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By M. Brennan, A. Majumder

"The mineral processing research community has devoted enormous efforts to the modelling of mineral separators over the last 50 years. While many efforts have been made to use only fundamental physics as a basis for the models, fluid flow patterns in separators are not simple while the motion of particle swarms through fluids is extremely complex.One of the simplest separators is the sluice, in which separation takes place in a flowing film a few millimetres thick. The particles are suspended by turbulent eddies and motion relative to the fluid takes place in so-called hindered settling mode.In an effort to test the extent to which fundamental fluid mechanics can provide a prediction of particle separation, a study on a simple parallel-sided sluice has been carried out using glass beads at various particle concentrations and flow rates.The actual particle separation taking place was compared to predictions made by first describing the fluid flow by a modified ‘law of the wall’ model.Particle concentration distributions were described using Hunt’s sediment transport model, assuming the particle eddy diffusion coefficient to be the same as that for the water. Settling velocities were calculated using either the Richardson and Zaki or Brauer and Thiele models. The latter takes into account particle size and density distributions and solids volumetric concentration.The resulting predictions of particle separation are remarkably good, even though the model does not model the local settling velocity in the slurry."

Jan 1, 2003

By Gülay Bulut, Ünzile Yenial

"Water composition has a significant effect on flotation. A better understand the role of water composition in flotation is very important for process quality and performance. Based on source of process water, the concentrations of the dissolved ions, hardness, contaminants and residual composition of reagents can varied in the process water. The various water resources employed during lead and zinc flotation caused activation, depression or selectivity effects on minerals due to the presence of different ions in solution. Metal ions in these water resources were adsorbed on the ore, whilst ions associated with gangue minerals were released into the flotation pulp.This paper includes a research for the effect of different water compositions on lead and zinc sulphide flotation. Experiments were carried out five types of water such as tap water, recycling water, tailing dam water, underground mine water and sea water. The ore sample was obtained from Esan Balikesir-Balya mine. The mineralogical study of ore sample shows that it contains mainly galena, sphalerite, pyrite, hematite, quartz, diopside, epidote, calcite, and chlorite. It is found that recycling water and tailing dam water has a positive effect on recovery of lead and zinc minerals and detrimental effect on content of both minerals. This result may explain with remaining reagents in these waters. Underground mine water and tap water gave nearly the same recoveries of zinc. Sea water affects the recovery lead and zinc negatively."

Jan 1, 2014

By Ronel du Plessis, John C. Davidtz, Jan D. Miller

Xanthates (dithiocarbonates) have been the collector of choice for the flotation of sulfide minerals ever since the flotation process was developed and have been studied quite extensively. Recently, trithiocarbonates have been shown to be very effective in bulk sulfide bench-scale flotation. Fundamental analysis using electrochemical techniques have supported these bench-scale results, illustrating that the trithiocarbonate (TTC) collectors are more readily oxidized to their corresponding dithiolates, than the dithiocarbonate (DTC) collectors, as is shown in Figure 1.

Jan 1, 2003

By Nóra Schreithofer, Kari Heiskanen

"INTRODUCTIONThe knowledge of the adhesion forces between particles and bubbles is of key importance in the improvement of flotation models. The development of the Colloidal Interaction Force Measurement Apparatus (CIFMA) opened the possibility to quantify the adhesion forces between particles and bubbles. The aim of this study was to examine the effect of three main variables: the approach velocity, the loading force and the loading time on the adhesion force between an air bubble and a hydrophobized silica particle."

Jan 1, 2003

By lves Matiolo, Jorge Rubio, Carlos Perez Guerrero, Denise G. Nunes, Fabiano Capponi

The problem (very old) of processing by flotation, the fine (“F”, 38-13 µm) and ultrafines (“UF”, < 13 µm) mineral particles continues to be a major technical challenge in the area of mineral processing. Many different processes have been proposed over the last three decades but, unfortunately, most of them have not found practical applications yet. This paper describes some advances in flotation research in fines (“F”, 38-13 µm) and ultrafines (“UF”, < 13 µm) mineral recovery by two different techniques. Extender flotation (EF) and flotation with bubbles with a wide size distribution-SB of copper sulphide ore fines (1 % Cu) were compared with a “standardised” (STD) mill laboratory procedure. Comparative evaluation was made through concentrate grade and recovery (size by size results included), “true” flotation recoveries and process kinetics. Copper porphyry samples, containing about 1/3 of the mass and about 37 % of the total copper content in the fines and ultrafines fractions were used. In the EF tests diesel-in-water emulsions were used as an extra reagent, added after the collector and frother addition. In SB tests, drawing air through an aspiration nozzle, directly in the pulp, generated bubbles sized about 50-200 µm. Both techniques yielded higher metallurgical efficiencies than the mill standard but better results were obtained with the extender flotation, with emulsified diesel oil (90 gt-1). Concentrate recoveries were of the order of 85-90 % (84 % in standard conditions), 14-17 % Cu grade (15 % Cu in the standard) and 3.5-4.5 min-1 for the Klimpel rate constant (3 min-1 in the standard). Size by size and true flotation results proved that metallurgical and kinetic gains, obtained with the two techniques, might be attributed to fines and ultrafines enhancement. Data obtained are discussed in terms of physical and physicochemical parameters related to aggregation and capture (of particles by bubbles) phenomena and the practical potential of these techniques are envisaged. The following figure shows size-by-size copper recoveries results, after classification of produced concentrates and tailings using micro-sieves. Results clearly show that flotation of the F and UF fractions is highly improved with injection of small bubbles (SB) and with the addition of an extender collector. It is believed that the techniques presented here will establish shortly a new basis for the flotation enhancement of the F and UF, aimed at solving a very old problem. This work is being continued, now at pilot plant scale, and results will be published en the near future.

Jan 1, 2003

By Ashok M. Raichur, S. P. Vijayalakshmi

"Recent advances in technology and the growing environmental concern have laid emphasis on the usage of fine clean coal. Coal fines are also generated in large amounts during various coal processing operations and are difficult to be treated, handled and disposed. Since most of the operations use water as the carrying medium, the fines so generated contain lot of moisture. Conventional technique for processing coal fines includes flocculation by a polymeric reagent followed by moisture removal using a rotary drum or a pressure plate filter, which produce a product with more than 30% moisture. Polymeric flocculants though required in small amounts, need high shear mixing for good performance and operate only in a narrow pH range. Moreover they increase the viscosity of the solution by many folds. Hence it becomes imperative to emulate newer methods for efficient fine coal processing. One such method is the novel use of microorganisms in coal processing.The present investigation deals with the utilization of two gram-positive microorganisms, Mycobacterium phlei and Bacillus subtilis, as bio flocculants and dewatering aids in fine coal processing. Preliminary characterization done on J-coal and M.phlei showed that they are negatively charged over most of the pH range. B.subtilis was found to possess positive charge below pH 4.7. The adhesion studies proved the high potential of microorganisms for adsorption to surfaces. Adhesion tests were conducted at various experimental conditions by varying parameters like pH and ionic strength of the suspension. Results obtained showed that adhesion is always higher in the presence of an electrolyte like KNO3 and most of the adhesion takes place in 15 minutes. With change in pH the amount of B.subtilis adhering on to J-coal remained almost constant and for M.phlei it decreased at higher pH values. Increase in ionic strength did not increase the adhesion of bacteria. Flocculation experiments confirmed the ability of the bacteria to act as bio flocculants. It was seen that about 98% of the coal settled in about 2 minutes on the addition of bacteria along with the addition of the electrolyte. The electrolyte was found to be indispensable for efficient flocculation of the coal suspension and for the attainment of a clear supernatant after flocculation. Also the optical micrographs revealed the formation of large, tight flocs when the electrolyte is present along with the bacteria. Further, flocculation tests were done in the pH range of 2-10 and the flocculation efficiency was found to be independent of pH."

Jan 1, 2003

By Sonia Helle, Ursula E. D. Kelm

"From a metallurgical point of view, acid leaching of oxide zone copper ore is considered in the first place in terms of ore grade, whole rock mineralogy, rock granulometry, crushed aggregate size distribution, porosity, permeability and corresponding fluid flow characteristics. Despite exhaustive physicochemical and general mineralogical characterization, drops in copper recovery occur and are often attributed to and “unquantified mineralogical misbehaviour” of the gangue phases. This study establishes an experimental link to monitor and quantify the chemical interaction of chrysocolla, malachite and atacamite with reactive gangue phases.Laboratory-scale sulphuric acid (15%) leaching in small columns is carried out using a synthetic rock composed of 57% quartz, 1% ore phase, and 42% reactive gangue. Gangue phases were selected to represent mineralogy found in argillic, advanced argillic, and propylitic alteration zones. Exotic copper ores of Codelco Norte (in particular Mina Sur) served as model for the modal distribution of mineral phases. Natural reactive gangue phases (<200 Tyler mesh) were chosen as pure as possible: Saz-1 and Swy-1 montmorillonites, KGa-1B and KGa-2 kaolinites (all Clay Mineral Society, USA, Source Clays) Nau-2 nontronite, Imt-1 illite (both Clay Mineral Society Special Clays) and Chilean mordenite, laumontite and bentonite. Ore phases are: mineraloid chrysocolla from Radomiro Tomic (Codelco Norte), reagent grade malachite and a purpose synthesized atacamite-paratacamite mixture. Complete chemical, physicochemical and mineralogical characterizations were carried out for ore and gangue phases.The following observations hold for samples with and without mineralization: Smectite-rich gangue increased up to 75% in volume. Once wet, air-filled vesicles formed due to the rapid sealing effect of overlying smectite in the column. No volume change occurred for illite, mordenite and kaolinite gangues. For kaolinite mixtures acid puddles formed on top of the column during initial leaching due to the delamination and barrier formation of this phyllosilicate. Fluid recovery was <60% for expandable phases and reached up to 80% for non-expandable mixtures. These observations confirm the very low extractions obtained in ore with a combined smectite – kaolinite gangue for Mina Sur. The most important mineralogical changes observed in the solid residues are: reduction of basal smectite phase reflections (XRD), and slight loss of illite intensities. This is also reflected in the major element composition of the residues, which demonstrate a loss of Mg, Mn, Ca, Na, K followed by Fe, Al and Si; with their proportions depending on the starting materials. Furthermore, alunogen and meta-alunogen formed, with related phases coquimbite (illite) and godichite (nontronite) also depending on gangue mineral chemistry. For laumontite samples, meta-alunogen and alunogen precipitated from the leachate."

Jan 1, 2003

By Zoran M. Stevic, Mirjana M. Rajcic-Vujasinovic, Zvonimir D. Stankovic

"Covellite is one of the most important copper sulfide minerals. Its behavior during the anodic polarization, as well as its rest potential in various solutions are of importance for flotation processes and also define the mechanism of electrochemical extraction of copper from sulfide minerals.Measurements of the rest potential have been performed in water solutions in presence of different cations and anions. The existence of the Nernst type dependence of the rest potential on the activity of some of the investigated ions was established. The biggest influence on the rest potential of covellite exhibits the change of activities of the ions forming the crystal lattice of investigated mineral, i.e. Cu(II) and S(II). The dependence of the rest potential of covellite on logarithm of Cu(II) activity has a shape of a strait line with a positive slope of 60 mV, while when concentration of S(II) anions increases ten times, the rest potential of covellite decreases about 120 mV. The first dependence was established in acid medium, while the second was registered in alkali one.Different electrochemical techniques have been used for the investigation of the behavior of covellite during anodic polarization. In the basis of all these measurements is a three electrode system consisted of: 1) working electrode made of the sample of natural covellite from the Bor mine (Eastern Serbia, Yugoslavia); 2) reference electrode (saturated calomel electrode) and 3) counter electrode (flat platinum). Electrolyte solutions used in these investigations are of the same kind as those used in the rest potential measurements. Sulfate and chloride solutions, as well as the mixed sulfate/chloride solutions, are the most perspective from the application point of view.Cyclic voltammetry has been the first used method that confirmed the formation of passive layer on the electrode surface during anodic polarization. Elementary data about the reaction mechanism have been obtained too. More detailed data needed for more precise conclusions about the mechanism and rate of the anodic process on covellite have been obtained by galvanostatic method. It was established that the essential difference exists between the reaction mechanism in sulfate and chloride solutions. Also, the difference is exhibited in the nature of reaction product directly connected with the passivation of the electrode surface. Chronopotentiometric investigations have been performed in sulfate, nitrate, chloride perchlorate and mixed sulfate/chloride solutions applying different current densities. The shape of obtained curves indicate if there was a surface passivation and how intensive it was."

Jan 1, 2003

By Lev O. Filippov, Inna V. Filippova

"Flotation of metal-organic precipitate in column is sensible to the hydrodynamic conditions determined by the diameter and dispersion of rising bubbles. Thus, the dissipation energy in the system resulting of ratio between superficial fluid velocities (gas and feed) in the column is main parameter for precipitate flotation in pilot or industrial scales. The knowledge of aggregate structure can lead to better understanding and control of many solid-liquid separation processes (sedimentation, filtration, thickening and flotation). Many authors highlight the importance of structural properties (i.e. floc density) for different industrial applications.This approach was applied to achieve the flotation of molybdenum precipitate obtained with organic collector a- benzoin oxime for molar ratio 1:2. Flotation tests were carried out at lab and pilot scale with rigorous control of the size distribution of bubbles and the values of dissipation energy in the systems.Structural evolution of Mo precipitate and the destruction phenomenon were described by using of the various fractal dimensions, which were deduced from experimental data of laser sizer devices and optic microscopy pictures of precipitate. The boundary fractal dimensions Df of large aggregates allows the detection of a structural scaled organisation of the precipitate flocs (elementary germs-microflocs-flocs). The fractal dimension of the particle mass distribution Dm=1.94+/-0.15 suggests, for the floc structure, a cluster-cluster aggregation process either with structuration or breakage phases.The destruction process characterised by a fractal dimension Dg controls the final size distribution. The floc destruction occurs for the lowest values of the dissipation energy (0.02 – 0.05 W/kg). The relation between Dg and the system dissipation energy Et (for various energy E dissipated in the agitation cell) involves the different breakage mechanisms of aggregates in function of the aggregation bonds (Fig. 1).Besides, the destruction of floc produces very fine particles (elementary germs or/and microfloc according to E), and thus induces the difficulties of bubble-particle attachment during the precipitate flotation in column. Yet, very small bubbles, which increase the collision probability, would make the flotation feed rate decrease. The optima ratio between superficial gas and feed velocities are determined for bubble size and flotation results control at pilot scale. Thus, it is possible to providing the catastrophically changes of precipitate column flotation by using the following established experimental relation: the mean diameter d of the flocs under the action of dissipated energy Et follows a power like law (d~Et -0.5). This allows estimating the size distribution of flocs for the E values which can appear in column."

Jan 1, 2003

By V. Dementiev

Pyrite cinders is a complex raw material, in which the valuable metals are: Au(2.04 g/t), Ag(35.0 g/t), Cu(0.38%), Zn(1.1%) and Fe(56.0%). In this case high costs for mining and ore processing are not required. The hydrometallurgical treatment of pyrite cinders using thiourea, sulfite and cyanide solutions are examined in this paper. It was found that thiourea leaching is ineffective due to high consumption of thiourea and sulfuric acid. The cyanidation on pyrite cinders was conducted. Various options of the preparation of the products for the subsequent cyanidation have been examined. The tests on cyanidation process optimization were conducted. The effectiveness of these cementation methods by metallic iron or extraction for copper recovery from water phase has been shown. Zinc precipitation can be found in the form of carbonate. In order to recover noble metals from cyanide solutions, the cementation onto zinc dust was tested. The developed technology provides for recovery at 49-56%Au, 20-25%Ag, 22.5%Cu and 50%Zn during NaCN (1.4 kg/t) and CaO (6 kg/t) consumption. The results of pilot plant testing are the basis for technical regulations in order to design a pyrite cinders treatment plant. The technology has been examined in pilot-plant conditions. In this case the recovery of 54.5%Au, 45.3%Ag, 30%Cu and 49.9%Zn was achieved to provide saleable products. Keywords: pyrite cinders, thiourea leaching

Sep 1, 2012

By Erico Tabosa, Kym Runge, Peter Holtham

"The distribution of turbulence in an industrial scale mechanical flotation cell and its effect on flotation kinetics was studied using a fully instrumented 3m³ cell. The cell was operated at a copper concentrator in Australia with different combinations of airflow rates, impeller speeds and sizes and cell aspect ratio providing a wide range of hydrodynamic conditions. An analysis of the overall flotation cell performance showed that the overall copper recoveries were very similar for the conditions tested. However, by decoupling pulp effects from froth effects it was possible to evaluate whether the changes made affected the pulp zone and/or froth zone responses. The analysis showed that the overall recovery has the potential to be up to 7% higher, if not limited by froth recovery. The cell hydrodynamics were characterised by the overall energy dissipation rate and, to understand the distribution of turbulence within the cell, localised measurements were made using a piezoelectric vibration sensor. By comparing the metallurgical performance of the cell with the different hydrodynamic conditions it was found that the collection zone flotation rate constant was not directly related to overall energy dissipation, as is commonly observed at laboratory scale, where energy is usually only changed by varying impeller speed. In this study, energy dissipation rate was varied using not only impeller speed, but also the size of the impeller and the cell aspect ratio. In these circumstances, the best correlation was found to be between the collection zone flotation rate constant and the distribution of turbulence, i.e. the size of the turbulent zone."

Jan 1, 2014

By David Deglon, Martin Harris, Mehdi Safari

This paper investigates the effect of energy input on the flotation of galena in a novel oscillating grid flotation cell (OGC). Oscillating grids generate near ideal hydrodynamic environments, characterised by turbulence which is relatively homogeneous and isotropic. Previous work was limited to quartz (-100 µm). Galena (-150 µm) was floated in the OGC at power intensities from 0.5 to 3 W/kg using 0.13, 0.58 and 0.82 mm mean bubble sizes (d10) and at three collector dosages. Results show that the effect of power intensity on flotation kinetics is strongly dependent on both particle and bubble size. For fine particles (-19 µm), increasing energy input increases the flotation rate at a low collector dosage but leads to an optimum at both the moderate and high collector dosages. For intermediate particles (+19-38 µm), increasing energy input has no effect on the flotation rate at the low collector dosage but leads to an optimum at both moderate and high collector dosages. For coarse particles (+38-150 µm), increasing energy input has no effect on the flotation rate at the low collector dosage and leads to a general decrease at both moderate and high collector dosages. Optimum conditions for fine particles are using small bubbles at a moderate collector dosage and an energy input of around 2 W/kg. Optimum conditions for moderate particles are using small bubbles at a high collector dosage and energy inputs of around 1.0 W/kg. Coarse particles do not benefit from agitation.

Jan 1, 2014

By T. Suzuki

Acid leach liquors from spent lithium-ion batteries employing lithium cobalt dioxide (LiCoO2) as a cathode active material contain valuable metals such as cobalt and lithium, as well as common metals such as aluminum and copper. The objective of this study was to develop a hydrometallurgical process providing a high selectively between aluminum, cobalt, copper, and lithium from hydrochloric acid solutions. Batch experiment results showed that selective hydroxide precipitation approach was not suitable because a significant amount of cobalt and lithium was co-precipitated with aluminum and copper hydroxides. In contrast, batch experiments and the subsequent data analysis indicated that the sequence of solvent extraction circuits using PC-88A, Acorga M5640, and tri-n-octylamine (TOA) extractants effectively separates four metals of interest. In the first circuit of the proposed process, copper was selectively extracted by Acorga M5640 at the pH 1.0-2.0. Aluminum, cobalt, and lithium remain in the aqueous phase. In the second circuit, aluminium is selectively extracted by PC-88A at pH 2.0-2.5. For the separation of remaining cobalt and lithium, Acorga M5640 provided higher separation efficiency compared to PC-88A single extractant and PC-88A/TOA mixed extractant. However, the stripping efficiency of cobalt from the Acorga M5640 organic phase was found to be low (less than 10 %). Therefore, PC-88A/TOA mixed extractant rather than Acorga M5640 is an advisable choice for the separation of cobalt and lithium. Keywords: spent lithium-ion secondary batteries, solvent extraction, lithium, cobalt

Sep 1, 2012

By Renato S. Costa, Afonso H. Martins, Flávia D. Mendes

"This paper presents some important aspects regarding the ion exchange treatment of an Ni and Co sulphate acidic liquor using chelating resins. Chelating resins possessing iminodiacetic or bis-picolylamine functional groups were compared in terms of nickel and cobalt recovery. The experimental resin with bispicolylamine functionalities (Dowex M4195) was evaluated with respect to Ni and Co uptake, and compared with Rohm &Haas commercial chelating iminodiacetate acid resin.The use of selective ion exchange to recover metals from sulphuric acid liquors has been studied for a long time. Companhia Vale do Rio Doce (CVRD) has been actively involved in hydrometallurgical applications of IX and has researched in selecting the optimum resin for nickel and cobalt selective removal and the design of a suitable IX circuit. When an IX operation is contemplated for a specific application, it is followed a screening process which generally involves consideration of the solution composition and pH, speciation of the metals in the liquor to be treated and other possible techniques that can be applied to achieve the desired specifications or recoveries. Once a decision has been made to use IX for a specific application, relevant testwork should be done to select the most cost-effective resin and design the IX circuit to achieve the desired criteria. The experimental part dealt with the working conditions, resin characteristics and the influence upon the resin maximum exchange capacity of the following factors: nickel concentration, temperature, pH and resin-to-solution ratio. Based on the experimental data obtained, it has been proposed a innovative technological treatment"

Jan 1, 2003

By Ibrahim Alp, Duygu Dinç, Ersin Yazici, Adularya Energy and Electricity Generation and Mining Inc., Oktay Celep, Haci Deveci

Thiosulfate leaching as an environmentally benign alternative to cyanide leaching has received particular attention in recent years. Most studies have focused on thiosulfate leaching of free milling ores and, to some extent, pregrobbing ores. In this study, leaching of gold from the refractory Kaletas ore (in which gold/silver alloy is the only mode of occurrence detected) in ammoniacal thiosulfate solutions was studied. A two-level full factorial design (23) was adopted to investigate the effects of concentration of thiosulfate (0.2-1 M Na2S2O3), ammonia (0.5-1 M NH3) and copper (0.05-0.1 M CuSO4) on the extraction of gold from a ground ore (-38 µm) over a period of 24 h. Statistical analysis confirmed the leaching of gold improved with increasing concentrations of thiosulfate and ammonia. In contrast, increasing the concentration of copper in the range tested adversely affected the extraction of gold. The magnitude of effects of parameters on the gold extraction was in the order of Na2S2O3 > NH3 > CuSO4. The maximum extraction of gold (69.6%) was obtained in the presence of 1 M Na2S2O3, 1 M NH3 and 0.05 M CuSO4. The negative influence of copper on gold extraction is most likely attributed to the decomposition of thiosulfate catalyzed by copper. Concentration of thiosulfate was also identified to be statistically the most important parameter for its consumption.

Jan 1, 2014

By H. J. Liang

In this study, zeta potential and contact angle measurements of pyrite before and after interacting with Thiobacillus ferrooxidans are carried out to investigate the changes in its surface properties. The results show that, after pyrite interacting with Thiobacillus ferrooxidans and its metabolite, the surface electricity of pyrite changes obviously, resulting in a notable shifting of isoelectric point to lower pH value, but the contact angle varies little. Bacterial protein plays the most important role in the interaction. This work is helpful to disclose the mechanisms of Thiobacillus ferrooxidans de-pressing pyrite flotability.

Jan 1, 2014

By Q. R. Chen

Wet processes are commonly used in coal preparation field home and abroad. Dry coal cleaning technology is very necessary in China because more than 2/3 of coal resources deposited in arid areas, which are difficult to separate coal with wet processes. A particular interest of this article is the discussion of separation research development of dry coal cleaning in China, such as an air dense medium fluidized bed (ADMFB) beneficiation technology, a compound dry (pneumatic) separation technology and a triboelectrostatic separation technology. The fundament and importance of dry coal cleaning methods above were introduced. On the basis of analyzing parameters which influence the performance of separators, including that of structure, technologic and operating, the separation research status of dry coal cleaning methods, especially the ADMFB beneficiation technology, were discussed and reviewed. The recent developments, especially the fluidized bed beneficiation for <6mm fine coal with an external energy assisted, including a vibrated fluidized bed and a magnetically fluidized bed, and the triboelectrostatic separation for <1mm fine coal are also introduced. The study and application results in China show that dry coal cleaning technologies can efficiently beneficiate different coals with different sizes. The ADMFB can separate >6mm coarse coal with an E value of 0.05~0.08, the external energy assisted fluidized bed can separate <6mm fine coal with an E value of 0.07~0.09, the triboelectrostatic technology can efficiently separate <1mm fine coal, and the compound dry separation technology can efficiently separate >3mm coal with a easy washability. The dry coal cleaning provides an efficient and lower cost clean approach for coals distributed in arid areas.

Jan 1, 2014

By M. Amani, A. Akbarinasab, M. Bazmandeh

The fourth iron ore production plant at Golgohar mining and industrial company consists a ball mill and wet medium intensity magnetic separators (3500 Guss). The ball mill product applies directly into MIMS and the MIMS concentrate directly transfers to the low intensity magnet separators (LIMS). The thickener is used for dewatering of MIMS and LIMS tailings. By sampling from streams of ball mill and separators, it was defined that the grade and iron recovery were 60% and 80% respectively while they would be 66% and 90% respectively based on the designing documentation. The average of P80 in the ball mill was 342 micron in a month that is coarser than optimized value. This problem was related to unsuitable operation of grinding. For optimization of grinding efficiency, the amount of charge volume was investigated as a variable. The measurements illustrated 32% charge volume. For increasing grinding efficiency, the charge volume was increased 1% and 1.6% in two stages. The results showed ball mill power draw average reached from 3650 kw to 4120 kw and the average P80 reached to 329 micron. All of parameters calculated about 35 day before and after change. Average of Ball mill feed iron ore grade before and after change was 49.24% and 50.32% respectively. This change caused an increase in efficiency of MIMS, where the enrichment ratio and concentrate grade increased 2% and 1.9% respectively. For increasing the iron grade of the MIMS concentrate; the orifice of the tail flow exit pipe changed from 4 inch to 6 inch and the concentrate grade increased 2%.

Jan 1, 2014

By Alejandro Rubio-Lopetegui, Gonzalo Montes-Atenas

The occurrence of Enargite (Cu3AsS4) in hydrometallurgical plants bears a number of challenges. One question lies on how Copper and Arsenic can be selectively extracted from the solid surface structure when immersed in aqueous solution. The leaching of Enargite in ambient conditions has been extensively studied; nevertheless, challenges regarding the process kinetics and surface structures enabling the extraction process remain ill-understood. Prior knowledge indicates that different conditions of pH, Eh, temperature, speciation of the aqueous phase or variations of these parameters defines the overall rate at which metals and other elements will be transferred to the solution. One aspect that the research studies have seldom addressed involves studying the main features of the surface structures formed during the leaching process at the outmost surface atomic layer. Such structures will mediate between the solid bulk and the solution to undertake subsequent stages of leaching process. The aim of this research work is to critically analyze the state of the art of Enargite electrochemistry from the mediator formation-dissolution mechanisms standpoint occurring at the mineral-solution interface. The most probable mediator species reported in the literature were found to be: Arsenolite and Sulfur (acidic media), Covellite (acidic and alkaline media) and Tenorite (alkaline media). As a result of this analysis, it was observed that the surface speciation was not in full agreement with the previous knowledge suggested to explain Enargite leaching reactions and, therefore, modified Eh-pH diagrams are presented. The acquired insights on the surface structures rising from the Enargite-aqueous solution interactions is further revised to provide new experimental protocols towards selectively separating Copper and Arsenic through hydrometallurgical paths.

Jan 1, 2014

By Andrzej Luszczkiewicz, Alicja Bakalarz

The application of new collectors is considered to improve the beneficiation of Polish sedimentary complex copper ores from the Legnica-Glogow Copper Basin (SW Poland), the upgradeability of which becomes more difficult because of ores grade declining. One group of potential collectors is represented by nonpolar reagents. According to the available literature data, only selected sulfides float in the presence of the nonpolar liquids. The paper presents laboratory scale flotation results for the main sulfide minerals occurring in the Polish copper ores applying either thiol (potassium ethyl xanthate) or nonpolar (aqueous emulsion of n-dodecane) collectors. The investigated feed contained 1.91 % Cu in the form of sulfide minerals such as bornite, chalcopyrite, chalcocite and covellite. Other sulfide minerals including galena, pyrite and sphalerite were also present in the samples. Several flotation tests with various doses of each collector were conducted. Mineralogical analyses of flotation products were carried out and the results compared. A comparison of the results indicated that all sulfide minerals floated well with high recoveries ranging from 91 % to 99 % in the presence of the thiol collector. The use of the nonpolar collector provided results in flotation of sulfide minerals with different terminal recoveries from 61 % for chalcopyrite to 100 % for sphalerite. The most significant differences in flotation in the presence of both types of collector under various conditions of upgrading were observed for chalcocite, chalcopyrite and covellite. Similar flotation results were observed for sphalerite and galena for both collectors. The studies confirmed selective flotation of sulfides from the copper sulfide ores of the Legnica-Glogow Copper Basin in the presence of nonpolar collectors.

Jan 1, 2014