Search Documents

By Y. S. Song

Flotation separation of galena, sphalerite with pyrrhotite in ethanoic acid + sodium acetate sys-tem was studied in this paper. The flotation tests show that galena and sphalerite can float but pyr-rhotite can hardly float in the pH range of 3.6 to 5.6 in ethanoic acid + sodium acetate system. The flotation tests also show that galena can be separated efficiently from pyrrhotite and sphalerite can be separated on the whole from pyrrhotite, but the separation efficiency of the former is much better than the latter. The flotation results have been investigated by ultraviolet spectrum and XPS analysis, and explained by theory of hard and soft acid and base (HSAB), hydrogen bond and galvanic coupling.

Jan 1, 2014

By Chandra Sekhar Gahan

The present study aimed to investigate the depyritization of coal using iron oxidizing and sulphur oxidizing acidophilic chemolithotrophic microorganisms in an iron free 9K growth medium at varying pulp densities. Three different coal samples used in the shake flask study from three different countries such as domestic anthracite (Korean), Indonesian and Chinese lignite. Experiments were carried out in batch at a controlled temperature of 35 °C and varied pulp densities of 8%, 10% and 12% (w/v) with a working volume of 100 ml containing 10 ml of microbial inoculum and 90 ml of the iron free 9K growth medium. Coal samples were crushed and ground to fine size fractions <100 microns by attrition mill wet grinding. The progress in the experiments were followed by the regular measurements of pH, redox potential, viable planktonic cell count, residual Fe2+ ion concentration, amount of acid addition and amount of slaked lime addition. The pyritic sulfur removal percentage varied between 90-95% with highest sulfur removal with China coal (95%) followed by Korean Coal (93-94%) and slightly lower level for Indonesia coal (90%). However the pyrite oxidation was highest for Korean coal ranging between 78-80% followed by a similar level of pyrite oxidation of 50-61% for the Chinese and Indonesian coal. It was found that 10-12% pulp density would be fairly feasible for all the different coals tested leading to an economical process. Another benefit from this study is the iron free 9K growth medium, which will reduce the operation cost and overall it?s an environment friendly process.

Sep 1, 2012

By ?. S. Potokin

"In general the electric pulse treatment is composed from two processes – electric pulse grinding and electric explosive (electric discharging) activation. The electric pulse grinding with the effect of selective destruction refers to a major slag fraction, more that 2-5 mm, when a one can see a physical prerequisites for electrical breakdown of the slag particles. The electric pulse treatment of slag is considered as an additional operation in the technological scheme extraction of non-ferrous metals from copper smelting slags in addition with mechanical grinding, flotation enrichment and acid leaching. Studies of the reprocessability of pelletized slag have shown the possibility of higher extracting non-ferrous metals via flotation after electric pulse disintegration (EPD) treatment. By using EPD of slag pellets, it is possible to increase the non-ferrous metal yield to concentrate by 18-24%.INTRODUCTION The use of the explosive effect of the plasma discharge channel at electrical breakdown of a solid at a pulsed high voltage is the physical basis of the electric pulse method of destruction of geomaterials (Vorobiev et al., 1971; Andres, 1995; Semkin et al., 1995). The flowing of a pulsed current in the discharge channel is accompanied by a sharp increasing of plasma’s temperature and pressure, leading to a supersonic expanding of the discharge channel, the formation in the solid of intensive mechanical tension field, spreading of cracks, and eventually, to the destruction of the solid. Fragmentation and grinding of ore and mineral aggregates by electric pulse method due to a number of specific physical mechanisms of power load and crack formation, belonging only to this method, is different by high selectivity of destruction. This ensures better liberation of grains of individual minerals and their better preservation from destruction and less formation of a slurry output. These features creates the prerequisites for a more complete extraction of a useful mineral in the concentrate during ore dressing, to improve the quality of concentrate. In contradistinction to mechanical methods, electric pulse technology does not contaminate the product with instrumental metal, material of lining and grinding bodies. The possibility simply and in a wide range to change the dynamic load character allows to regulate the granulometric composition of the product of grinding, to obtain a narrower class of product size than is possible with mechanical grinding. The product of electric pulse grinding is characterized by less roundness and large specific surface area of grains due to the trait, that a new surface is formed in the process of splitting during the expanding of cracks. The degradation product has increased sorption and reactivity in chemical reactions. Liberation of the mineral inclusions particle better corresponds to their initial natural state in the rock and this is the great importance for the study of mineral raw materials."

Jan 1, 2018

By Tiago de Souza, Caroline Gonçalves, Carlos Eduardo de Castro, Everton Dias

The Miraí plant from Votorantim Metais, has two screening stages to the bauxite classification after scrubbagem. The final product has hold the larger size of 0,84 millimeters (20 mesh). There are two sieves Haver & Boecker (model R-TE 3658 x 10972), which have four modules with cut size sieves of 10 mesh (1.70 millimeters) and ¼ inch (6.35 millimeters) on the primary screen, and 20 mesh (0.84 mm) on the secondary screen, all made of polyurethane. The actual configuration was established according to the empirical tests realized that aimed the increase material load on the secondary screening. However, having in mind the screening process improvement and use the fractions smaller than 0,84 mm, which have significant levels of Alumina and low reactive silica, was proposed a new test of change the sieves opening for 6 mesh in the primary screen and 28 mesh for the secondary sieve. The objective of this work was measure the ore alterations on the particle size distribution on the screening steps, increasing the plant weight and ratio recovery before and after the test.

Jan 1, 2014

By Mohammad Noaparast, Hamed Haghi

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

Jan 1, 2014

By Kevin Galvin, James Dickinson

"The Reflux Flotation Cell (RFC) consists of an inverted fluidized bubbly bed, located above a system of parallel inclined channels. This novel arrangement enhances the hydrodynamics of flotation in a number of ways. Firstly, the free-surface of the flotation cell is enclosed via a fluidized bed distributor, ensuring uniform application of wash water to promote desliming. Secondly, foams can be variable in their stability and lack permeability. Hence this system is preferably operated in the absence of foam, instead opting for a bubbly zone of high bubble volume fraction through the use of a very high gas flux and/or wash water flux. Discharge of the product proceeds upwards through a central port of narrow cross-sectional area. Thirdly, the system of inclined channels increases the segregation between the bubbles and the water, ensuring the bubbly zone remains contained within the upper vertical zone, and the liquid flux reporting to the overflow is low. The expanded operating regime of the RFC was investigated by covering a very broad range of gas fluxes, up to 5.5 cm/s, and wash water fluxes, over 2 cm/s. Both laboratory and pilot plant performances were quantified by the recovery of hydrophobic particles and rejection of the hydrophilic particles, using model and industrial coal feeds. The new system hydrodynamics was found to offer significant advantages in flotation kinetics, desliming, and particle recovery across a broad particle size range. Bubble interfacial flux reached as high as 600 m2/m-2s-1. The upper system consists entirely of a high volume fraction of bubbles, which provides a "safety-net" to protect against the detachment of coarse particles, and a highly permeable zone for counter-current washing of the flotation product. Under relatively extreme conditions selective stripping of the product is observed, resulting in grade recovery values beyond those achievable using the tree flotation method."

Jan 1, 2014

By M. A. C. Goes

Over the last twenty years CETEM has been involved in the preparation and certification of reference material from ores, concentrates and related materials for use in analytical laboratories associated with mining, metallurgy and earth sciences. CETEM provides reference materials from mineral samples, not available from the existing sources, to meet the needs of the mining and mineral industry. CBPA-1 and CBPA-2 respectively are samples of copper sulphide ore and flotation concentrate, both originated from Sossego Mine, located in Pará State, Brazil. The batch characterization was carried out by a collaborative measurement program with participation of 20 industrial, commercial and government laboratories from South America, North America, Europe, Asia and Africa. The certificates and report were issued in December 2007 (http://www.cetem.gov.br/ mrc/ing). The objective of this paper was to describe the procedures for material preparation and the collaborative approach involving a collaborative interlaboratory measurement program to certify CBPA-1 and CBPA-2.

Jan 1, 2014

By Ali Dehghani Ahmadabadi

Gravity separation technique is widely used in mineral beneficiation practices for its low cost, ease of operation, and eco-friendly nature. The most commonly used gravity methods for beneficiation of iron ores are shaking table, jig and spiral methods which aren&apos;t very effective for iron mineral fines. Multi-Gravity Separator (MGS) is suitable for the treatment of fines with a maximum particle size of approximately 500 µm. This study was performed on a low-grade haematite iron ore from Chadormalutailing dams. Detailed mineralogical, physical and chemical characteristics of a goethiticlateritic iron ore showed that the sample contained porous and friable oxides and hydroxides of iron. The ore sample had a feed grade of 35% Fe, 1.1% FeO, 15.5% SiO2 and 8.5% Al2O3. Haematite and goethite are the main iron-bearing minerals while quartzite and kaolinite are the major gangue mineral constituents. In this paper, the statistical evaluation and optimization of grade and recovery using a factorial design of experiment (DOE) method has been carried out for Chadormalulow grade ultra-fine haematite tailings with MGS. The input process parameters considered are drum speed, tilt angle, wash water flow rate and pulp density. The results show considerable improvements in both the quality characteristics such as grade and recovery with statistical evaluation and optimization, compared to the initial value of grade and recovery. Keywords: multi-gravity separator, mgs, statistical evaluation, optimization, haematite

Sep 1, 2012

By S. V. Balikov

A close association of fine gold with pyrite and arsenopyrite is observed at many gold ore deposits. Gold forms and gold distribution in sulfides of different deposits were studied. A comprehensive method including optical spectrum analysis, X-ray diffraction analysis, high-temperature X-ray diffraction analysis, scanning electron microscopy and X-ray microanalysis was used for the de-tailed investigation of the defectiveness of the structure of pyrites from different deposits. It was shown that gold content in pyrites varied from 1.5 g/t to 100 g/t compared with 30 g/t to 230 g/t in arsenopyrites. The nature and arrangement of inner inclusions of foreign minerals in pyrites as well as occurrence of fractures and macro- and microdefects were examined. Iron, sulfur and arsenic contents in all pyrite samples were determined. Chemical composition of the detected inclusions was defined. Parameters of submicroscopic structure of pyrite were estimated using X-ray Fourier line shape analysis. Estimation of real structure characteristics with calculation of submicroscopic D-blocks and root-mean-square lattice deformation 2()evalues was performed. Defectiveness of the pyrite structure can be defined by the following parameters: K?1 and K?2 reflection resolution and its intensity within high Bragg angles; S content to Fe content ratio; the arsenic content and the local variation in its concentration within the pyrite grain; the size of submicroscopic blocks and the crystal lattice disturbance value. The performed analysis showed some relationships between these parameters and refractoriness of the pyrite concentrates. It was observed that at S/Fe ratio close to or higher than stoichiometry and at a low arsenic content (less than one per cent) pyrite concentrates were not refractory. The analysis results showed that concentrates of the Beryozovskaya gold plant (0.02~0.07 As%) and concentrates of the Kommunarovskaya gold plant (less than 0.05 As%) as well as some pyrites from the Sukhoi Log gold deposit had the same characteristics. It was found that average arsenic content did not excess 0.2 per cent. Arsenic distribution within the grain of all these pyrites was uniform. At S/Fe ratio lower than or close to stoichiometry and at high arsenic content (more than one per cent) and significant variation in its distribution within a pyrite grain the pyrite concentrates are refractory. Such pyrites are typical for the Sukhoi Log, the Bakyrchik and the Mardzhanbulak gold deposits. Direct correlation between arsenic and gold contents, in particular, in fine grains of pyrite was observed by many researches. This correlation was explained by impregnation of arsenopyrite which normally contains more gold than arsenic pyrites. The analyses results showed that no arsenopyrite impregnation was found: arsenic distribution within the grain was uniform but its concentration varied greatly. Such pyrite concentrates are more refractory. This paper presents the results of analyses of sulfides to determine the amenability of the associated gold recovery by hydro- and pyrometallurgy.

Jan 1, 2014

By K. B. Venkatesh

In view of the importance of size reduction operations from the energy consumption point of view, during the past one hundred and fifty years a number of attempts have been made to develop a general energy-size relationship. However, to date it has not been possible to clearly define the scope of application of various laws proposed, though Bond law has been widely used with a moderate degree of success. It was, therefore, decided to carry out a thorough analysis of the batch ball mill grinding data available in the literature on several minerals that were ground under normal range of operating conditions. In particular, relationships proposed by Bond, Kapur, Voller and Morrell were examined in detail. It was found that the relationship proposed by Kapur in terms of the mean particle size provided a more accurate basis for estimation of energy. The value of the exponent of mean particle size in this relationship, a, is same as the exponent of the particle size in the expression for the grinding rate parameter: ki = Axia, provided the particle size distribution range corresponds to the self-similar or self-preserving regime. In this regime, mean particle size is proportional to the 80 percent passing size and size modulus. From the practical application point of view, it was possible to describe the variation of µ1 over limited, but significantly broad size range, using a constant value of a. Thus, by carrying out a few batches, grinding experiments over the size distribution range corresponding to the contents of the production mill, the applicable value of a can be determined. Keywords: energy-size relationship, ball mill, mean particle size, self-similar size distribution regime

Sep 1, 2012

By B. K. Mohapatra

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

Sep 1, 2012

By L. X. Zeng

The low-grade nickel resource is very abundant in the western China. To make the best of the low grade nickel ore, we studied the process mineralogy and beneficiation process of nickel ore in the Yangliuping of Sichuan province. The grade of nickel is 0.45% in the nickel ore and the nickel exists in the sulfide including pyrrhotite, nickel pyrite, chalcopyrite primarily. The nickel distributes in pentlandite and magnetite over 90% and the copper does in the chalcopyrite over 95%. The carri-er minerals of useful elements tend to concentrate in space for concentration. As is shown in grain-size analysis of three major minerals, there are some more fine-grained grains with the percent of from 10% to 15% (<40 µm) and from 3.5% to 5.5% (<20 µm). So there will be some smaller size disseminated sulfide minerals difficulty to liberate in the process of concentration. The liberation degree of sulfide is not high and there are much associativity in the sulfides. Therefore, we should reclaim the sulfide aggregation dominating as pentlandite when concentrating. In addition, the sulphide minerals is so brittle that we should take care of preventing from crashing when grinding. The gangue mineral is mostly metamorphic mineral including serpentine, chlorite and amphibole which is easy to be argillization and difficult to concentrate.

Jan 1, 2014

Deposit of Jinbaoshan platinum group metals (PGMs) ore ranks the second in Yunnan province, China. Study on process mineralogy shows that the grade of copper and nickel in the ore is just 0.16% and 0.20% respectively. Moreover, the impurities magnesium and iron and silicon are rich in the ore. The copper and nickel sulfides in the ore, especially the minerals of PGMs, are finely disseminated with the gangues like serpentine. Fine grinding is necessary, and high content of MgO in the concentrate is inevitable through traditional flotation technology on the basis of process mineralogy studies. Achievements on process mineralogy studies are applied to guide the flotation tests successfully. Progress on flotation has been made through comprehensive studies. Suitable grinding by two stages is required for the finely disseminated PGM sulphide ore to achieve the liberation of valuable minerals. Stage grinding and stage separation shall be conducted for a better recovery of valuable minerals. The oxidation state of mineral surface of sulphides has a profound effect on flotation of the ultra-fine sulphides, especially the pentlandite. The combined use of collectors enhances the selective collecting on the ultra-fine and low-grade valuable minerals, is of critical importance.

Jan 1, 2014

By Mingfei He, Xianyang Qiu, Hongling Wang, Yude Gao, Qingbo Meng, Xiaoping Xu, Hao Bu, Guosheng Wang

"By the “HGMS-gravity separation-Flotation” combined technology, a Ta-Nb-Sn concentrates with the grades of Ta2O5 10.00%, Nb2O5 9.13% and Sn 42.62% at the recovery of 44.73% , 9.13% and 14.62% respectively , a Li concentrate with Li2O 3.11% at the recovery of 51.83% and a feldspar product are obtained. All the value elements are recovered. INTRODUCTION Tantalum and niobium widely used in electronics, atomic energy, aerospace and chemical industries with the advantages of high melting point, good ductility, low vapor pressure, strong corrosion resistance and high thermal conductivity are important strategic resources which occupy a special position in rare metals (Crockettand, 1993; GAO and Zou, 2004; LI and HUANG, 2012). China's tantalum and niobium resources are poor, and most of them are low grade, fine grain size, complex composition, and polymetallic associated with refractory mineral resources (GAO and Qiu,2013; Liang and Li, 2016). The Yifeng Lion Ore containing lithium and other rare metals (tantalum, niobium and tin) is an important lithium polymetallic deposit in China Jangxi province. While extensive previous studies have settled on the recovery of lithium minerals by flotation only for the extremely low grade of Ta, Nb and Sn. The regular gravity concentration is used, and the recovery of tantalum niobium is low. This paper sets out to investigate a combined flowsheet of “high gradient magnetic separation (HGMS),-gravity separation-flotation” to effectively comprehensively recover the Li, Ta, Nb, Sn and the feldspar resources in the Yifeng Lion Ore.ORE PROPERTIES Chemical analysis The run of mine comes from Jiangxi Yifeng Lion ore. The main elements of the ore chemical analysis results shows in Table 1. From Table 1, the grades of lithium and rubidium are relatively high, but the grades of tantalum and niobium are lower than the industrial grade; feldspar with high content of potassium and sodium and low content of iron can be used as high-quality ceramic raw materials."

Jan 1, 2018

By Ritva Tuunila, Antti Häkkinen

"Wet stirred milling is one of the most effective grinding methods in producing very fine powders in large scale production, but it is a very energy-intensive process, especially when pigments or even nanosize particles are produced. Therefore, optimization of the process parameters is important. The specific energy consumption is a function of stress energy, which describes the effect of tip speed and media characteristics on grinding, and is related to the number of stress events during milling. In effective grinding, a large number of stress events are necessary. The stirrer configuration influences the number of stress events and their intensity, because the motion of the media and the suspension are caused by a rotating stirrer. The effect of three different stirrer types: a pin, disc and blade stirrer, on the grinding rate, i.e. product fineness and specific energy were investigated in this study with a batch laboratory stirred mill with two tip speeds. The material used in the experiments was calcium carbonate. The results obtained with the smaller tip speed showed approximately similar grinding rates and energy consumption for all stirrer types. When looking in closer detail, the blade stirrer seemed to give the finest product with all grinding times compared to the pin and disc stirrers, which both produced almost the same product fineness with the same grinding times. The specific energy consumption was the smallest with the disc stirrer, but not very big differences were obtained. With a higher tip speed, greater differences occurred both in grinding rates and specific energy consumption. The blade stirrer was the most effective in grinding, and the pin stirrer the next. The net energy consumption for the blade and pin stirrers was almost the same, but it was slightly higher for the disc stirrer."

Jan 1, 2014

By T. Wang

A new kind of mixed chelate collector was prepared in order to separate gangue (mainly quartz) and iron minerals efficiently during room temperature. The collector was was composed of two chelate collectors DES and DA. Floatation tests were carried out under 28?C with mixed magnetic concentrates of Donganshan sintering plant, and optimal test conditions were obtained, in which a direct floatation was undertaken and then a close circuit reverse floatation tests ( one roughing, one cleaning, and two scavenging), and a result of Fe grade of siderite concentrate 36.49%, Fe recovery of siderite concentrate 10.65%, Fe grade of final concentrate 66.27%, Fe recovery of final concentrate 78.25%, and Fe grade of tailing 16.22% was gained. Fourier Transform Infrared (FTIR) Spectra and Zeta potential of siderite and quartz adsorbed the DES and DA collectors before and after were measured in order to analyze the reaction between the collector and minerals. Furthermore, Molecule electronegativity of DES and DA was calculated by Electronegativity Proportionality Principle (EPP). Results showed that DES molecules have five active adsorbing sites and their molecule electronegativity was stronger while DA molecules have three active adsorbing sites, which all make the mixed collector has efficiently collecting performance. Zeta potential results of Quartz and siderite showed that at pH 11.5 both kinds of particles charged negative and at this pH value the mixed collector were mainly anionic collectors which indicate that electrostatic adsorption does not exist at this situation. FTIR results showed that characteristic peaks of the mixed collector changed before and after adsorption with the minerals, which means that there exists chemical adsorption or hydrogen bonding.

Jan 1, 2014

By D. W. Lui

Tangdan copper oxide deposit, the largest independent copper oxide deposit in China so far, is located in the northeast of Yunnan province, southwest of China. The ore from this deposit is characterized by: low grade, high oxidation ratio, and fine-grained disseminated, resulting in poor flotation recovery. The analysis of plant flotation data reveals that the loss of copper minerals into tailings is mainly with coarse size fraction due to insufficient liberation and fine size fractions, especially minus 10 microns fraction, which proved to be very difficult-to-float with conventional flotation. In order to improve the performances of plant flotation of this ore, laboratory flotation tests were conducted with the pure malachite, which was the principle copper oxide mineral in Tangdan copper ore deposit. The results obtained suggested that a finer grinding (up to -325 mesh or even -400 mesh) be needed to decrease the loss with coarse size fraction, and vacuum flotation is a promising way for the recovery of the fine malachite effectively.

Jan 1, 2014

By B. Y. Tuo

This article presents results of studies performed to determine the feasibility of using low-grade limonite ores from Yunnan deposit. Laboratory tests and commercial trials have shown that the magnetic separation can be performed with a high high-grade concentrates. The quality of the preparation concentration meets the requirements for blast-furnace smelting. The results as a whole demonstrate the feasibility of utilizing bank limonite ore. Yunnan limonite?s magnetic separation, roasting magnetic separation and flotation experimental study shows that grade of the ore grade can been from 32.91% heighten to 58.64%, a recovery of iron is 90.87%, productive rate is 51% through orthogonal test optimization; roasting magnetic separation can obtain that the grade of iron preparation concentrate is 61.16%, recovery is 67.39%, productive rate is 36%. As far as economy and enviormental protection is concerned, strong mag-netic suggested a more appropriate process. Keywords: Limonite; Roast; Magnetic separation

Jan 1, 2014

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 Chuanyao Sun, Zhenguo Song, Aijun Ren

"Lead-Zinc sulfide ore containing carbon is one of typical refractory resource. Because organic carbon has natural floatability, it has a great negative influence on flotation of lead-zinc sulfide ores. It is difficult to obtain qualified lead concentrate by flotation using conventional reagent system. In order to reduce the influence of organic carbon on lead and zinc flotation, two alternatives, i.e., removal carbon by pre-flotation or adding carbon depressant in lead and zinc flotation, can be adopted. If the flowsheet of removal carbon by pre-flotation is adopted, the lead and zinc recovery which are lost in the carbon tailings is often higher. If conventional carbon depressants are used in lead and zinc flotation, it has a certain depression effect on lead minerals. In the paper, the ore grading 1.49% Pb, 5.73% Zn, and organic carbon 3.14% has been carried out by flotation tests. In the bulk flotation of carbon and lead minerals, Pb concentrate is only about grading 21.00% Pb. In the pre-flotation of organic carbon, it’s hard to remove most of carbon, and the loss of Pb and Zinc recovery is too much in carbon tailings. New carbon depressant BK547 and three conventional carbon depressants including sodium lignosulphonate, Fe-Cr-lignosulfonate, and dextrin, are used in comparative tests of Pb flotation, and results show that BK547 have a good depression on organic carbon, and have almost no effect on Pb flotation. Using BK547 as carbon depressant, cycle test in flowsheet of two-stage Pb and Zn selective flotation and coarse concentrate regrinding have been carried out, and results show that concentrates grading 62.51% Pb and 41.33% Zn are obtained at recoveries of 70.24% Zn and 83.58% Zn.INTRODUCTION Lead and zinc mineral resources are still the main source of lead and zinc metal. For lead-zinc sulphide ore, lead concentrate and zinc concentrate are obtained by flotation. Minerals separation by flotation is a complex process and is dependent, along with other factors, on the relative distribution of hydrophobic (e.g. collector, polysulfide, elemental sulfide) and hydrophilic (e.g. metal hydroxide) species between different particles of the same mineral phase and between mineral phases (Piantadosi & Smart, 2002; Boulton, Fornasiero & Ralston, 2003). Organic carbon has natural floatability and is often associated with lead-zinc sulphide ores. If the content of organic carbon in the ore is more, the flotation index will become worse. The main problem is that the concentrates has too low grade to be qualified grade products for sale. At the same time, the consumption of reagents is more than ordinary ores without organic carbon. In order to solve the influence of organic carbon on the flotation of lead and zinc and other sulphide ores, a lot of research works have been carried out."

Jan 1, 2018