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By E. V. Leont'eva, V. R. Vafin, A. G. Kladova, N. L. Medyanik, A. P. Ponomarev

"The thermochemical technology for recovery gold, silver from mature flotation tailings of pyrite ores by two-stage sintering with chlorine-ammonium reagents and leaching of bakes with water is developed. The economic effect of the technology implementation is calculated. KEYWORDS Gold and silver recovery, mature flotation tailings, pyrite ores, chlorine-ammonium reagents, leaching INTRODUCTION The relevance of processing the technogenic reserves of ore mining and concentration wastes accumulated over the long operation periods of mineral deposits is beyond doubt. Hundreds of tons of precious and non-ferrous metals that were not recovered during the primary processing are concentrated in the accumulated wastes. For example, tailing storage facilities (TSF) at the mineral processing plants in the Southern Urals region of Russia hold more than one hundred tons of gold and thousands of tons of silver in the mature flotation tailings (MFT) of copper pyrite ores. Secondary processing of technogenic materials is facilitated by the fact that they are preprocessed — ground and compactly deposited on the surface, but complicated by the low (compared to ores) grade, uneven distribution of the valuable minerals, altered and unstable material composition of the mineral feed. This leads to low recovery rates of valuable metals by conventional physicochemical processing methods. It is necessary to develop innovative methods of deep integrated processing of industrial wastes, based on the study of the material composition of specific technogenic mineral resources and based on ensuring the environmental safety of the resulting secondary waste. MATERIALS AND METHODS The object of this study are the MFT of Uchaly Concentrator in the Republic of Bashkortostan, Russia, which processes copper pyrite and copper-zinc ores. According to the data of particle size, chemical, X-ray phase, and optical microscopic analyzes, samples of the MFT are characterized by considerable particle fineness (the fraction of the size class -41 µm is 89-90%), low grade of gold (1.7-2.0 g/t) and silver (18.0-19.5 g/t), comparable grades of iron (23.0-26.8%) and sulfur (28.1-28.7), but different grades of copper (0.12 - 0.73%) and zinc (0.34-1.26%). The most common mineral in the samples is pyrite (50 - 56%), found in intergrowth with grains of other sulfides (0.1-4%) - sphalerite, tennantite, chalcopyrite, covellite. Quartz and aluminosilicates are the dominant rock minerals (21-36%). Based on the results of the phase analysis, it was established that approximately 70% of the gold in the samples is found in a form that is resistant to cyanidation — it is locked in sulfides (52-56%) and quartz and coated by films. Along with sulfides and rock minerals, common gold and silver minerals (chalcogenides, intermetallides, native gold) in the form of submicro- and nano-sized precipitates are also associated with intergrowth of a complex structure. These features of the material composition allow classifying the MFT of the pyrite ores as a mature feed material and emphasize the need for chemical non-cyanide methods to recover gold and silver from the mature minerals and associations they are found in."

Jan 1, 2018

By Luis Bergh, Francisco Díaz, Isabel Panire, Juan Yianatos, Luis Vinnett

"In this paper, industrial data of residence time distribution (RTD) in mechanical flotation cells is presented, which were measured in several plants using the radioactive tracer technique. Results include mechanical flotation cells of 100 - 300 m3, which were operated with effective residence times from 2 to 7 min. Tests were developed in forced air and self-aerated cells for liquid and solids (non-floatable) per size classes. Different RTD model structures including perfect mixer, Large and Small Tanks in Series (LSTS) and by-pass flow plus perfect mixer were evaluated and compared for single cells. The actual mixing regime was related to the effective residence times and cell designs based on the model fitting. By-pass flow percentages were typically lower than 10%, therefore the perfect mixer model plus a dead time was a suitable model structure in most cases. On the other hand, the arrangement of cells in series was effective in reducing the bypass flow observed in single cells (by-pass flow percentage greater than 5%). In addition, the N perfectly mixed reactor in series model was used to represent the mixing regime along flotation banks, which were operated with mean residence times in the range of 18 - 53 min. Results showed that N tanks in series plus dead time were closer to the actual values in most cases."

Jan 1, 2014

By Z. F. Dai

The Thompson Nickel Belt, located in Northern Manitoba, Canada, contains more than 400 million tonnes of ultramafic low-grade nickel mineralization in 9 deposits. One of these, the Pipe deposit, was mined between 1971 and 1983 producing a Ni concentrate of 9% to 11% grade at 73% recovery. The operation was terminated because the poor recovery, low Ni grade and high MgO con-tent of the concentrate made processing this ore uneconomical at the metal prices prevailing at that time. Vale Inco is now re-evaluating these deposits with the intent of providing a new source of concentrate for its Thompson integrated processing facility. It is known that serpentine slimes coating on the sulphide surfaces are responsible for the poor metallurgy. Various desliming methods were tested including settling, elutriating, tabling and hydrocycloning. Hydrocycloning has proven to be the most effective means of desliming. With a feed assaying 0.70% Ni and 33% MgO, a rougher flotation concentrate assaying 22.7% Ni and 4.8% MgO was produced at a Ni recovery of 62.5% (relative to the feed to grinding), with no utilization of slimes dispersants. Another 16.6% Ni recovery was achieved in scavenger flotation resulting in a total Ni recovery of 79.1%. Batch flotation tests using various flowsheet configurations to optimize the grinding, desliming, reagentizing and flotation conditions are in progress.

Jan 1, 2014

By U. A. Peuker, A. Grohme

"The cost effective and sustainable production of high purity rare earth elements is still challenging. The most commonly used method, known as solvent extraction, is very complex, needs many steps to separate the rare earth ions in a high purity, and causes a lot of environmental problems. An alternative method is the utilization of a separation column filled with adsorbent material. Within the column the desired target substance can be selectively separated out of a multi component mixture in one step. To find out if an adsorption material is suitable for the separation with a column, a characterization of the adsorbent material is necessary. In this paper are described the usage of modified activated carbon (oxygen and phosphorous modification), Di-(2-ethylhexyl) phosphoric acid (D2EHPA) impregnated Amberlite® XAD 7 HP and Lewatit® VPOC 1026 as adsorption materials. Several adsorption isotherms and process conditions are determined to characterize the selectivity between the adsorption materials and different rare earth ions. With these isotherms the mass specific rare earth ion retention for the adsorbent materials can be calculated. In addition to the isotherms the best working parameters are investigated e.g. pH-value and the time for equilibration. Finally, the separation factors between two rare earth elements are determined. The separation factor shows how successful a two element system can be separated."

Jan 1, 2014

By I. A. S. Brum

Brazil has large reserves of coal (7 billion tons) located predominantly in the south of the country. Despite the relative ease of exploitation, these resources have been underestimated as a source of energy. Currently the country's strategic planning aims to minimize the gap between coal and resources traditionally used in the Brazilian energy matrix. In this scenario, the processing of Brazilian coal in columns flotation appears as a promising alternative, having been successful in mineral systems containing fine particles. This paper aims to study the overall process of column flotation for a specific Brazilian coal. The material used in this study consists of fine coal that is usually discarded as waste and deposited in the tailings dams. The objective is to recover carbonaceous matter and reduce the ash and contaminants grade. The variables analyzed were the influence of size particle, solids concentration in the pulp, the concentration of surfactants and frothers to be used. The first results showed that it is possible, from a feed with around 65% of ashes, to obtain products with approximately 35% ashes with mass recoveries about 40%. Concentrations of frother (pine oil) ranged from 50 g/t to 300 g/t and the collector (diesel oil) of 50 g/t to 200 g/t. Keywords: column flotation, coal, tailings, negative bias

Sep 1, 2012

By D. W. Feng

Thiosulfate as an alternative lixiviant for gold has received much attention in recent years due to the growing environmental concerns over the use of cyanide. Cu(II) is used to catalyze the oxidation of gold and ammonia is to stabilize Cu(II) in alkaline solution. However, the presence of Cu(II) enhances the oxidative degradation of thiosulfate. The high consumption of reagents has made thiosulfate leaching uneconomical to date, in comparison with cyanide leaching. Therefore, further investigations are required to leaching under low reagent concentrations over extended periods where reagent consumption is low and gold leaching rate is low as well. It is thus necessary to find additives to enhance gold dissolution in the thiosulfate leaching system. Electrochemical studies using rotating gold disc showed that the oxidation of gold in ammoniacal thiosulfate solution increased significantly in the presence of a small amount of thiourea and the beneficial effect increased with an increase in the thiourea concentration up to 7.2 mM. Thiourea improved gold extraction in the thiosulfate leaching of gold from a sulfide ore under low reagent concentrations. Excessive addition of thiourea inhibited the leaching of the gold ore. The decomposition of thiosulfate was reduced with the presence of thiourea. X-ray Photoelectron Spectroscopy (XPS) results indicated that thiou-rea likely decomposed to sulfide and formed gold sulfide species at the gold surface.

Jan 1, 2014

By Zhen Hu, Chenghang Wang, Jianjian Zou, Xiujuan Li, Hanwen Li, Tai Wang, Baoxu Song

A rare earth deposit in Sichuan, China is rich in wulfenite in addition to bastnaesite and barite, and the deposit has already adopted high intensity magnetic separation technique to pre-enrich the rare earth minerals, during which the wulfenite and barite have been left in the non-magnetic product. Because the chemical composition of wulfenite is only 24.17% Mo and 42.15% Pb, it is very hard to obtain qualified lead or molybdenum concentrates by physical separation process. To pre-concentrate the wulfenite, flotation has been first introduced, and by using 100g/t GY-7 as dispersant, 1500g/t CP-1 as depressant and 160g/t GY-4 as collector, a floatation concentrate assaying 15.52% Mo with a recovery of 86.12% was obtained. Then a leaching process was conducted by using 1.2mol/L Na2S as leaching agent, and molybdenum can be translated into sodium molybdate in the solution with a leaching recovery of 99%. To recover rare earth oxides (REO) and lead from the leaching residue, a magnetic separation process has been conducted. Above all, by the whole process, the molybdenum recovery reaches 86.16% in the sodium molybdate solution, and a lead concentrate assaying 54.59% Pb with recovery of 27.99% could be obtained.

Jan 1, 2014

By Herman Mendoza, Luisa Ciriaco

An investigation of the seeding mechanism using heterocoagulation and the Derjaguin - Landau - Verwey - Overbeek (DLVO) theory of two pure minerals is the objective of this study. The seed was silica with modal diameters of 4.529 µm (fine particles) and 5.216 µm (coarse particles) and the test powder was hematite with a modal diameter of 0.883 µm. Parameters included variations in the concentration and particle size of silica. Both particles were first subjected to size determination and settling tests via Stoke’s Law. The silica-hematite system was then suspended in 0.001 M KNO3 electrolyte solutions where pH was adjusted. Afterward, the suspension was agitated and was allowed to settle at different time intervals while obtaining turbidity readings. Results through settling time and turbidity vs. pH graphs indicate that maximum heterocoagulation occurs from pH 2 to 5 while minimum heterocoagulation occurs from pH 6 to 7. At these pH levels the zeta potentials of silica and hematite are negative and positive, respectively. On the contrary, repulsive forces dominate the system from pH 8 to 11 due to the particles’ negative zeta potentials. Furthermore, the attachment of hematite to silica increases when the parameters become larger thereby increasing the effectivity of heterocoagulation at acidic to neutral pH levels.

Jan 1, 2014

By Tong Guan, Guangqin Fu, Lihong Deng, Xiaotong Zhou

The raw ore contains many kinds of valuable metal elements. It is rich in valuable metal elements, such as Pb, Zn, Sb, Sn, Ag and is associated with S,Fe,As,In,Cd. A new technology which includes medium intensity magnetic separation, preferential lead flotation, zinc-sulfur bulk flotation, zinc-sulfur flotation separation, tin gravity separation and tin slime flotation was chosen to recover valuable minerals in the raw ore, and the iron concentrate, the lead concentrate, the zinc concentrate, the sulfur concentrate and the tin concentrate can be obtained. To the zinc-sulfur bulk concentrate with the grade of 7.21%Zn and 16.97%As acquired by the zinc-sulfur bulk flotation, two different flowsheets of Zn-S separation technology are chosen to study. The lab test shows that when using lime+DM as depressor, the zinc concentrate assaying 47.73%Zn with the recovery of 95.01% can be obtained and the test indexes are better than that of the flowsheet which using lime as depressor. At the same time, the Cd content is 0.65% and the In content is 0.10% in the zinc concentrate, so they all reach the smelting comprehensive recovery standards and the As content decreases to 0.23%. Finally the flotation separation of christophite and As-bearing sulfide minerals can be successfully realized.

Jan 1, 2014

By Sergio Castro, Janusz Laskowski

In the flotation in highly concentrated electrolyte solutions not only ionic strength but also chemical composition, flotation pH and mineral properties play an important role. As a result an improvement or depression may be observed. A strong improvement in the floatability of hydrophobic bituminous coals and other inherently hydrophobic minerals is characteristic for the "salt flotation process", which is carried out in concentrated electrolyte solutions at natural pH and without addition of any flotation reagents. However, pH is a key factor when saline and hipersaline waters with hydrolyzing metallic ions are used as process water. In the case of flotation of Cu-Mo sulfide ores in seawater, molybdenite is strongly depressed by the Mg2+ hydrolysis products when pH is raised to depress pyrite. Process water with a high content of Mg2+ and Ca2+ ions, and this also includes seawater, have deleterious effect on the anionic flotation of industrial minerals with fatty acids. The flotation of potash ores is carried out in NaCl-KCl saturated brine (6-7 mole/L of NaCl and KCl) with the use of long-chain amines. Such a high electrolyte concentration changes dramatically the amine Krafft point making the collector insoluble in brine; its mode of action entirely differs from that in conventional flotation processes.

Jan 1, 2014

By Montaser S. EL Salmawy, Takahide Wakamatsu, Eishi Kusaka

"The effect of ethoxylation of long chain alcohols on the quartz and feldspar flotation behaviors was investigated for surfactants of the octylphenol polyoxyethylene family. The tests were carried out with an average of 7.5 (Triton X 114), 9.5 (Triton X 100), 30 (Triton X305), 40 (Triton X 405) and 70 (TritonX 705) oxyethylene units. Electrokinetic and flotation studies on quartz have shown that at low pH values, quartz could be floated with polyoxyethylene octytlphenols. The possible mode of adsorption of these nonionic surfactants may be suggested to be due to hydrogen bonding between the oxygen of ether and hydrogen in the free hydroxyl groups on quartz surface. It was anticipated that the selectivity between quartz and microcline would decrease as oxyethylenic chain length increases.Electrokinetic methods were used to elucidate the mechanism of surfactants adsorption and to correlate these findings with the optimum flotation conditions. The polymeric surfactants would reduce the negative values of zeta potential of quartz while the reverse is true for feldspar. After treatment of microcline with HCl, its electrokinetic behavior in the presence of Tritons X surfactants became similar to that of quartz.The obtained results would suggest that separation of quartz from feldspar using nonionic surfactants of octylphenoletoxylated alcohols having shorter oxyethylenic chain length at low pH could be achieved"

Jan 1, 2003

By H. Y. Ran

Magneto-flotation is a recently developing technology of mineral processing. This technology combines the traditional magnetic separation with flotation and provides a magnetic and buoyant force compound field to separate different minerals based on the differences of magnetism and floatability between useful minerals and gangue minerals. Magneto-flotation is applicable to separate ferromagnetic from nonmagnetic minerals, especially to separate magnetite from floatable gangue minerals. This magneto-flotation column is equipped with pulse electromagnetic field in the middle upper part, a bubble producer and upwell water inlet in the bottom. The pulse time of electromagnetic field in the column can be optionally adjusted and the maximum intensity of pulse magnetic field can reach 35mT. Pulse magnetic field can not only restrain magnetic minerals from entering into the froth layer, but also destroy the magnetic agglomeration to ensure a higher recovery and a higher concentrate grade. The results of experiments with magnetite show that pulse magnetic field has remarkable influence on flotation process. Compared with single flotation, magneto-flotation column can improve the recovery of magnetite by ten percent. After one time?s magneto-flotation processing, the grade of iron concentrate can increase from 65.4% to 69.5%, the content of silicon oxide in iron concentrate decreases from 7.4% to 2.8%, and the total recovery is above 96%.

Jan 1, 2014

By Ali Dehghani Ahmad Abady

In Sarcheshmeh Copper Complex, a molybdenite floatation concentrate was produced by selective removal of copper impurities with 7 stages of cleaner floatation banks, that contains around 1% copper and 54.5% molybdenum. Mineralogical studies, physical and chemical characterizations showed that the sample contained 91% molybdenite (MoS2), 2.6% chalcopyrite (CuFeS2), 0.2%Covellite (CuS), 1.6% pyrite (FeS2) and 4.6% other nonmetallic impurities. In the present study, the removal of the copper sulfide was investigated by ferric chloride leaching for achieving a molybdenite concentrate with copper content less than 0.3%. Therefore, the application of response surface methodology (RSM) for modeling and optimization of the influence of three control parameters including ferric chloride (FeCl3) dosage, reaction temperature and reaction time, each varying in five levels, were considered, and the mathematical model for prediction of %Cu removal from molybdenite concentrate was derived. This study showed that RSM could effectively be applied for the modeling of ferric chloride leaching, and finding an economical optimum condition to achieve maximum de-coppering efficiency while consuming minimum leaching reagents. The optimum dissolution conditions to maximize Cu extraction were 200 kg/t for ferric chloride dosage, 120 min for reaction time and 80 °C for reaction temperature with a prediction of 83.20% Cu dissolution. Keywords: molybdenite concentrate, ferric chloride, leaching, response surface methodology (RSM)

Sep 1, 2012

By Sergio Castro, Janusz Laskowski

In the flotation in highly concentrated electrolyte solutions not only ionic strength but also chemical composition, flotation pH and mineral properties play an important role. As a result an improvement or depression may be observed. A strong improvement in the floatability of hydrophobic bituminous coals and other inherently hydrophobic minerals is characteristic for the "salt flotation process", which is carried out in concentrated electrolyte solutions at natural pH and without addition of any flotation reagents. However, pH is a key factor when saline and hipersaline waters with hydrolyzing metallic ions are used as process water. In the case of flotation of Cu-Mo sulfide ores in seawater, molybdenite is strongly depressed by the Mg2+ hydrolysis products when pH is raised to depress pyrite. Process water with a high content of Mg2+ and Ca2+ ions, and this also includes seawater, have deleterious effect on the anionic flotation of industrial minerals with fatty acids. The flotation of potash ores is carried out in NaCl-KCl saturated brine (6-7 mole/L of NaCl and KCl) with the use of long-chain amines. Such a high electrolyte concentration changes dramatically the amine Krafft point making the collector insoluble in brine; its mode of action entirely differs from that in conventional flotation processes.

Jan 1, 2014

By Subrata Roy

The slag produced in metallurgical smelting processes contains considerable amount of valuable metals. The present work focuses on recovery of copper value by froth flotation from the copper smelter slag. The slag samples are subjected to detailed characterization including automated quantitative mineralogy using Mineral Liberation Analyzer (MLA). The MLA results indicate that a part of copper is present in sulphide form in fayalite (iron-silicate) matrix and very poorly liberated. SEM-EDX analyses also reveal that a certain part of the copper is present within the fayalite (Fe-Ca-Si) matrix, perhaps either as lattice impurity or very fine grain. The copper smelter slag is ground in wet mode in a ball mill, to particle size distribution suitable for flotation. Subsequently, the slurry undergoes froth-flotation in lab-scale Denver-type flotation cell. Various combinations of sodium iso-propyl xanthate (SIPX) and alkyl hydroxamate collectors are used along with pine oil as frother. A multiple collector dosage is required for effective flotation of non-sulphide copper phases associated fayalite and alkyl hydroxamate is identified as one of the effective collector for non-sulphide minerals. The copper recovery can be enhanced by 7-10% with a combination of collectors as compared to conventional sulphide flotation using only xanthate type collectors. However, the grade decreases to some extent. Keywords: flotation, copper slag, hydroxamate collector

Sep 1, 2012

By P. M. Solozhenkin, V. P. Nebera, V. D. Skorohodov

"Aqueous dispersions of air specifically prepared on the basis of flotation reagent solution acquire active properties in relation to the solid surface and may be imply as a medium for flotation ores. Such aerated solutions were defined as Activated Aqueous Dispersions of Air (AADA) with 5-6:1 ratio and the procedure of mineral separation was defined as a flotation process under non-equilibrium conditions. These procedures differ from conventional flotation mainly in reagent interaction with minerals through their adsorption films on gas bubbles simultaneously forming flotation complexes. Gas-liquid mixture is kept till an adsorption layer of reagents close to the limit value is formed on the surface of bubbles. The theoretical basis of AADA is considered.The results of two-phase gas- liquid mixture (AADA) were studied. Two versions of AADA were considered. Electrical surface potential of the air bubbles with the surfactant film, untreated minerals particles in Coulomb attraction and flotation complex formation were analyzed. Special flotation machines were constructed for separation of mineral suspension in air dispersion, whose effectiveness was confirmed in flotation of apatite waste of Kovdor ore. FK-9 machine introduced on Kovdor iron ore combine for the recovery of apatite from current departures of factory. The froth product contained 7.2 % P2O5, with recovery of 44 % apatite, which was processed for getting melted phosphoric-magnesium fertilizers."

Jan 1, 2003

By K. Anand Rao

Extensive exploration for uranium metal by the Department of Atomic Energy in India led to the discovery of new carbonate host rock uranium deposits Tummalapalle and Gogi, which are located in southern part of India. Alkaline leaching at elevated temperature and pressure using autoclave is the central unit process in the overall strategy for extraction of uranium from these new ore deposits. Autoclave leaching is affected largely by intrinsic nature of the ore and by the process variables and the equipment settings. In order to understand such a complex process and also to be able to predict the process performance under randomly (ore grade, uranium mineralogy, gangue mineralogy etc.) and consciously perturbed variable settings (partial pressure of oxygen, temperature, concentration of leachant, stirring speed and time), sound and robust mathematical models are needed with analytical and/or numerical solutions. Present investigation attempts to develop a variant of kinetic correlation based on shrinking core model for batch alkaline leaching process in an autoclave. The leaching model incorporates particle size distribution according to popular Gates-Gaudin-Schuhmann equation, rather than assuming a single mean size for all the particles participating in the leaching reaction. Kinetic experimental data generated in autoclave leaching of Gogi uranium ore was used to evaluate various constants of the model developed. The ore sample used for experiments assayed 0.21 per cent U3O8. The partial pressure of oxygen was varied between 0.15 to 6.5 atm, temperature from 343 K to 438 K, concentration of sodium carbonate leachant in the range 0.47 g mol/L to 0.94 g mol/L, and the stirring speed from 573 rev/min to 900 rev/min in the experiments. The goodness of fit of the leaching process model is found to be >0.95. Morphological studies carried out on leached and unleached particulate matter using Scanning Electron Microscopy with Energy Dispersive X ray analyses are included in the present study.

Sep 1, 2012

By Qiuyang Hong, Bo Li, Dongyun Liang

"A weathering-crust niobium-rare earth-bearing deposit is formed by the weathering alterationof alkaline granite, which is refractory due to its low grade and complex properties. The mineralogicalcomposition of ore, the grain size of majorminerals as well as the occurrence of Nb and REE are studied in this paper using MLA combined with traditional methods. According to the results, a wide range of Nb-Ti-REE compound minerals including titanoniobite, uranpyrochlore, chlopinite, samarskite and ilmenorutile areformed for the extensive isomorphous substitution among the elements of Nb, Ta, REE, Fe, Ti, U and Th. The occurrence of REE is extremely complex. First, the REEexistingas rare earth minerals including monazite, xenotime, bastnaesite,etcaccount for 21%. Second, the REEexist in the lattice of Nb and Zr minerals as isomorphous additions, which account for 18%. Third, the REE exist as colloidal rare earthdeposit,which account for 35%. Last, the REEexistingas ion-adsorption-type rare earth,which are adsorbedon the kaolinite in the form of hydrated hydroxyl ions, account for 26%.Based on the results of mineralogy, the combined technology process "mineral processing- hydrometallurgical separation" is needed to deal with the ore. The mineral processing tests show that, for the feed ore containing Nb2O5 0.065%, REO 0.13%, an Nb-REE concentrate containing Nb2O5 1.38%, REO 1.98% can be gained by "high intensity magnetic separation-regrinding-gravity separation by shaking tables". The recovery of Nb2O5 and REO is 62.42% and 44.78% respectively. The Nb-REE concentrate is dealt by the metallurgy process "direct reduction- acid leaching & acidulating- neutralization & precipitation- chlorination & distillation" and the rare earth oxide of 92.34% REO as well as the Nb-Zr oxide of 24.62% Nb2O5 can be obtained with the recovery of REO and Nb2O5 being 76.83% and 59.37% respectively."

Jan 1, 2014

By Y. Q. Yang

Jig is a kind of conventional and old gravity concentration equipment, which is mainly used in the preconcentration of medium particle fractions of low-grade nonferrous metal ore. In the past ten years, studies on jigs have made a great progress on separation of coarse and fine particle fractions. At the same time, pneumatic jigs have been widely used in processing of nonferrous metal ore. In this paper, pilot test study and commercial operation for several kinds of jigs used in Dachang con-centrator has been introduced. Dachang mining area of GuangXi province is rich in mineral resources, where many minerals exhibit various complex properties. Many elements such as Sn, Pb, Sb, Zn, S, As, Cu, In, Ag, Gd, Ag, Ga, etc, were found in the run-of-mine ore (ROM). At present, Dachang concentrator produces tin concentrate, lead-antimony concentrate and zinc concentrate. Other valuable elements get enriched and recovered in above concentrates. Dachang Mine with a mining operation history more than sixty years will end its mission on the rich ore resource exploitation. At present, low grade ores are porcessed in the concentrator, and the concentrator begins to run at a loss upon processing the low-grade and refractory ore. Heavy medium cyclones were used in the concentrator for processing 20-4mm fraction, Gongge jigs for 4-2mm fraction, and cone concentrator combined with spiral concentrator to discard the tailings of -2mm fractions. In order to minimize overgrinding, table flotation was used to recover some liberated cassiterites in a complex processing technology. Therefore, the concentrator is difficult to manage. Many years studies have shown that heavy medium cyclones and spiral concentra-tors can be replaced by sawtooth shaped wave jigs to discard 50% tailings to improve the grade of feed. And the feed was divided into rich one and poor one to minimize overgrinding of valuable cassiterites with different grinding mills. In this case, the recovery of the valuable minerals can get improved and the flowsheet was simplified. Moreover, the concentrator begins to make profit with reduction of cost. At the same time, pollution in the concentrator gets reduced and the working condition gets improved too. Sawtooth shaped wave jig was used in Changpo concentrator for replacement of cone separators and spiral concentrators to discard the tailings of -4mm fraction, and the yield of discarded tailings got improved from 30% to 50%. Grades of tin, lead and zinc in the tailings are lowered from 0.15%, 0.12%, 0.7% to 0.06%, 0.05%, 0.4% respectively, with a loss reduction of tin, lead and zinc from11%, 9%, 10% to 7.5%, 6%, 5% respectively, corresponding to the increasement of recovery of tin, lead, zinc from 52%, 50%, 55% to 56%, 56%, 62% respectively. Cost of the concentrator for processing one-ton ROM was therefore dropped from RMB155.28 to RMB119.74. Changpo concentrator lost RMB2.5 million in 2003, but it made a profit of RMB0.92 million in 2006 after the innovations. Batac jig was used in Tongkeng concentrator to discard the coarse tailings at 20-4mm, with a discarded yield of 35%, and the grades of valuable minerals got improved with a recovery above 95% for tin, lead-antimony and zinc.

Jan 1, 2014

By Mikhail I. Khersonskaya, Mark I. Mantsevich, Andrey V. Tarasov

"Due to depletion of the existing ore deposits the content of non-ferrous metals in produced ores has an inevitable tendency to decrease. Processing of low-grade ores is associated with production of low-grade selective concentrates and poses serious problems for the non-ferrous industry relating to increasing losses and expenses for waste treatment and disposal.The main component contaminating selective concentrates is iron sulphides.Copper-nickel, copper-zinc and polymetallic ores have been discussed as most important types of ores requiring removal of iron sulphides.One of the promising directions for development of selective flotation technology is the use of lowmolecular organic depressors. An advantage of organic compounds is in the possibility of multi-variant structure of organic molecules, which permits selection of a most efficient reagent for the given conditions. Another way for improving the grades of concentrates is concentration of valuable constituents associated with iron sulphides in an intermediate product and upgrading of the latter by chemical and metallurgical techniques.In order to remove iron from nickel concentrates, specific process conditions have been developed for selective flotation of pentlandite with the use of an organic depressor for pyrrhotite. The use of a depressor for flotation of high-grade copper-nickel ores ensures an increase in the nickel content up to 14% in the nickel concentrate produced as compared with 7% when using conventional process conditions, along with a nickel recovery improved by 2-3%."

Jan 1, 2003