A. thiooxidans and elemental sulfur percentage was reduced from 6,8 to 3,9%. The bacterium concentration is increased from 2,5x107 to 2,5x109 cells/ml. Almost full oxidation of arsenopyrite (98%) and pyrrhotite (98- 99%) is finished in 120 hours. We have studied two pretreatment methods for elimination sulfur content influence on cyanidation: aeration in lime and electrochemical treatment. Such an operation enables to have supplemental oxidation of elemental sulfur and other oxygen and cyanide absorbants before cyanidation. After treatment the content of elemental sulfur was decreased from 3,9 to 2,2%, gold recovery from 90 to 96% and cyanide consumption from 16,2 to 7,2 kg/t decreases. The content of toxical rhodonite-ions in liquid phase of pulp and disinfection reagents consumption were decreased as result of elemental sulfur content in bio-oxidated product reduction."

Jan 1, 2003

By G. Onal, E. Gock, H. Dincer

"During the conventional chromite dissolution process; chromite ore concentrates were ground to at least <90 µm using the ball mill and after adding Na2CO3 and leaning material (instead of CaCO3 also residues from the leaching process) at a ratio of e.g. 1:0.75:1.2, depending on the specific formulation, follows the digestion e.g. in a rotary kiln, heated in countercurrent. The reaction time is 3 to 8 h. The calcine is comminuted and leached with water or a thin chromate solution. Depending on the type of ore used, the yield is 75-85 %. The disadvantage of the above-described process is that it is impossible, using current technology, to achieve a complete digestion of the chromite ores and that the chromium-containing residues have to be disposed of as special waste, involving a large financial effort.The subject of this study was the investigation of the effect of vibratory milling on chromite dissolution. The impact stress in the vibratory mill causes the mechanical activation of chromite, which is characterised by lattice defects. The degree of mechanical activation can be quantified by measuring the relative X-ray diffraction intensities at the parent lattice planes of chromite. Each relative X-ray diffraction ratio correlates with a certain degree of digestion. As opposed to vibratory milling, a complete digestion of chromite using a ball mill is impossible, even at extremely long grinding times.In this investigation a chromite concentrate from Pulpinar/Kayseri with content of 44 % Cr2O3 and the Eccentric Vibratory Mill was used. In order to verify the effect of the vibrating grinding on chromite dissolution, reaction kinetic investigations on digestion and leaching were carried out."

Jan 1, 2003

By E. Yigit

In the flotation process the physical factors such as agitation intensity air flow rate and pulp solid concentration are very important parameters affecting grades and recoveries of the products. The flotation tests were performed on the samples of Bigadic colemanite ore to investigate these factors at different agitation intensities ranging from 1000 rpm to 1400 rpm and air flow rates from 1000 cm3 ? min-1 to 2000 cm3 ? min-1 with 2 L and 4 L cells using Denver laboratory flotation machine. The tested collectors were Cytec Aero 825, Aero 848N and their mixtures. Agitation intensity and air flow rises resulted in more floated material although sometimes being not selective. These effects can be attributed to powerful collisions between air bubbles and mineral particles. The first observation was the effect of pulp solid concentration on the floatability of minerals, while the flotation in 2 L cell (0.125 g.L-1 solid concentration) at 1000 rpm (0.176 wattcm-3), and with 1000 cm3.min-1 air flow rate was possible, the flotation in 4 L cell (0.0625 g.L-1 solid concentration at 1000 rpm. (0.176 wattcm-3) and with 1000 cm3 ? min-1 air flow rate was not possible. However at 4 L cell after increasing agitation intensity to 1200 rpm (0.201 wattcm-3) keeping 1000 cm3 ? min-1 air flow rate, flotation was possible again. This shows that these factors such as agitation intensity have some critical role on flotation. Another interesting observation was on the selective effect of pulp solid concentration. In the flotation with 1 L cell at different agitation intensities, collector Aero 848N was selective against calcite, while in the flotation with 2 L cell at different agitation intensities the same collector Aero 848N was selective against colemanite.

Jan 1, 2014

By Rakesh Kumar, S. Mehrotra, T. Alex

"This paper, an overview of our recent findings on mechanical activation (MA) of minerals/materials, emphasizes the importance of sample genesis, mechanically induced/altered porosity and surface area, mineral interactions, glassy inorganic materials, surface alteration and zeta potential changes and, mechanically induced reactivity of solids. Materials investigated include Al-oxyhydroxides (gibbsite and boehmite), fly ash and granulated blast furnace slag. Planetary, attrition, vibratory and conventional ball mills are used for MA. Genesis of boehmite has been found to have a bearing on its behavior during MA. For example, BET specific surface area (SSA) of boehmite prepared by thermal dehydroxylation of gibbsite has decreased from 265 to 65 m2/g during 240 min of MA; in contrast, the SSA of hydrothermal boehmite has increased from ~2 to 36 m2/g. Mechanisms of surface area changes in the two boehmite samples differing in their genesis is explored. MA of gibbsite and its mixture with hematite/quartz using attrition milling has indicated that the presence of the second phase promotes amorphisation. Amorphous materials show signs of change in nature of bonding during milling as illustrated by MA of fly ash. Physicochemical changes during MA are not restricted to the bulk phase. The genesis of zeta potential changes in MA solids can be many; for example texture induced zeta potential changes as observed for gibbsite, while zeta potential changes during the milling of granulated blast furnace slag occur due to enhanced (possibly non-congruent) dissolution. Interesting manifestations of mechanically induced reactivity and its correlation with physicochemical changes are illustrated in terms of alkali leaching of boehmite and, hydration of MA slag. New findings reported here are associated with several open questions and are highlighted in the paper."

Jan 1, 2014

By W. Yao, M. L. Li, R. Cui

"The purpose of this research is providing a flotation process for recovery of lead in lead silver residues from a zinc smelting plant with a new hydroximic acid collector (HOC). The collecting ability and reaction mechanisms of HOC were investigated based on pure mineral flotation tests and analysis of zeta potential. The pure anglesite flotation tests showed that HOC has a good collecting ability for anglesite. Zeta potential measurements proved that HOC adsorbed on anglesite surface physically and chemically. The lead-silver residues in a zinc smelting plant with lead grade about 17.5% was used as raw material for flotation tests. The results showed that under the conditions of pH adjusting agent lime 500g/t, inhibitors sodium silicate 2000g/t, collector HOC 1500g/t, frother alcohol oil 75g/t, the grade of the lead concentrate can be reach 47.18% and the recovery was 76.39% by one roughing-one scavenging-three cleaning and middles back to the flow-sheet in turn after the lead-silver residues were washed by two stages of cleaning-filtration.INTRODUCTION Nowadays, with the increasing exploitation of natural mineral resources, more and more secondary resources should be researched and used for the valuable metals. Hydrometallurgical zinc leaching residues, which contain a considerable amount of zinc, lead, silver and other minor valuable constituents, is a typical and abundant secondary resources (Zhang et al., 2011). A huge quantity of zinc leaching residues (ZLR) had been produced per year (Jiang et al., 2017).This residues not only occupy precious land but also threaten the environment because of the dissolved heavy metals especially in rainy days (Han et al., 2014). Consequently, it is urgent to develop an effective and clean method to dispose this residues which can benefit not only in economical but also in environmental (Jha et al., 2001)."

Jan 1, 2018

By Caroline Olsen

The objective of this study was to achieve Zn activation control in an industrial Cu flotation circuit treating a complex Cu-Pb-Zn-Au-Ag sulphide ore at the LaRonde Division of Agnico-Eagle Mines Limited (Québec, Canada). In order to provide process relevant results, a laboratory testing methodology was developed to duplicate plant operational parameters. The action mechanisms of ZnSO4, TETA (triethylenetetramine) and NaHSO3used as depressants in order to limit ZnS activation were examined. Laboratory flotation tests with and without Cu2+ions (the ions responsible for ZnSactivation)were performed. All depressants showed equal or better flotation performance relative to baseline tests. For example, in the presence of Cu2+, the Zn recovery reduced from 32.2% to 17.5% in the copper rougher concentrate after adding ZnSO4(with a dosage of 50 g/t). TOF-SIMS analyses confirmed the relative decrease in surface Cu in the samples treated with ZnSO4and TETA. For the case with TETA, this indirectlyindicates the formation of a copper TETA complex in solution, thereby reducing available Cu for transfer. The TOF-SIMS and XPS data indicate that the addition of ZnSO4and NaHSO3resulted in an increase in surface metal hydroxides and SO3 on ZnS grains inhibiting both Cu and collector attachment.ZnSO4 was subsequently tested with success in the LaRonde concentrator. The addition of 40 g/t ZnSO4 at the conditioning (or cleaning) stage led to a 2% reduction in Zn content in the Cu and PM concentrate. Recently, improvements towards the control of copper sulfate addition in the downstream Zn flotation circuit resulted in the reduction of Cu2+ in process water. As this ?cleaner? water is recycled for Cu and PM flotation, the addition of ZnSO4 is now used only as required.

Sep 1, 2012

By Y. E. Brylyakov, M. A. Kostrova

"The apatite staffelite ores (ASO) of Kovdor deposit are dominated by fine-grained aggregates containing both phosphate minerals and silicates — vermiculite, smectites, zeolites. These minerals exhibit ion-exchange properties. In this connection, the magnitude of cation exchange, along with the content of P2O5 and CO2, is an important property of apatite staffelite ore, which largely determines the performance of concentration processes. In the examined ASO samples, the magnitude of cation exchange measured in terms of Ca2+ is 2546-4790 g/ton. The conducted laboratory studies of apatite staffelite ores with various mineral composition showed that adding calcined soda improves the recovery of P2O5 into the flotation concentrate. Full-scale tests were conducted at the apatite staffelite ore processing plant at Kovdor Mining and Concentration Operation and demonstrated the high performance of the proposed agent treatment. The test results showed that in the flotation of apatite staffelite ores, the optimum calcined soda feed rates range from 3,500 to 5,000 g/ton of the flotation feed. With close flotation feed and concentrate grade parameters, the P2O5 recovery gain is 17%. The effect of adding soda is associated with the desorption of the hardness cations from the minerals, precipitation in the pulp and, accordingly, improvement in the ionic composition of the liquid phase of the flotation feed due to a reduced content of calcium ions. Based on the test results, the proposed agent treatment with an elevated feed rate of calcined soda was implemented at Kovdor Mining and Concentration Operation in the flotation of apatite staffelite ore. The apatite staffelite ores (ASO) of Kovdor deposit are dominated by fine-grained aggregates containing both phosphate minerals and silicates — vermiculite, smectites, zeolites. These minerals exhibit ion-exchange properties. These minerals have certain properties requiring a more detailed examination for a better understanding of the flotation properties of apatite staffelite ores."

Jan 1, 2018

By Felix M. Zhuravlev, Vitalij P. Lyalyuk

"While there is a large industry based on recovering platinum group metals (PGM) from complex sulphide ores containing Cu and Ni, there is little information on the thermodynamics of the matte phase formed on smelting the ore and in particular the behaviour of Pt in the matte.The equilibrium between metallic platinum and industrial mattes was studied in a series of experiments where a small quantity of matte (5g) and platinum foil of 150 µm thickness (0.5 g) were equilibrated in magnesia crucibles at 1400°C under conditions where the oxygen and sulphur partial pressures were controlled.The chemistries of the phases were determined from EDS analysis using a scanning electron microscope. Line scans were initially used to confirm that the foils had attained a uniform composition across their thickness. The composition of the matte was determined from analysis of an area of matte including the small metallic droplets and dendrites.The distribution of Pt in Cu-Fe-Ni mattes (Pt: 1.5 – 3%) and solid Pt alloys (Pt: 54 – 72%) was studied. For a matte of base composition 45%Fe, 11%Ni, 8% Cu, at a base temperature 1400°C, the following behaviour was observed for the Pt distribution ratio:Under the range of sulphur and oxygen pressure of this study (log pO2 from -15 to –12.2, log pS2 from – 4.6 to –2.2), the distribution ratio of Pt between matte and alloy was independent of the sulphur and oxygen pressure, having a value of 0.04.The Pt distribution ratio increased from 0.02 to 0.08 with temperature from 1340 to 1450 oC.For mattes where the proportions of Cu and Ni were kept constant, a slight decreasing trend was shown for the distribution ratio of Pt versus the iron content (37 to 53%) in the matte but was not confirmed because of experimental uncertainty."

Jan 1, 2003

By Steven Wright, Sharif Jahanshahi, Shouyi Sun

"While there is a large industry based on recovering platinum group metals (PGM) from complex sulphide ores containing Cu and Ni, there is little information on the thermodynamics of the matte phase formed on smelting the ore and in particular the behaviour of Pt in the matte.The equilibrium between metallic platinum and industrial mattes was studied in a series of experiments where a small quantity of matte (5g) and platinum foil of 150 µm thickness (0.5 g) were equilibrated in magnesia crucibles at 1400°C under conditions where the oxygen and sulphur partial pressures were controlled.The chemistries of the phases were determined from EDS analysis using a scanning electron microscope. Line scans were initially used to confirm that the foils had attained a uniform composition across their thickness. The composition of the matte was determined from analysis of an area of matte including the small metallic droplets and dendrites.The distribution of Pt in Cu-Fe-Ni mattes (Pt: 1.5 – 3%) and solid Pt alloys (Pt: 54 – 72%) was studied. For a matte of base composition 45%Fe, 11%Ni, 8% Cu, at a base temperature 1400°C, the following behaviour was observed for the Pt distribution ratio:Under the range of sulphur and oxygen pressure of this study (log pO2 from -15 to –12.2, log pS2 from – 4.6 to –2.2), the distribution ratio of Pt between matte and alloy was independent of the sulphur and oxygen pressure, having a value of 0.04.The Pt distribution ratio increased from 0.02 to 0.08 with temperature from 1340 to 1450 oC.For mattes where the proportions of Cu and Ni were kept constant, a slight decreasing trend was shown for the distribution ratio of Pt versus the iron content (37 to 53%) in the matte but was not confirmed because of experimental uncertainty."

Jan 1, 2003

By P. Hatefi, O. Rahmani, M. Noor-Mohammadi, A. Tohry

"Depressants play an important role in the selective separation of impurities from iron minerals by direct or inverse flotation. In this paper, sodium co-silicate (SCS) with chemical formula of Na4SiCO6, as a depressant with pH modifier features, has been synthesized and used for phosphorus bearing iron ore in anionic reverse flotation. In this process, phosphorus is floated while the iron oxides are depressed, using fatty acids as collector in an alkaline medium. Batch and pilot flotation tests were carried out to investigate the influence of SCS on the separation selectivity of phosphorus from iron oxides. Also the depression mechanism of SCS on hematite surface was investigated by microflotation tests and FTIR spectroscopy measurements. The flotation results indicated that without using a pH modifier, the pH of pulp could reach up to 10, when SCS was used as depressant. In this condition, a better depression performance was observed compared to sodium silicate (SS). Using the SCS, phosphorus and total iron content in iron concentrate was 0.046% and 66.29%, respectively, compared to SS where these values were 0.06% and 65.75%, respectively. As a result the weight recovery of iron concentrate was increase by around 4%, compare to SS. It was also found that 238 g/t of SCS could be used instead of 3 types of reagents collectively, including 32 g/t of NaOH, 151 g/t of Na2CO3 and 400 g/t of SS gel.INTRODUCTIONThe phosphorous presents in iron ore minerals is mainly in the form apatite mineral. The apatite is separated from iron ore minerals by magnetic and flotation methods (Ma, 2012). The reverse anionic flotation of iron is one of the most common method for removing phosphorus from iron concentrate (Araujo et al., 2005).Usually apatite minerals are readily floated using anionic collectors such as fatty acids (Ma, 2012). In this conditions the iron minerals are depressed and by modifying the pH in alkaline conditions, selective separation of phosphorus from iron concentrate can be achieved. Depressants including; sodium silicate (SS), starch, dextrin, guar gum and humic acids play an important role in the reverse flotation of iron ore (Kar et al., 2013; Ma, 2012; Turrer and Peres, 2010; Araujo et al., 2005). Studies have shown that SS has a better effect on iron recovery and phosphorous reduction, compared with other depressants (Tohry and Dehghani, 2016; Qi et al., 1992, Dho and Iwasaki, 1990). In most flotation processes, modifiers are used to adjust the pH. To achieve alkaline conditions, mainly pH modifiers such as carbon dioxide and sodium carbonate are used."

Jan 1, 2018