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By Cesar Gomez, Jaime Álvarez, Lisette Wen

Laboratory mechanical flotation cells are extensively used in mineral processing for running batch flotation tests with a variety of purposes, for example assessing feed characteristics, determining flotation kinetics, or testing of new reagents and procedures. It is known that the results obtained depend on the operation mode, self-aerated or forced-air, and on the operating conditions selected for the test, in particular impeller speed and gas flow rate. These different results are a consequence of changes in size distribution of the generated bubble population and its remarkable evolution when frother is added; variations in bubble size affect the bubble surface area flux (BSAF), a variable which correlates with particle collection. Results of a two-phase (air-water) test program designed to measure bubble size for a variety of combinations of impeller type and speed, gas flow rate, frother type and concentration, and cell volume, demonstrated the wide range of BSAF with which the cell can be operated. The results obtained in this work demonstrated that gas dispersion in a forced-air lab flotation machine is significantly affected by the characteristics of the impeller/cell geometry combination in use, and the operating conditions selected. Operating conditions must be carefully selected and maintained during the test to ensure consistent results, in particular when frother concentrations below 20 ppm are used, because frother concentration in the cell will decrease as flotation progresses as a consequence of frother partitioning to the concentrate.

Jan 1, 2014

By A. N. Zagorodnyaya

The results of research on recovery of ammonium perrhenate from by-products of lead production are presented in the article. For the first time in the world, rhenium marketable products were produced on lead plant of Kazakhstan. Distribution of rhenium in lead production has been studied. During oxidative agglomeration of lead charge, 67.43% rhenium is sublimated and at that, 67.27% rhenium is trapped by the dusts of a system of gas purification, out of them 6.21% is transferred into different solutions, 1.89% into slime and only 0.09% into exhaust. During agglomerate smelting, up to 72% of its rhenium is sublimated and condensed in the dusts also. At leaching stage of sulphate product under dusts processing by the method of sulphatization, 15% of its content in the dusts is transferred into solution. The sorption method on the ionite ?N-21 has been proposed for recovery of rhenium from the solutions of sulphatization. At the same time, rhenium recovery from the dusts into commodity was 13%. The method of direct hydrometallurgy - leaching the dusts by solutions of sulphuric acid under presence of manganous concentrates, has been proposed as alternative method of recovery. This method has allowed increasing recovery of rhenium and other metals from the dusts into solution up to 90%~95%. Rhenium solvent extraction by ??? has been proposed for recovery of rhenium from the leaching solutions. At that, extraction of rhenium into marketable product from the dusts is above 75%.

Jan 1, 2014

By Liu Xiaodong, Jiang Meifen, Kou Xiaoyan, Lu Xianjun

"Two low grade calcium bentonite (Ca-bentonite) from Hejia (HJ) and Boteng (BT), Shandong, China, mainly contain montmorillonite and quartz, and slightly feldspar, calcite, and montmorillonite content are 32.37% and 55.90%, respectively. In order to examine the feasibility to produce activated clay to meet the standard of Chinese chemical industry (HG/T2569-2007), three types of industrial production process are proposed according to the screen test. The first process is grinding the raw bentonite to -200 mesh and then activating. The second is to activate the raw bentonite directly, and then classify the production to -200 mesh. The last one is to activate the raw bentonite after purification. Three parameters, which are sulphuric acid strength, ratio of liquid to solid and treatment time, are optimized via orthogonal test for every process. The preliminary test indicates that it is feasible to produce acid bentonite of difference quality. For the sample of HJ, milling the raw to -200 mesh first and activating afterwards, the parameter is at 9% acid strength, the ratio of liquid to solid 2, the time 4 hours, temperature 90°C, the products can meet the qualified grade for animal and vegetable oils refining. For the sample of BT, milling the raw to -200 mesh first and activating afterwards, the parameter is at 15% acid strength, the ratio of liquid to solid 2.5, the time 2 hours, temperature 90°C, the products can meet the qualified grade of high decolorization for petroleum products refining. For the sample of BT, adopting purification to -325 mesh by stirring and rubbing at first, and then activating at 9% acidity, under the conditions that the ratio of liquid to solid is 2, the time is 2 hours and the temperature is 90°C, the products can meet the first grade of high activation degree for petroleum products refining."

Jan 1, 2014

By V. F. Petrov, O. D. Khmelnitskaya, V. M. Lodeischikov

"The processing of copper gold ores is one of the major issues of gold-mining industry in the beginning of 21st century. In this case, application of ammoniacal cyanidation process for this type of mineral raw materials deserves special attention. Irgiredmet Institute has carried out a wide range of studies on ammoniacal cyanidation process and evaluated the possible commercial use of this process for Russian operating plants. Copper gold-bearing ore from a current Russian deposit was studied. The content of gold and copper in the ore is 9.4 g/t and 0.4%, respectively. Based on the results of preliminary test work, a flow sheet was developed. It included gravity concentration followed by cyanidation of the gravity tailings using CIP. The flow sheet provides sufficient total gold recovery (91.5%). However, this technology has disadvantages, such as high specific consumption of reagents: 4.5 kg/t NaCN and 13.5 kg/t Ca(OCl)2 the ore and the increased gold grade in the liquid phase of the tails slurry (0.2 g/t). In order to improve cyanidation process parameters, test work on gravity tails processing using ammoniacal cyanidation was carried out. After the test work was completed, ammoniacal cyanidation flow sheet for copper gold-bearing ores was developed. Leaching conditions were optimized using recycled process solutions. Recommendations on gold recovery from the solutions and tails detoxification are given. The proposed flow sheet provides 91.5% gold recovery. In this case, compared to the conventional option, the consumption of cyanide decreases by 3-4 times.INTRODUCTION The technology of precious metals recovery from gold-bearing ores is based on cyanidation. This accounts for up to 90% gold recovery and the significant portion of silver (Lodeischikov, 1999). However, there exists the category of so called refractory ores. They are not amenable to direct cyanidation (without using additional expensive processes), or they are characterized by low indices of metals recovery and high consumption of reagents. The latter also include copper gold ores. Their deposits can be found extensively both worldwide and in Russian Federation. A feature of these ores is that they contain not only gold but also copper. It is considered not only as a valuable component but also as a harmful mineral impurity that hinders gold cyanidation, gold precipitation from the solutions by cementation onto zinc dust and adsorption onto activated carbon (CIL, CIP-processes), detoxification of cyanide bearing effluents."

Jan 1, 2018

Results of the study of chloride ion concentration effect to osmium-187 sorption from wastewater of copper smelter with use of foreign anionites of different basicity and structure are presented in the work. It was found that the increase of chloride ion concentration in the wastewater helps to increase recovery of osmium using anionites regardless of their nature. The ways to intensify the process are proposed. The method for osmium elution from the anionites with potassium iodate solution was developed. It was shown that the method lets to recover up to 99% of osmium from the anionites. Recommendations for selection of eluents for better recovery of osmium from the anionites were elaborated. Keywords: wastewater, osmium, anionite, sorption, chloride ions, eluent, recovery

Sep 1, 2012

By A. Ceylan

Pyrite is present as a gangue mineral in most of complex sulphide ores and most of time it is removed as a tailing. However, beneficiation of pyrite from tailings could be important because of its economical value and for the environmental concern. This paper presents a research for the enrichment of pyrite from copper plant tailings, Siirt-Madenköy/Turkey, by flotation. Siirt-Madenköy copper mine is operated by Park Electric Production Mining and Industry since 2005. In order to obtain the pyrite from Siirt-Madenköy copper plant tailings, laboratory test work has to be done to determine the type of reagents, the flowsheet, and the size of plant. The objective of the company is to develop a pyrite flotation circuit for concentration of pyrite from tailing dam to obtain 500.000 tons/year pyrite concentrate. There are about 3 million tons tailings that containing pyrite in the tailings dam. Experiments were carried out with three different types of tailings from the plant and the pyrite concentrate with 46% Fe and 90% recovery has been obtained. Therefore, this study shows that high grade pyrite with high recovery can be recovered from the tailings.

Sep 1, 2012

By Yessengaziyev Azamat, Almagul Ultarakova, Onayev Murat, Barkytova Botakoz

"The research is aimed at solving the processing of waste products of titanium-magnesium production. The aim of the research is to extract synthetic carnallite and industrial products enriched with niobium from waste products of titanium-magnesium production. Analysis of the current state of the titanium and magnesium production in the Republic of Kazakhstan shows the availability of significant amounts of chloride wastes containing valuable components. Magnesium and potassium with an average content of 50-60 % potassium chlorides, 25-30 % magnesium chlorides are lost in the wastes of the titanium and magnesium production. Ust-Kamenogorsk Titanium Magnesium Plant JSC (UKTMP) currently imports carnallite that is a KCl-MgCl2X6H2O binary salt at quite high prices. Chloride wastes of titanium production, sublimates of dust chambers of a titanium chlorinator, sludge of a magnesium electrolytic unit and sludge of a carnallite chlorinator has been chosen as the targets of our research. The analysis of niobium distribution within any solid waste of a titanium slag chlorination process in UKTMP JSC showed that the highest content of niobium is in the sublimates of dust chambers of a titanium chlorinator in the ranges from 0.26 to 1.0 %. Among the niobium extraction technologies considered, the known method of its extraction from the chloride waste is leaching with weak solutions of mineral acids or alkalis in order to convert the chloride salts into the solution to be used for synthetic carnallite extraction. Besides niobium, other impurity components are transferred to the cake and shall be separated from the base metals. The effective method for this is the chlorination process enabling niobium transfer into the sublimates in the form of chlorides and thus separating it from the majority of the impurities. The condensed niobium chlorides are hydrolyzed transferring it into sediments in the form of oxide. New ways of processing waste products of titanium-magnesium production with the receipt of marketable products are relevant in the development of processing of mineral raw materials."

Jan 1, 2018

By Carlos Rabanal, Levi Guzmán

"The concern regarding grinding media consumption rates is as old as the invention of the tumbling mill. Over the years many materials and grinding media shapes have been tested. Through all this, steel grinding balls have proved to be the most effective media for comminution in tumbling mills. There have been several approaches to be able to the wear of the ball charge in any given mill, all of them based on limited empirical evidence. The mining industry still utilizes the Bond Abrasion test that was developed in the 1960´s. As shown later in this publication, this approach shows error greater than 60%. More recently, in 2007, Radziszewski from Mc Gill University has proposed a different total grinding media wear model; this model is based on decoupling the effects of the abrasion, corrosion and impact wear mechanisms. Also in 2007, Benavente, at the time with Moly-Cop Peru, presented an empirical model relating both ore properties and operational conditions that would affect the wear performance of the grinding media. For the database referred in this article, Benavente’s model shows an average error of ± 12% which represents a higher degree of improvement in comparison with previous approaches. The current paper presents recent research conducted by Moly-Cop Peru to further improve the capability of the empirical model originally developed by Benavente to estimate grinding media wear rates. The work performed included extensive determinations of the Bond Abrasion Index for different types of ores, collected at industrial grinding applications where the corresponding operational information was made available. The results of the research enabled us to build a more robust and confident model for estimating the grinding media wear rate. The average error of this new model is in the order of ± 9% which is a significant improvement over the models previously described."

Jan 1, 2014

By T. Tsurata

The accumulation of thorium and/or uranium by various microorganisms from a solution containing thorium and/or uranium at pH 3.5 was examined. Among the species tested, a high accumulation ability for thorium was exhibited by strains of gram-positive bacteria, such as Bacillus megaterium, B. subtilis, Micrococcus luteus, Rhodococcus erythropolis, and actinomycetes, such as Streptomyces levoris. On the other hand, though uranium was accumulated in small amounts from the solution containing both elements by all these microorganisms, actinomycetes, such as S. flavoviridis and S. levoris had relatively high uranium accumulation abilities. In these high performance thorium- and uranium- accumulating microorganisms, S.levoris, which accumulated the largest amount of uranium from the solution containing only uranium at pH 3.5, could accumulate about 300 µmol thorium and 133 µmol uranium per gram dry wt. of microbial cells from a solution containing both thorium and uranium at pH 3.5. After this screening, uranium accumulation using microorganisms isolated from uranium mines, among which some strains of bacteria were found to possess an extremely high uranium accumulating ability. The microbial biomass may thus be considered for use as an accumulating agent for the removal and recovery of uranium that may be present in metallurgical effluents, mine tailings, seawater and other sources. Recently, hundreds of micro-organisms existing in uranium deposits located in North America, Australia, and Japan for their ability to accumulate large amounts of uranium, and found some new strains which accumulated large amounts of uranium, such as Lactobacillus and Bacillus sp. from Japanese and Arthrobacter and Bacillus sp. from the United States uranium deposits. These bacterial strains can accumulate about 2500 µmol uranium per gram dry wt. of microbial cells within one hour. These microbial cells can remove uranium from the uranium refining wastewater with a high efficiency. The accumulation of uranium from a solution (pH 5.8) by microorganisms contained Arthrobacter nicotianee, which is a similar strain mentioned above Arthrobacter sp., was examined. Among the 76 strains of 69 species tested (23 bacteria, 20 actinomycetes, 18 fungi, and 15 yeasts), high uranium accumulation ability was exhibited by strains of the bacteria, A. nicotianae, B. subtilis, and M. luteus. A. nicotianae cells, which showed the best performance, could accumulate about 698 mg uranium (2.58 mmol) per gram dry wt. of microbial cells.

Jan 1, 2014

By S. Dwarapudi

Pelletizing of Indian chromite ores is more challenging due to their high refractory nature. High Cr/Fe ratio and high MgO content in these ores demand high firing temperatures and longer firing cycles but often result in low strength fired pellets. Aim of this study was to develop cold bonded chromite pellets for smelting in submerged arc furnace (SAF) from chromite fines using suitable binder that induce less gangue into the pellets but cures quickly. Different binders were studied through laboratory pelletizing experiments for their suitability for cold bonding the pellets. As result, a composite binder comprising organic and inorganic binding agents was found to be suitable and pellets made from the same were tested for their low and high temperature behavior. High temperature reduction studies revealed that pellets were resistant to disintegration up to 1200 °C.

Sep 1, 2012

By S. H. He

This paper focuses on the influence of pulp oxygen content on the separation of pentlandite and pyrrhotite contained in two different process streams in the Clarabelle Mill, operated by VALE INCO. The two process streams are (i) the scavenger cleaner feed, the product of regrinding the thickened, scavenged concentrate and the B cleaned tail produced from treating the non-magnetic fraction of the milled run of mine ore, and (ii) the magnetic concentrate which is the magnetic fraction of the milled run of mine ore. In the Clarabelle Mill, sulphite (SO32-) and TETA (triethylenete-tramine) are together used to selectively depress pyrrhotite from pentlandite. These depressants pre-ferentially decompose collector and/or sulphur enriched species at the pyrrhotite surface, an appar-ently essential mechanism in pyrrhotite depression from the scavenger cleaner feed stream. Selectivity in depression may originate from differences in the electrochemical rest potential of pyrrhotite and pentlandite. As both minerals are susceptible to oxidation, aeration in conditioning and flotation plays a key role in selective flotation. Important differences in the flotation behavior of pentlandite and pyrrhotite from the magnetic concentrate and the scavenger cleaner streams are highlighted in

Jan 1, 2014

By A. V. Bogdanovich, S. A. Urnysheva

"Gravity and X-ray luminescent separation methods are generally used in processing of diamond ores. X-ray luminescent separators are applied for separation of diamonds from feed ore size fractions from 100 to 10-15 mm size and for final treatment of gravity concentrates over 1 mm size. Gravity separation machines used for diamond ores separation are heavy medium cyclones and pneumatic jigs. One of the most important parameters, to a large extent affecting diamonds recovery from ore, as well as processing costs and specific energy costs, is size of dumped final tailings. That was exactly the purpose of the studies, the results of which are described in this paper. The studies performed on different ores showed that in order to achieve high recovery results, it is necessary to dump heavy medium separation tailings below 3 mm size. Separation of final tailings of -6+1 mm size by jigging is also possible. However, with a view to exclude extra diamond losses, it is necessary to maintain jigging concentrate yield approximately at 25% from starting feed, with its subsequent cleaning in heavy medium cyclones. In addition to that, it is necessary to separate 20-25% of middling with diamonds content at feed ore level and subject it to grinding. This will complicate flow sheet, but reduce load on heavy medium cyclones. In order to develop optimal flow sheet for gravity concentration of a specific diamond ore, large-scale investigations are required, including technical-and-economic assessment of separation alternatives. INTRODUCTION In processing of diamond ores, gravity and X-ray luminescent separation methods are generally used. It is common practice to apply X-ray luminescent separators for separation of diamonds from feed ore size fractions from 50 (100) to 20 (10) mm size and for final treatment of gravity concentrates over 1 mm size. With that, X-ray luminescent separation tailings larger than 3-6 mm size are reground and delivered to gravity concentration processing stage. Middling products circulation volumes are large, not infrequently necessitating decrease in concentrating plants capacity, especially with mining in deep workings of deposits, where ore strength is higher (Goryachev, 2010; Vaginov, 2000; William, 1988). Diamond ores preparation before processing is accomplished by means of roll-screw crushers and autogenous grinding mills, minimizing probability of free diamonds damage in liberation from barren rock. In re-crushing of size fractions that are “difficult” in autogenous grinding, high pressure grinding rollers are applied (Burchardt, 2014; Fedotov, 2011a, 2006b)."

Jan 1, 2018

By J. Quinteros-Riquelme

Despite increasing demand, the supply of silver is limited by its inadequate recovery from ores. Due to its scarcity in nature, its high ore variability and complex mineralogy of current silver deposits, its recovery by mineral froth flotation presents a real challenge for process engineers. A silver ore often contains significant amounts of five to ten different silver minerals and knowledge of the mineralogy is a necessary precursor to flotation testing. In this work, a complex silver ore was treated by using batch laboratory flotation, in an effort to find the best experimental conditions to obtain the optimal silver recovery. Mineralogical characterization of feed samples was conducted using a combination of tools, namely the mineral liberation analyzer (MLA) and inductively-coupled plasma mass spectrometry (LA-ICP-MS), which indicated that most of the silver occurred as ?invisible? silver in pyrite. The pyrite assay of the ore was approximately 4%. The ore was processed in two flotation stages. The first stage, sulphide flotation, aimed to collect a sulphide mineral rich concentrate; then a non-sulphide flotation stage was performed by adding sodium sulphide and collectors to the remaining material in the cell, with the objective of obtaining a pyrite rich concentrate. To reach the target recovery of silver from this ore, various variables were altered. In the first stage: mixed collectors, activators, dispersants, and mainstream inert grinding were assessed. In the second stage, the conditions for sulphidization and subsequent flotation were both investigated. From the grade recovery curves for silver and sulphur it was possible to observe a similar metallurgical behavior for both elements, supporting the results given by the mineralogical characterization. Keywords: silver oxide, froth flotation, sulphidization, mig

Sep 1, 2012

By S. M. Narbekova, L. V. Semushkina, Zh. A. Aldybayeva, A. A. Mukhanova, D. K. Turysbekov, N. K. Tussupbayev

More complex and heavy-concentrating oxidized and mixed sulphidic oxidized ores are involved in production because of increasing needs for non-ferrous metals. Extraction of the oxidized minerals of non-ferrous metals general in current technologies. The problem is solved to oxidized polymetallic ores the preliminary selective sulphidization is the oxidized surface of the components by means of hydrogen sulphide, sulphide and calcium polysulphide, and also element sulphur, etc. The task was to study details of flotation of the oxidized lead-bearing ore of the Zarechnoe deposit (Kazakhstan) with calcium polysulphide as sulphidation. According to chemical analysis initial samples of the ore contain, %: lead – 4.8; zinc – 0.1; general iron – 0.81. Phase structure lead is characterized by the oxidized form of 57.76-64.5 %; by sulphidic status of 25.5-32.64 %, and by plyumboyarosite of 19.0-23.02 %. Researches on flotation beneficiation of oxidized lead-bearing ore are conducted. Sulfidizator gave to grinding, on flotations there were used the collector (butyl xanthogenate of sodium) and frother (T-80). Experiments were conducted with a new sulphidizator – calcium polysulphide. The latter is obtained by interaction of CaO calcium oxide with sulfur and water at high temperature with component mass the following ratio S:CaO:H2O =10:5:85. Extent of mill initial ore was optimised (88.42 % of - 0.071 mm grade). For the optimum mill mode the studies on matching of an optimized consumption of sulphidizators are performed. On basic technology with application of sulphide of sodium as a sulphidizator at optimum mill, the lead concentrate with the content of lead 64.4 % at extraction of 86.6 % was received. The content of cerussite in a concentrate was about 14.9 % at extraction of 82.1 %. With application of calcium polysulphide as a sulphidizator the lead concentrate with the content of lead 68.0 % at extraction of 88.0 % was received. The content of cerussite has made 16.0 % at extraction of 85.0 %. Application of a new sulphidizator allows to increase extraction of lead in a concentrate from 64.4% to 68% and cerussite extraction – from 82.1 % to 85.0 %. The possibility of application of calcium polysulphide as a sulphidizator instead of conventional sulphurous sodium was established.

Jan 1, 2018

By Ashok Kumar, Subhadra Sen Dr. Rashmi Kumari, D. P. Chakraborty, Ajay kumar Gupta

The blast furnace coke quality depends on the characteristics of coal blend, pre-carbonization techniques adopted such as stamping, vibro-compaction etc., and coking conditions. Of the above, coal blend plays a significant role in the production of quality coke. Furthermore, the quality of the blend depends on the quality of individual coals and their interaction making up the blend. During transportation and stockpiling, coal is in contact with air for periods of time that sometimes may exceed several months. During this time, reaction with oxygen in the presence of water, sunlight and possibly elevated temperatures may take place. The resulting weathered coal suffers some alteration in its technological properties, which in turn affects coke quality for Blast Furnace operation adversely. In this work, two coals belong two prime hard coking and hard coking categories, namely Illawarra & Tahmoor, were selected and kept in a separate stock pile for a period of 120 days. These coals were characterized in terms of proximate, rheological & petrography properties at a regular frequency. This report summarizes the impact of weathering on the coking behaviour of these two coals in terms of properties as mentioned above. However, the rate of loss depends on coal rank and storage conditions. This has been seen through this study that the loss depends on coal rank and storage conditions. Deterioration of rheological and petrographic properties were seen higher for low-rank coals, like in this study Tahmoor, which is a low rank coal, compared to Illawarra showed higher degree of coking property degradation within the same time period under identical environmental conditions. Low-rank coal deteriorates at a faster rate because low-rank coal is dominated by free radicals that easily react with oxygen, resulting in the creation of various carbonyl groups and weathered vitrinite. Based on the results obtained in this study, size of stock piles for different qualities of coals to be maintained, at a specific time, can be determined and recommended, so that target coke quality is always achieved.

Jan 1, 2018

By E. V. Zelinskaya, A. E. Burdonov

"In the raw aluminum production processon electrolyzers with self-baking anodes, during technological operations, flow-regulated treatment and adjustment of the electrolyte chemical composition produced by diesel and floor-rail technology, alumina and fluoride salts are spilled through the ventilation grilles of the electrolysis body aeration channels, leakage in the equipment, machinery and motorized machinery construction elements. As a result of this, a part of the bus station and the level of ± 0 (the basement room with supporting curbstones and columns, on which electrolyzers are installed) are covered with technological raw material. This leads to an increase in the conductor (bus) resistance, contact connections deterioration, cell insulation failure and raw materials loss. For example, at the enterprise ""Rusal-Bratsk"", as a result of the activity, up to 420 tons per month of alumina-containing wastes that are unfit for use in the electrolysis process are formed due to the alumina substantial contamination with salts of iron and silicon, which significantly reduces the technical and economic parameters of the electrolysis process. The main scientific problem is the processing (cleaning) of alumina-containing waste having a complex mineralogical composition of the material, its contrast and structure, from oxides of iron and silicon, using ""dry"" separation methods. In the performing the work process, qualitative indicators of the influence of external factors (air humidity, precipitation, sunlight, temperature difference) on the physicochemical processes components interaction during the raw materials storage are established. On the basis of laboratory tests, it is established that alumina-containing raw materials can be separated and photometric and gravitational separation methods used. A mathematical model of the motion of particles of primary and prepared alumina-containing raw materials in a cascade-gravity classifier was developed. The criteria and factors characterizing the alumina-containing material, as well as influencing and determining the maximum material enrichment, are established."

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 I. N. Gridasov, T. V. Ignatova, E. V. Shelepov

"Mikhailovskoye Deposit is comprised of oxidized ferruginous quartzites among other minerals with the former widely spread across the ore field and mined along with unoxidized ferruginous quartzites. PJSC Mikhailovsky GOK strategic development is aimed at iron ore product output expansion. The most critical item of the strategic development road map is process design for separation and concentration of hematite contained in oxidized ferruginous quartzites mined as the by-product in the open pit. Laboratory and semi-industrial tests run to design cost-effective, reliable and environmentally friendly process for Mikhailovskoye deposit oxide quartzites processing included the following processes: - pilot tests of reverse anionic flotation which produced concentrate with Fe60.1-62.9% total iron content and 66-69%Fe recovery rate; - combined magnetic separation and flotation (Marubeni Company, 1982) including slime removal, reverse anionic and cationic flotation and high gradient magnetic separation resulting in 48.9% yield, Fe55-57.5%, up to Fe60.5% hematite concentration ( Fe64.1% being the best result achieved),76.2% recovery. - magnetic separation and flotation, including separation in low gradient (1400 Oe) and high gradient (up to 17000-18000 Oe) magnetic field. The results of these pilots tests were the following: 57-59%Fe concentrate, 70%Fe recovery (Mekhanobrchermet, MGOK, 1982-1988): - magnetic roasting in pilot plant of TSGOK (Krivoy Rog) which produced 61.1%Fe concentrate with 81.5% recovery. Major disadvantage of the method is high generation of dust equal to 20% of the initial feed. Based on third-party studies flotation of oxidized ferruginous quartzites process embracing best processing methods was designed by Metalloinvest Management Company LLC and PJSC Mikhailovsky GOK engineers. This paper discusses analysis of studies aimed at intensification of separation processes with application of mechanical slurry agitation (MSA), addition of alkali (NaOH) and elutriation of initial feed. The paper also provides a design solution to implement a lab-scale MSA plant. It includes feasibility evaluation of MSA inclusion in flotation of oxidized ferruginous quartzites process to enable higher yield and hematite recovery in resultant concentrate. Outcomes of the discussed work have significant value for prospective engineering of OFQ concentration process."

Jan 1, 2018

By N. I. Tyushkova

This is a brief survey of the technological solutions for deep cleaning of baddeleyite?s concen-trate. Baddeleyite is part of kamaphorite of composite apatite-magnetite ore in alkaline-ultrabasic mountain range of Kola Peninsula. Baddeleyite?s concentrate, which is being produced by Kovdorsky GOK (ore mining and processing enterprise) contains 97.5%~98.5% (Zr?2 + ?f?2) and also scandium oxide (0.05%). It is usually used without any chemical processing. World demand for zirconium oxide is constantly on the rise. Its consumption grows in ceramic industry and in high-tech spheres. That?s why the demand for its purity is very important nowadays (additional cleaning from phosphorus, silicon, iron, aluminium). The mechanism and kinetics of interaction of activated bad-deleyite with sulphuric acid has been studied as well. While applying method of mathematical mod-eling it was confirmed that the mechanical activation is effective for the intensification of sulphation process of baddeleyite?s concentrate with the aim of purifying it from dirt. The use of the calculated optimal regimes allows to reduce such technological operation factors as: temperature, the du-ration of process, the sulphuric acid consumption. It also allows to determine the conditions of anticipatory abstraction of scandium and main dirt in solution in comparison with zirconium oxide.

Jan 1, 2014

By Ingvar Ymén, Alexandra Lundmark

"The depressive effect of dichromate on galena was evaluated with chemical experiments and small-scale depression experiments at a pH of around 12. The depressive effect was measured with Dynamic Vapor Sorption (DVS), which measures the water uptake of a sample as a function of the % relative humidity. The depression is due to PbCrO4PbO, which in the presence of Ca2+ and SO4 2- forms a more or less amorphous solid solution (Pb,Ca)(CrO4,SO4)PbO, which clogs to the surfaces of the ore. Free Pb2+- ions are needed for depression to occur and at high pH Pb2+ - hydroxide complexes, supply enough Pb2+ - ions to precipitate (Pb,Ca)(CrO4,SO4)PbO. In air CO2, precipitates CaCO3 which slowly lowers pH, precipitating more (Pb,Ca)(CrO4,SO4)PbO and CaCO3 on galena. Similar experiments where performed with five alternative depressors whereof phosphate, which forms an amorphous precipitate of approximate composition Pb3Ca13(PO4)10SO4, gave the best results. Phosphate was tested in laboratory scale flotation experiments giving encouraging results.INTRODUCTIONWithin Boliden Mines, concentrating of copper and lead from complex sulfide ores is performed by bulk flotation, meaning that copper and lead are concentrated in the same product. Downstream, copper and lead are separated by flotation of chalcopyrite, and depression of galena. Boliden has for many decades used sodium dichromate to depress galena (reference 1). Dichromate is both toxic and carcinogenic and Boliden has for many years investigated several alternatives. So far, however, the best results have been obtained with dichromate. Since Boliden has tried to substitute dichromate for a long time without finding a process as effective, Boliden decided to clarify the function of dichromate in the depression of galena and on the basis of the outcome find alternative depressors with similar function. This study was given to RISE Bioscience and Materials / Surface, Process and Formulation as a consultancy assignment. This paper presents the outcome of the combined efforts of Boliden and RISE."

Jan 1, 2018