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By A. N. Lazaridi, S. I. Afanasenko, S. A. Safonov

"This scientific work presents the results of comparative tests of centrifugal concentrators with a vertical and horizontal axis of rotor spinning. The article also describes a qualitative analysis of the behavior of mineral particles in a centrifugal field with fluidization for concentrators with a horizontal axis. The analysis was aimed at studying the effect of the quantity of fluidization holes in the basket on the degree of magnetite recovery and the effect of the basket rotation speed. The experiment demonstrated that concentrators with a horizontal axis are not inferior in terms of the qualitative recovery indicators to concentrators with a vertical axis. The researches also provided facts about the wide practical application of concentrators with a horizontal rotor arrangement. INTRODUCTION Methods of gravity concentration, attractive for their ecological purity, see their revival. First of all, this is due to the invention of centrifugal concentrators with the fluidization of the layer of mineral particles. The creation of a new generation of gravitational centrifugal concentrators made it possible to achieve a tangible breakthrough toward the direction of reducing the grain size of mineral particles effectively divided by density, up to 10 µm at a high production capacity of the vessels reaching 500-600 tons per hour. Such separators enable dressing a crushed ore of flotation grade and in some cases replace the flotation process, especially in primary operations, in joint cycles. In the long term, this method of concentration should significantly expand the scope of application, replace partial flotation processing, ensure the introduction of environmentally friendly energy-saving technologies at ore-processing plants. Summarizing the international experience for the last 30 years, it can be confidently asserted that centrifugal methods of concentration have become widely used in processing schemes for the separation of finely dispersed products in the mining industry. An interest in the creation of installations using centrifugal acceleration for the separation of mineral particles by density has particularly grown. Dozens of centrifugal concentrator designs have been created and continue to be improved. Centrifugal concentrators have a number of obvious advantages, which are especially relevant today. Losses of fine gold during the process of mining are well known. Gold washers screw, spiral separators, jigging machines and sluices traditionally used in the production processes do not solve the problem of fine gold recovery. Table 1 shows the results of comparative tests of a centrifugal concentrator (ITOMAK) and other concentration equipment. According to the research conducted by the company GMRS, Krugersdorp, South Africa (2000), the initial material was gold ore crushed to 150 micrometers (90%)."

Jan 1, 2018

By M. Salas, L. Rolf, H. Z. Kuyumcu

"Industrial use of abrasive stress takes place in the area of fine grinding and in order to clean surface contaminated soil particles. Also, at many mechanical operations abrasion emerge as an unintended side effect. At the Institute for Raw and Waste Material Processing at the Technical University of Berlin systematic investigations are carried out, to further the knowledge regarding the dependence of abrasion process on material properties and stress mechanisms as a basis for design and optimisation of the process or for its avoidance. Aim was the mathematical description of the abrasion process running with a single particle stress. For the purpose of investigations an agitation cell was used. Limestone served as material.In the presented work the influence of particle diameter, particle shape and collision energy on the mass and particle size distribution of abrasion debris was studied. The surface geometry of the particle was characterised by the particle shape factor ?x and the particle diameter dp. The particle shape factor ?x stays in direct relation with the abrasion degree x, and can be substituted by a mathematical function of x.Physically the particle shape factor ?x can be interpreted as a proportion of abrasive stresses splitting up complete crystals and crystal fragments. These two different mechanisms depend on different ratios of curvature of the surface of the stressed particle. The abrasive stress of strongly curved surface areas leads to separation of complete crystals. The abrasive stress of weakly curved surface areas leads to separation of crystals fragments. Accordingly the mass of wear mAb and median diameter of the abrasion debris d50 are different. The derived mathematical relations consider this behaviour by the particle shape factor ?x , which represents the ratio of the mixture of both stress mechanisms. The dependence of the particle shape factor ?x on the abrasion degree x with an exponent approximately to 2/3 points out that the particle shape factor ?x represents a surface area relation."

Jan 1, 2003

By Jyotirmoy Sarma

The surface-active agent, like sodium humate, in the present systems produces a very stable suspension. Saline water up to 15% is used as a flocculating agent for recovery of water from the dispersed phase containing 1% solid. The recovered water in the presence of humate at different concentrations (0.05-0.2%) and at pH 8 also produces a stable dispersed phase of iron ore fines and tailings. The beneficiation techniques adopted in the present systems can increase ~ 2-3% iron value in the concentrate of iron ore fine, however, for tailings the increase in iron value is not observed. Keywords: beneficiation, flocculation, iron ore fines, tailings, saline water, surface-active agent

Sep 1, 2012

By P. V. Ramesh Babu

India has sizeable number of small deposits of rare metal bearing minerals namely columbite-tantalite, pyrochlore, beryl, lepidolite, spodumene and rare-earth element bearing minerals monazite and xenotime in different parts of India. Pegmatites, alkali granites, syenites, and the derived gravel and soils are the main source for rare metal and rare earth (RMRE) minerals. The RMRE minerals are classified as prescribed substances under Atomic Energy act of 1962 keeping in view the uses in the nuclear industry. Atomic Minerals Directorate for Exploration and Research (AMD) of Department of Atomic Energy (DAE) is the sole agency responsible for exploration and exploitation of RMRE minerals in India since 1962. The latest technological advancements and applications in the field of space, electronics, communication and nuclear industry using these metals and their products have necessitated to augment the production and resources of RMRE minerals. These minerals play a significant role in the technological and economic development of the country. The light rare earth element bearing mineral monazite is abundant in beach sands and is being commercially exploited by Indian Rare Earths Limited (IREL). Columbite-tantalite, the chief source for niobium and tantalum (Nb-Ta) metals present in the pegmatitic gravel are recovered by small mobile beneficiation plants. The cut and fill method of mining adapted is eco-friendly. The processing flow-sheet consists of wet sizing at ½ in and 1mm followed by gravity concentration of sized products utilizing the methods of jigging and wet tabling. The sizing equipment include revolving and vibratory screens and the gravity concentration equipments used are fixed sieve jigs and movable screen jigs for final upgradation. The gravity concentrates are further upgraded by magnetic separation to remove magnetite and ilmenite. The mine areas are reclaimed by filling with coarse size tailings and new plantations. These mobile pilot plants were earlier operated in Bihar mica belt, Jharkhand, pegmatitic areas of Karnataka and Nellore mica belt of Andhra Pradesh and are currently being operated in Bastar-Malkangiri pegmatite belt of Chhattisgarh and pegmatites of Jharsuguda district of Orissa. Beryl, lepidolite, spodumene and cassiterite minerals are the by-products from the pegmatite ores. Fergusonite, samarskite and microlite are the other niobium and tantalum bearing minerals which are also found in pegmatites. Xenotime, the chief mineral of yttirium generally occurs in granites and pegmatites. The rare earth minerals monazite and xenotime get concentrated in the placer sands of rivers and beaches. The occurrence of xenotime in the sands of Siri River and Deo River in Chhotanagpur granitic terrain in parts of Chhattisgarh and Jharkhand are an example. These sands containing 1-2.5% heavy minerals were processed by a mobile plant consisting of sizing equipments like trommel and vibratory screen followed by separation of heavy minerals from -1mm sands with the help of gravity concentration equipments like spirals/shaking tables/sluices. Xenotime and monazite from the pre-concentrate are further separated by magnetic and electro-static methods. The present paper illustrates the pilot plant set up for the recovery of rare metal as well as rare earth minerals. Keywords: small deposits, rare earths, rare metals, pegmatite, Nb-Ta, xenotime, India

Sep 1, 2012

By Tsvetan Damyanov, Nikolay Hristov, Evgeni Kostadinov, Lubomir Kuzev

"The process of grinding in drum mills is directly dependent on the dynamic forces occuring in the free movement of the grinding bodies in the volume. The subject of discussion are the potential impacts between grinding bodies having different shape and equal weight in a drum mill that has a diameter of 1200 mm. The study utilised ball grinding bodies having a diameter of 120 mm and tetraballpebs having a conditional diameter of 120 mm. The study was carried out using the Solid Works Simulation software. The following cases were examined: • Impact between two identical grinding bodies; • Impact between a grinding body and the shell surface. The study provided numerical values for impact force, kinetic energy of impact, linear speed of impact, linear displacment of the center of gravity and duration of impact deformation. The impact force of the balls is approximately 179000 N. The impact force of the tetraballpebs with typical vertices, edges and spherical surfaces depending on the area of impact is 177000 to 185000 N. Upon impact between two spherical surfaces of tetraballpebs, the contact area is 263.55 mm2, which is more than twice the contact area upon impact between two balls, which is 114.39 mm2. This determines better conditions for the transmission of kinetic energy by reducing opportunities for overgrinding."

Jan 1, 2014

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 Zhou Yucai, Zhou Jing, Zhu Yimim

"In this article, research on the beneficiation process for a carbon containing copper-molybdenum deposit in Henan province has been carried out. The elemental composition, mineral composition and relative content, characteristics of embedded and symbiotic relationship, the forms of the copper-molybdenum particlesand their disseminated features, and particle size distribution have been worked out. Flotation behavior of copper-molybdenum mixed flotation and separate flotation in the systems of different collectors and depressants, and the influence of different flowsheets on the copper-molybdenum separation have been investigated by actual ore flotation tests, The separation problem of copper-molybdenum for a carbon-containing copper molybdenum ore has been solved.INTRODUCTION In this paper, technological mineralogy, small experiment of flotation recovery of copper Molybdenum ore, and expansion experiment of mineral recovery of copper Molybdenum ore are introduced. (1) The results of process mineralogyshow that the content of copper is 0.42%, molybdenum is 0.021% in the raw ore, they mainly exist in the form of sulfide, with low oxidation rate. (2) The results of flotation tests show that the combination of butyl xanthate and isopropyl ethylthionocarbamate reagents is effective collector of copper-molybdenum mixed flotation for carbon copper molybdenum, (3) When the grinding fineness is -0.074 mm taking 71%, and regrinding fineness is -0.038 mm taking 85%, the flowsheet of first copper-molybdenum mixed flotation and then regrinding rough concentrate and last separation flotation of copper and molybdenum was adopted, which realized the effective separation of copper and molybdenum for carbon-containing copper molybdenum ore. Good flotation indexes were gained with copper concentrate containing 20.96% copper and copper recovery 90.06%, and molybdenum concentrate containing 46.78% molybdenum with molybdenum recovery 72.68%. (4) The seven days average concentration index of scale-up the continuous dressing experiment is as follows, the copper content in copper concentrate is 20.16% and the recovery is 89.61%; the molybdenum content in molybdenum concentrate is 24.58% and the recovery rate is 74.09%."

Jan 1, 2018

Domestic production of pyrite cinder reached 15.95 million tons in 2006. The pyrite cinder used in cement industry accounts for about 30% production. The remaining 70%, 11.17 million tons, was thrown away as waste, which brought serious environmental and resource problems. A new technology of composite pellet reduction-roasting and magnetic-separation (CPRM) has been developed to utilize pyrite cinder. Pyrite cinder sample, containing 45.56% Fe, 0.19% Cu and 1.45% S, was pelletized with addition of 5% composite agent in balling disc. Green balls after drying were reduced at 700~800 ? for 20min with 300kg/t pulverized coal injection in rotary kiln. After two stages of ball-grinding and four stages of magnetic separation, high grade magnetite concentrate with 66.23% Fe was produced, meanwhile impurities Cu and S was decreased to 0.048% and 0.053%, respectively. And iron recovery of 73.32%, desulphurization of 96.34% and decoppering degree of 74.73% were achieved. Key technique is that composite agent can accelerate reduction of Fe2O3 and transfer copper and sulphur into water-soluble inorganic salts, which can be removed during wet magnetic concentration. This new technology can be used to fully utilize pyrite cinder and produce high quality magnetite concentrates as feed for steel industry, which will help tackle the pollution problem of pyrite cinder and extend iron-bearing raw material sourcing for Chinese steel industry.

Jan 1, 2014

By Qingming Feng, Dongbo An, Jinhong Zhang

The application of the geopolymerization of mine tailings as construction materials is hindered by the fact that conventional geopolymerization of mine tailings usually involves a large consumption of alkali, i.e., sodium hydroxide (NaOH), and a long curing time, because of the relatively low reactivity and the high Si/Al ratio of mine tailings when compared to other active resources such as fly ash and clay minerals. The present study is focused on efficiently activating mine tailings, reducing alkali consumption, decreasing curing time and improving compressive strength. We firstly investigate the temperature effect on the alkali activation of mine tailings during geopolymerization. Secondly, the impact of additives, i.e., calcium hydroxide and aluminium oxide, on the compressive strength of samples is studied. Thirdly, different forming pressures have been applied to remove excessive alkali and its impact on samples' strength has been investigated. The results show that a 40 MPa unconfined compressive strength (UCS) can be achieved with the geopolymerization samples after mine tailings are activated by sodium hydroxide at 170°C for 1 hour with the addition of calcium hydroxide and alkali dissolved aluminium oxide, further compressed with a 10 MPa forming pressure and finally cured at 90°C for 3 days. The findings of the present work provide a novel method for the geopolymerization of mine tailings as construction materials.

Jan 1, 2014

By Hakan Gürsoy, Kemal Bilir, Bahri Öteyaka, Ali Uçar, Oktay Sahbaz

"Jet diffuser flotation column (JDFC) is a flotation device resembles Jameson flotation cell (JFC) in terms of operational principles, but it has significant difference structural points of view. This difference is an essential change in the downcomer to provide an increase of performance of the cell for the coarse particles. By using the hydrodynamic and fluid mechanics principles (Sahbaz et al., 2013) downcomer was changed as a diffuser, therefore turbulence at the end of the downcomer was reduced to provide better aggregate stability and homogeneous dissipation of the bubbles in the separation tank. According to experimental results carried out in pilot scale JDFC, not only higher flotation performance comparing with Jameson cell was obtained specifically for coarse talc particles in a certain condition, but also quiescent froth layer was acquired. In these tests, vertical pipe of JDFC having inlet diameter of 6 cm and outlet diameter of 11.5 cm was used with separation tank having the diameter of 39 cm. By using the data, pilot scale JDFC having 410 cm vertical pipe integrated with separation tank was produced and flotation tests were carried out by using the naturally hydrophobic mineral (talc). Finally, the talc recovery obtained as 94% by the use of JDFC, while Jameson cell has the recovery of 77% for the particle size of 250 µm."

Jan 1, 2014

By K. Haga

In this study, flotation tests were performed to investigate behavior of enargite and chalcopyrite ores from different sources synthetically mixed for separation of enargite. Flotation parameters including; pH, reagents concentration and type and flotation time were controlled to obtain an enargite concentrate with high As and chalcopyrite concentrate and final tailings with As content below 0.5 mass%. The copper ore bearing enargite mineral contained 5.87 mass% As and 16.50 mass% Cu while the chalcopyrite bearing ore contained 0.32 mass% As and 21.63 mass% Cu. The two ore samples were mixed at 7:3 (enargite: chalcopyrite) by weight ratio to prepare a mixed ore sample with arsenic content at 3.16 mass% and 18.25 mass% Cu for the flotation study. For enargite single ore flotation, the 38?75 µm size fraction showed that over 98% enargite was selectively recovered within 5 min at slurry pH of 4.0 and arsenic content in the final tailings was reduced to 0.22 mass%. In mix ore (enargite+chalcopyrite) flotation, 97% of enargite was first removed at pH 4 followed by chalcopyrite flotation at pH 8, and over 95% recovery was achieved in 15 min flotation time. Arsenic content in the final tailings was reduced to 0.1 mass%.

Sep 1, 2012

By S. N. Shahcheraghi

Grinding is one of the most important and expensive steps in mineral processing allocating about 70% of the total energy in the plants. Nowadays, for optimizing the energy consumption in these equipments, the rock behavior and its size distribution during the crushing have been normally predicted using selection and breakage functions to be able to optimally model and design grinding circuits. For this purpose, the case of Miduk copper mine in Kerman province, Iran was studied. The samples were taken from primary ball mills feed and were subject to screen analysis to achieve uniform sizes. Then, selection and breakage functions were determined at the different grinding times in laboratorial ball mill. Finally, ore size distribution in different grinding times was estimated using MATLAB software

Sep 1, 2012

By Christopher Davey, Marian Mazurkiewicz, Yvonne Spooner, Dana Hankey, Peter Kanjorski

The concept of a high pressure liquid jet mill with highly unsteady flow, has been developed and recently patented by Cornerstone Technologies LLC (CsT), for the purpose of micron and sub-micron particle production. The mill accepts feed sizes up to 15mm for some materials and is capable of submicron production. Successful material range from precious metals such as silver and palladium to dielectric components and substrates such as Barium Titanate slips, zirconia, titania and silica. The process has also succeeded in adequately reducing the size of materials for both composite and coating uses such as anthracite coal, graphite, chrome metal, wollastonite and more than 100 others. A unique design approach has been taken to highly maximize the energy flux in the mill's chambers to create extremely high turbulence and destructive cavitation. This is assisted by shock waves, which cause particles to undergo ultra high frequency pressure pulsations, high frequency collisions and then abrasion, shattering, hydrowedging, and cleaving. Due to these extreme properties the processing occurs in just fractions of a second allowing throughputs up to 1000 kg/h on a machine that would fit inside a 60 cm box. The technical solution and technological advance, as well as the environmental and economic benefits indicate rapid and wide spread acceptance of high-pressure liquid jet milling in modern industry.

Jan 1, 2003

By Rogerio C. Guimarães, Antonio E. C. Peres

The production of phosphate concentrates (phosphate rock) from slimes represents a challenge to the phosphate industry worldwide. In Brazil, this challenge is particularly relevant due to the consistently low P2O5 grades and the igneous origin. The Brazilian phosphate deposits are made up mainly of apatite from alkaline chimneys, representing 80% of the Brazilian reserves; phosphorite deposits contribute 18% and the other sources (guano, phosphorous bauxite, and apatite from veins) reach 2%. The Brazilian industry beneficiates phosphate ores that are mineralogically more complex than those treated elsewhere. The low apatite grade, inherent to the igneous Brazilian ores, added to the mineralogical complexity due to the weathering of the rocks, has represented a permanent challenge to the Brazilian process engineers. The total measured, indicated and inferred reserves reach 3.2 million tonnes of ore with average grade 11% P2O5. The development of local technology for phosphate beneficiation started in the 1960’s. The first industrial experiences of producing phosphate rock from slimes (-44µm or -325#) fraction started at Arafertil, currently Bunge Fertilizantes SA, in Araxá, in the 1980’s. This production corresponds, nowadays, to 35% of the total production of concentrates in the company (830,000t/year) in a circuit consisting of desliming in 40mm microcyclones and concentration in flotation columns, corn starch and fatty acids being utilised as reagents. The performance of slimes flotation and the process control, at Bunge´s concentrator, were improved after the introduction of column flotation technology, in the first half of the nineties. The column flotation technology for slimes recovery was expanded to other Brazilian concentrators still in the nineties: Ultrafértil (Catalão – GO), Bunge (Cajati – SP), Fosfértil (Tapira – MG) and Copebrás (Catalão – GO). At present the Brazilian production of phosphate rock is approximately 4.5Mt/year, with a strong trend for a market increase. Genuinely Brazilian technology is utilised for the processing of slimes, the concentrates representing 11 to 13% of the overall production. The production of apatite concentrates from slimes was relevant to increase phosphate recovery and reduce costs, contributing to the growth of the fertilisers industry.

Jan 1, 2003

By I. N. Gridasov, V. V. Khromov, R. I. Ismagilov

"Over 50 mln.tonne is the average annual output of oxidized ferruginous quartzites typically containing Fe38-42% mined in Mikhailovskoye ore field. The most promising and cost-effective way to expand concentrate production maintaining current output of magnetite is processing of oxidized ferruginous quartzites (hereinafter referred to as OFQ) mined as a by-product in Mikhailovsky open pit mine. However lack of proven and efficient concentration and agglomeration process applicable to OFQ and insufficient data for feasibility evaluation of OFQ processing encumbered full-scale addressing of this challenge. Based on outcomes of multiple various studies a process of OFQ flotation was designed by MC Metalloinvest LLC and PJSC Mikhailovsky GOK engineers. Comparative lab-scale testwork have proven performance and production advantages of flotation against magnetic separation feat. flotation and combined concentration method. Lab-scale flotation enabled production of concentrate with average Fe66.6% and SiO22.06% grade. Geologic and production contouring of OFQ deposited in Mikhailovskoye ore field has been carried out to collect necessary data for feasibility study incl. cost and profit estimation. As a result processing and production properties of over 3,500 samples have been analyzed which is characteristic of every direction and level of prospective ore mining within Mikhailovskoye ore field up to year 2032. Following work has been done based on geologic and production contouring of OFQ: – natural typification of ores and production (processing) grades of OFQ, determination of average mineral composition; – correlation analysis of interrelation between OFQ mineral composition and processing features thereof for overall bulk of samples as well as for individual data blocks related to various directions of open pit mine; – OFQ classification with account of defined mineral composition and processing features. Total Fe weight content in chamber product of flotation (known as concentrating ability in flotation) was applied as a major differentiating factor for processing grade classification of OFQ. Three processing grades of oxidized ferruginous quartzites were established as a result of flotation tests run to determine concentrating ability, i.e. – hard concentrated OFQ (TO) – below Fe65%; – mean concentrated OFQ (SO) – Fe65-67%; – easy concentrated OFQ (LO) – over Fe67%. Outcome of this work provided initial data for feasibility study and cost-profit evaluation of medium-term industrial-scale processing of OFQ deposited in Mikhailovskoye ore field."

Jan 1, 2018

By H. Koca, D. O. Aksoy, S. Koca

"In this work, modelling of galena flotation from a disseminated Pb-Zn ore containing 0,83 % Pb was investigated. In the flotation experiments, collector type and amount, frother type and amount, particle size, pH and flotation time were kept constant while solid ratio and depressant dosage were varied. The varied parameters are described with a mathematical model by using statistical design method. Central Composite Design method which is one of the principle response surface methodologies was used in statistical design and analysis of experimental results and process modelling. Two quadratic mathematical models were developed for Pb grade and Pb recovery. Analysis of variance showed that models obtained for Pb grade and Pb recovery were found to be statistically significant in 95 % confidence level. Predicted vs actual values of Pb grade and Pb recovery were also given in graphically. It was found that predicted values represent approximately 96 % and 94 % of actual Pb grade and Pb recovery, respectively. The high values of R-square (R2) indicate that the models developed for two response variables are powerful. INTRODUCTION Lead and zinc metals are widely used in electrical, mechanical, military, metallurgical, chemical, light and medical industries. Sulfide ores are the main resources for lead and zinc minerals (Luo et al., 2016). Froth flotation is the main beneficiation method to achieve lead and zinc sulfide minerals separation. There are two basic approaches to obtain lead and zinc separation by flotation. First, flotation of lead sulphide is carried out and zinc sulphide and gangue minerals are depressed by various depressants and second, the bulk flotation of lead and zinc sulphide is carried out to obtain a concentrate of lead and zinc, followed by the lead and zinc separation. (Yuan et al., 2012). Apart from concentration of lead and zinc by flotation, there are a few literature that reports concentrations of lead and zinc by gravity and hydrometallurgical methods with remarkable success (Goktepe, 2005; Yanbo et al., 2016)."

Jan 1, 2018

By S. L. Gao

Based on the differences in hardness and density of diaspore and aluminosilicate minerals which usually exist in Chinese diasporic bauxite, It is analysed that bauxite of this kind is difficult to be separated effectively by classification and normal gravity separation, and cyclonic centrifugal field was thus proposed to be carried out, in which classification can take place according to the size difference resulted from selective grinding, what?s more, the separation function according to density difference is expected to take effect at the same time, and then cyclonic separation tests for diasporic bauxite of A/S 4.39 and grinding 61.90% -0.074mm were carried out tentatively. The results showed that the accretion of feed pressure can offer strong enough centrifugal force for cyclonic separation with result that separation function according to density difference is effectively intensified, by the increasing of this parameter, underflow and overflow of A/S 7.64 and 2.18 can be processed at the same time, and the recovery of Al2O3 in underflow is 71.78%; judging by the flow ratio, the spatial locality envelope of zero vertical velocity do constringe to the vertical axis by the enlarging of diameter ratio of apex to vortex, overflow and underflow of A/S 2.08 and 6.5~7.0 can be processed by controlling this structural parameter, which is possible to offer a technological method for low-grade bauxite to de-sliming beforehand.

Jan 1, 2014

By Naoki Hiroyoshi, Hajime Miki, Tsuyoshi Hirajima, Masami Tsunekawa

"Acidic solutions for chalcopyrite leaching contain cupric, ferrous and ferric ions, and the oxidizing power of the solutions is indicated by a redox potential, determined by the concentration ratio of ferric to ferrous ions. This paper reviews our previous studies on the redox potential dependence of chalcopyrite leaching in sulfuric acid solutions. Leaching experiments indicated that copper extraction rate was accelerated by cupric and ferrous ions , and was higher at potentials below a critical value (critical potential) in the presence of these ions. This was in good agreement with the results of electrochemical measurements, i.e. in the anodic polarization curves for a chalcopyrite electrode, active-passive behavior was observed when cupric and ferrous ions co-exist, but the activation at low potential did not occur without these ions (Figure 1). The analysis of the AC impedance spectra for chalcopyrite electrode indicated that a high-resistance passive layer grows on the chalcopyrite surface without cupric and/or ferrous ions, and that coexistence of these ions causes a formation of another product layer and inhibits the passive layer growth at low potentials.These results are explained by a reaction model assuming the formation of intermediate Cu2S due to chalcopyrite reduction and the subsequent oxidation of the Cu2S. The critical potential corresponds to the redox potential of the Cu2S formation and the predicted values matched with the experimental results. Catalytic effect of silver ions on chalcopyrite leaching was also discussed based on experimental results and the above model."

Jan 1, 2003

By Urs Peuker, Thomas Mütze, Thomas Leißner, Inga Osbahr

"Mining in the Ore Mountains (Germany) has taken place for hundreds of years and in several stages. Each stage produced its typical tailing. According to the former state of the art of the processing much valuables could not be extracted. This was caused by complex interlocking of the minerals, low grades, or by upgrading processes not sensitive enough for small particle sizes. Thus the processing of these fractions was uneconomically. Modern technical possibilities and rising metal prices bring these tailings disposals into focus for the processing of strategic metals. Mining disposals can be classified in heaps with coarse, low grade rock, which did not undergo a processing, slags and fly ash from smelting, and tailing disposals from mineral processing. The latter is under investigation to integrate them into a geographic information system. The mineral processing of the material will be part of the database besides information about geographic location, structure, amount of valuables, and details of ownership of the heap. The tailings disposals are composed of coarse, low grade, interlocked particles from density separation processes, medium size particles, which are tailings from flotation, and fines with liberated valuables, which did not undergo flotation. Therefore mineral processing of these disposals comprises the separation of liberated gangue by classification, beneficiation of locked valuables with density separation, liberation of valuables by grinding, and beneficiation of liberated valuables with flotation. The presented work shows the results concerning the composition of one exemplary heap and results of the classification as well as of the preconcentration tests achieved with material from this heap."

Jan 1, 2014

By Tatyana A. Ivanova, Tamara N. Matveyeva, Anatoliy A. Fedorov, Tatyana V. Nedosekina, Valentin A. Chanturiya

"As native gold resources are depleted, gold-sulfide ores consisting essentially of pyrite, arsenopyrite, chalcopyrite and pyrrhotite become one of the major precious metal sources, being presently estimated to account for more than 40 % of total world gold reserves.The prime objective of flotation of gold-bearing sulfide minerals is to increase precious metal recovery and to reduce the product cost through achieving maximum selectivity of the separation process. Separation of Fe and As sulfides in selective flotation of gold-bearing ore products helps reduce the amount of ore mass being sent to hydrometallurgical processing, and permits of recovering gold from As-free gold-containing pyrite concentrate by using a lower-cost pyrometallurgical procedure.An ingenious methodology purposively developed by the authors for comprehensive assessment of flotation properties, electrophysical characteristics and crystal chemistry features of natural gold-bearing sulfide minerals of different origin has been used in this study.Comparison of the elemental composition, non-stoichiometry and flotation activity of different monomineralic pyrite varieties revealed that samples with negligible impurity content are nearly stoichiometric, and show similar flotation properties. Samples with high Cu, As and Au contents are highly non-stoichiometric and readily floatable, even in a highly alkaline environment. Samples with high surface sulfur content and the most defective crystal lattices, with iron atoms substituting for sulfide vacancies, are most sensitive to depressive action of the alkali.A correlation between the type of conductivity, impurity content and amount of xanthate compounds on the surface and in the liquid phase of monomineralic pyrite and arsenopyrite suspensions detected by the extractive spectrophotometric analysis has been identified and analyzed. Experiments showed that the proportion of cuprous xanthate in sorptive collector layers on sample surfaces increases with increasing Cu content of the samples. In general, pH for pH, overall xanthate sorption on pyrite is greater than on arsenopyrite, with the difference increasing in passing from an acid environment to an alkaline one. A significant decrease in xanthate sorption on arsenopyrite is observed at pH greater than 9, whereas sorption on pyrite markedly decreases at pH greater than 11."

Jan 1, 2003