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By S. Banisi, A. Ghasemi, S. Zare, F. Ghorbani, E. Nematollahi

"Comminution is the most energy intensive operation which constitutes the major portion of operating and capital costs of the mineral processing plants. Working at the maximum operating capacity of comminution equipment plays a significant role in the efficiency of the circuit. Appropriate design of cone crusher feed box could result in the choked crushing which is known as a vital factor in achieving target performance. In this research, the non-uniform cone crusher feeding at the Sarcheshmeh copper complex was studied and appropriate measures were taken to solve the problem. The problem resulted in non-uniform and high wear rate of crusher liners and high fluctuations of power draw. In recent decades, the discrete element method (DEM) has been applied successfully to study the effect of different factors on the behavior of particles flow. In this work, various designs for feed box were tested using an in-house developed DEM software called KMPCDEM©. The best feed box design obtained from the simulations was selected, constructed and installed in of the cone crushers. The new feed box design resulted in more uniform feed distribution on the distributor plate of crusher indicated by a reduction of the standard deviation of power draw measurements from 13 to 3 kWh. The better control of the crusher, 49% increase in the crusher throughput and finer and narrower product size distribution were the main outcomes of the installation of the new feed box. Furthermore, the wear rate on mantle and liners of crusher not only decreased but it also became uniform.INTRODUCTION Comminution is the most energy intensive operation which constitutes the major portion of operating and capital costs of the mineral processing plants. Reaching the maximum operating capacity in comminution equipment is a key sign of approaching the nominal productivity of the process. Due to the effect of crusher efficiency on the downstream circuit performance, optimization of the crushing circuits has received considerable attention (Herbst, & Potapov, 2004; Quist, & Evertsson, 2010; Li, Huiqi, McDowell, Glenn R, Lowndes and Ian, 2014; Cleary, Sinnott, Morrison, Cummins, & Delaney, 2017). Improving the performance of crushing circuits not only decreases the capital and operating costs but also increases the efficiency of downstream grinding circuits (Jacobson, Janssen, & Urbinatti, 2010). In cone crushers particles can be subjected to two different modes of compressive breakage; single particle breakage (free feeding) and inter-particle breakage (choked feeding). The maximum efficiency of the cone crusher directly depends on the type of feeding arrangement (Operation Manual, 1977). An appropriate feeding must lead to choked feeding the cone crusher. If the rate and method of feeding the crushers are correctly be executed, the maximum efficiency of the crushers could be available (Jacobson et al., 2010)."

Jan 1, 2018

By T. Leißner

This contribution deals with the mineral processing of Li-silicate greisen-type ores comprising of quartz, topaz and zinnwaldite (lithium-rich mica). The origin of the greisen-type ores processed is the Ore Mountains(Germany) where it is explored as a potential resource for the production of lithium carbonate. The goal is to develop a process chain for the enrichment of zinnwaldite using dry techniques only. As basis for the investigation the process chain, which historically was used to process the greisen focused on cassiterite and wolframite, is taken and modified towards the zinnwaldite. Starting with crushed material with particle sizes smaller than 35 millimeters, investigations on different approaches to grinding to liberate mica from gangue are carried out. Concentrates of zinnwaldite are then produced by magnetic separation of size fractions. To assess the success of grinding, classification and separation, mineral liberation analysis and atomic absorption spectroscopy are used. The amount of lithium measured in the sample with atomic absorption spectroscopy was used to calculate the content of zinnwaldite based on its known mineral chemistry. Combined with the particle size distributions, product qualities are determined. Altogether, this allows the thorough evaluation of the success of comminution with focus on following steps of concentration. Keywords: lithium, zinnwaldite, magnetic separation, mineral liberation analysis

Sep 1, 2012

By S. M. Javad Koleini

The aim of this research is to extract zinc from complex lead and zinc ore containing carbo- silicate lead and zinc minerals using sulphuric acid. Samples, taken from Dareh Zanjireh mine located in central of Iran, were subjected to leaching tests after calcination process. Experiments were planed according to the Taguchi statistical method. The calcinated samples contain 13.98% zinc and the constituting minerals being dolomite and zincate. Various parameters such as speed of mixing, temperature, time, solid to liquid ratio and pH were studied. A solid to liquid ratio of 1:6 gr/mL at 60°C and 800 rpm speed of mixing for 30 minutes resulted in a product with 85.96% zinc recoveries.

Jan 1, 2014

The present work aimed to verify the possibility of using froth flotation to separate the constitu-ent minerals of an alaskite rock waste having as the main minerals quartz and feldspar. Three stages of tests have been carried through. In the first stage was used an amphoteric collector for flotation, diaminedioleate, at pH 2 and, after analyzing the final products, a very poor selectivity was ob-served. During the second stage of tests, different collectors that are normally used for feldspar flo-tation as amines, sulfonates and soybean oil were employed. Once more, it was not possible to reach reasonably selective results. In the third stage, a salt of etheramine was used for collecting the feldspar and hydrofluoric acid was utilized for controlling the pH. A simulation of three flotation steps (rougher, cleaner and recleaner) permitted to attain a good selectivity in the flotation process. A concentrate with a silica content of 96.6% was obtained. The low concentrate content of alkalis totalizing 0.52% and a feldspar recovery of 96.5% to the tailings, represented by the high content of Na, K, and Al2O3 in the tailings, demonstrated that the concentrate was made up mostly of quartz.

Jan 1, 2014

By Atsushi Shibayama, Toyohisa Fujita, Wan-Tai Yen, Ke-jun Liu

"In this investigation, a sponge gold sample (99.92% pure) has been leached with ammonium thiosulfate solution. The effects of CuSO4, (NH4)2S2O3, (NH4)2SO4, NH4OH and retention time on the gold leaching % (extraction %) and thiosulfate oxidation % (consumption %) have been studied. The copper ion acts as a catalyst in gold dissolution. On the other hand, it also accelerates oxidation of thiosulfate (consumption). It was found in this study that the best gold extraction condition was obtained at 43.2 ks with 0.8 - 1.0x103 mol/m3 (NH4)2S2O3, 30 mol/m3 CuSO4, 2.0 - 3.0x103 mol/m3 NH4OH, 0.4 - 0.5x103 mol/m3 (NH4)2SO4, pH 10±0.2 and 150 rpm. The thiosulfate oxidation was about 15% of the initial concentration.For gold recovery from the pregnant solution, the trioctylmethyl ammonium chloride (TOMAC) was used in this investigation as an extractant to extract gold from thiosulfate solution without pH and other adjustments. The effects of gold concentration, pH, ammonia, thiosulfate and cupric ionic concentrations, extractant and diluent concentrations and retention time on the gold recovery were investigated. It was found that 180 mol/m3 extractant diluted with n-octant at O/A ratio of 1:1 could extracted the 100% of gold in 0.6 ks from the thiosulfate solution at the optimal leaching conditions. The 100% gold stripping could be achieved with 10% thiourea solution."

Jan 1, 2003

By Evgeni Visariev

In Copper smelter Aurubis Bulgaria, the copper in slag from the Flash furnace and the Converters is recovered by flotation. The scheme of slag processing is crushing, two stage grinding (autogenous and ball) and flotation. Till 2010 the grinding cycle in the plant was in two stages ? the first stage was AG Mill 70?23 working with a classifier and the second stage was ball mill 3.6/5.5 working with a hydrocyclone. In 2010 a new AG Mill 65x78 with hydrocyclone was put in operation in order to increase throughput of the plant. The second stage of the grinding cycle is with the same ball mill. The Old AG mill processed 55-60 t/h slag with particle size -0.071 mm 75% or over 75 t/h with particle size -0.071 mm 35-40 %. The new AG Mill was designed for throughput 105 t/h with particle size -0.071 mm over 70 %. After the start up there were some problems with the equipment ? block up of AG Mill?s grate and pumps for feeding of the hydrocyclone. For optimization of the process a lot of general tests were made. One of the main problem was a wearing of the sand nozzles. Different materials (rubber, ceramic, metallic) for the sand nozzles were tested. The openings of the grate were changed. This paper presents: ? the optimal technological parameters of the old and the new working scheme; ? how the technological parameters of the new working scheme change when the hydrocyclone works with different nozzles; ? the effect of the change of the grate openings; ? the problems with the increase of the circulating load; ? a final decisions for improvement of the grinding cycle work. Keywords: spigot, wearing, classification, effectiveness, circulating load, autogenous, mill

Sep 1, 2012

By Zhazira Baltabekova, Nina Lokhova, Arailym Dukembayeva, Madali Naimanbayev, Zhantore Dzhurkanov

"Recycling of industrial waste provided they have enough valuable elements requires significantly less investment compared with ore processing. Promising man-made raw material of rare earth elements (REE) is phosphogypsum. Works performed on dismantlement of phosphogypsum only concern phosphogypsum processing from apatit. In this case, none of the developments was tested for phosphogypsum produced on the basis of phosphorus, chemical and mineralogical composition of which is markedly different from apatites. We studied phosphorites of the Karatau deposits (Kazakhstan). Mineralogically all the layers are composed of three main ingredients: phosphate, silica and carbonate, which account for 92-96% of the weight composition of ores and rocks. Of impurities there are pyrite, iron hydroxides, gypsum, fluorite. In addition to the actual composition of phosphate in phosphate ores there are apatite. Silica is presented by chalcedony and quartz, and carbonate by calcite and dolomite. The results of X-ray microanalysis showed point cluster of grains containing REE in samples of phosphate ore. Analysis of the grain showed that REE are in the form of phosphates and silicates, which according to numerous studies in apatites is not established. In this regard, phosphogypsum, waste of production, is different in composition too. Therefore, transfer of techniques proposed in the methods for extracting REE from phosphogypsum after processing apatite, did not give the stated results and required a new approach to solving this problem. It should be noted that the phosphogypsum produced by processing the Karatau phosphorite contains medium group of REE by 2.6, and heavy group by 7.8 times more than phosphogypsum from processing the Kola apatite. Given the shortage of medium and heavy REE research to develop ways to extract the lanthanides from phosphogypsum after processing phosphate rock is promising. The study contains results of studies on the extraction of REE from phosphogypsum after Karatau phosphorite processing."

Jan 1, 2014

By I. Yochev, A. A. Metelev, A. I. Stepanenko, A. P. Purgin, M. G. Viduetsky, A. M. Panshin, I. F. Garifulin

"Pneumatic flotation cells of the KFM series are intended for the concentration of various ores and technogenic materials. The key operating principle of the flotation cells of the KFM series is the preparation of pulp for flotation in blowers, offering the advantage of an accelerated flotation process. This report summarizes the results of the implementation and operation of the KFM cells at Russian and foreign projects in 2015-2017. Their implementation and operation experience confirms the high efficiency of the KFM cells, especially in the head of lead and zinc, copper and zinc, copper flotation processes:head KFM-600 flotation cell recovers from lead-zinc ore (Rudmetal AD modular concentrator) up to 52% of lead in a single stage (without cleaner flotation) in a salable lead head concentrate (grade of up to 76% lead, 2.3% zinc). The flotation cell proved to be much more efficient than originally expected in the process description in the design phase,flotation of the copper-zinc ore from Novo-Shemur deposit (PJSC Svyatogor concentrator) in two KFM- 1400M cells recovers up to 65% of copper and 70% of zinc in a high-grade bulk copper-zinc concentrate containing 12-13% of copper and 25-28% of zinc. The installation of the two cells with a combined capacity of 26 m3 allowed to decommission 4 chambers of FPM-16 pneumomechanical flotation cells with a combined capacity of 64 m3 with a simultaneous increase in the overall ore concentration performancea single KFM-1400M cell (13 m3) recovers from the copper ore from Volkov deposit (PJSC Svyatogor concentrator) 53% of copper into a high-grade copper head concentrate (33.0% of copper). For comparison, the pneumomechanical cell FPM-16 (capacity 16 m3) has almost half as low copper recovery (29.6%) into the head concentrate at a lower grade (30.5% copper)."

Jan 1, 2018

By Kazutoshi Haga, Batnasan Altansukh, Atsushi Shibayama

Results of flotation process combined with atmospheric leaching and high pressure leaching for recovery of copper from low-grade silicate-copper ore have been presented in this study. The low-grade copper ore contains a small amount of copper (0.42 mass %), iron (4.6 mass %), and silicon (38.3 mass%) and among many other metals (Al, Ca, K, Mg, Na, Ti and S). The main mineralogical constituents of the copper ore are calcium aluminosilicate (Ca(Al2Si2O8), amesite (Mg2Al2SiO5(OH)4) and gypsum (CaSO4×H2O), and quartz (SiO2) as resulted from the alteration of an igneous rock. This study has shown that about 92 % of copper can be recovered with a maximum copper grade of 4.2 mass% in a froth after 15 minutes of flotation at pH 8.5 when used potassium amyl xanthate as a collector, MIBC and Dow-froth 250 as frothers and calcium hydroxide as a pH adjuster. The images of bubbles in the cell during flotation process were recorded using a high-speed camera. The froth concentrate with average copper grade of 2.0 mass% was further processed by atmospheric leaching with a solution of sulphuric acid and hydrochloric acid, respectively and high pressure leaching with sulphuric acid solution. As a result, leaching rate of copper reached about 60 % in an atmospheric leaching with sulphuric and hydrochloric acid respectively, whereas the rate can reach more than 70 % by autoclave leaching at high pressure under the controlled condition. A process to recover of copper from its low-grade ore can be developed by optimization of flotation combined with acid leaching.

Jan 1, 2014

By R. Van Hille, S. Lacour, A. E. Lewis, K. Peterson

"This paper compares thermodynamic modelling results with those obtained for copper sulphide removal from aqueous solution in both continuous and batch experiments at laboratory scale. Sulphide precipitation has been demonstrated to be an effective method for heavy metal removal. The advantages are the high degree of metal removal at low pH, low quantities of sludge formation, the ease of de-watering and low reactor retention time due to high reactivity of sulphides. However, the low solubility of sulphide species can lead to spontaneous nucleation of fines, which are difficult to remove.Theoretical modelling was carried out using ASPEN Plus™ and OLI systems Inc ESP process. Both simulators were used to calculate copper sulphide recovery and residual aqueous copper concentrations as a function of pH. The predictions were then compared with experimental results. Both simulations predicted that all of the inlet copper would precipitate as sulphide CuS(s) over the entire range of pH and sulphide: copper molar ratios investigated. The simulation results were strongly influenced by the number of chemical species and the nature of the thermodynamic data taken into account in the model.The continuous experimental studies were carried out in a fluidised bed reactor (FBR). Copper sulphate was reacted with sodium sulphide and the solid precipitated directly onto the pellets in the FBR. The use of sulphide was motivated by high sulphate levels in local acid mine drainage and the ultimate integration of the metal removal process with a biological sulphate reducing process that produces an aqueous sulphide stream.A maximum dissolved copper removal of 99.6% (AAS analysis) and a total removal efficiency of 95.8% (dissolved copper minus fines) were achieved. Stoichiometry of the precipitate was determined from SEM/EDAX analysis (atomic percentages) and was close to 1/1, irrespective of the operating conditions."

Jan 1, 2003

By P. Hatefi, O. Rahmani, M. Noor-Mohammadi, A. Tohry

"Depressants play an important role in the selective separation of impurities from iron minerals by direct or inverse flotation. In this paper, sodium co-silicate (SCS) with chemical formula of Na4SiCO6, as a depressant with pH modifier features, has been synthesized and used for phosphorus bearing iron ore in anionic reverse flotation. In this process, phosphorus is floated while the iron oxides are depressed, using fatty acids as collector in an alkaline medium. Batch and pilot flotation tests were carried out to investigate the influence of SCS on the separation selectivity of phosphorus from iron oxides. Also the depression mechanism of SCS on hematite surface was investigated by microflotation tests and FTIR spectroscopy measurements. The flotation results indicated that without using a pH modifier, the pH of pulp could reach up to 10, when SCS was used as depressant. In this condition, a better depression performance was observed compared to sodium silicate (SS). Using the SCS, phosphorus and total iron content in iron concentrate was 0.046% and 66.29%, respectively, compared to SS where these values were 0.06% and 65.75%, respectively. As a result the weight recovery of iron concentrate was increase by around 4%, compare to SS. It was also found that 238 g/t of SCS could be used instead of 3 types of reagents collectively, including 32 g/t of NaOH, 151 g/t of Na2CO3 and 400 g/t of SS gel.INTRODUCTIONThe phosphorous presents in iron ore minerals is mainly in the form apatite mineral. The apatite is separated from iron ore minerals by magnetic and flotation methods (Ma, 2012). The reverse anionic flotation of iron is one of the most common method for removing phosphorus from iron concentrate (Araujo et al., 2005).Usually apatite minerals are readily floated using anionic collectors such as fatty acids (Ma, 2012). In this conditions the iron minerals are depressed and by modifying the pH in alkaline conditions, selective separation of phosphorus from iron concentrate can be achieved. Depressants including; sodium silicate (SS), starch, dextrin, guar gum and humic acids play an important role in the reverse flotation of iron ore (Kar et al., 2013; Ma, 2012; Turrer and Peres, 2010; Araujo et al., 2005). Studies have shown that SS has a better effect on iron recovery and phosphorous reduction, compared with other depressants (Tohry and Dehghani, 2016; Qi et al., 1992, Dho and Iwasaki, 1990). In most flotation processes, modifiers are used to adjust the pH. To achieve alkaline conditions, mainly pH modifiers such as carbon dioxide and sodium carbonate are used."

Jan 1, 2018

By X. Tong

Flotation researches of indium-beard marmatite from the Dulong multi-metallic ore in the Yunnan Province of China has been studied in order to improve the grades and recoveries of zinc and indium of the zinc concentrate in the Dulong concentration plant. The experimental results have shown that copper sulfate mixed with a chemical reagent X-1 as the activator in the marmatite flotation produced a much better beneficiation than copper sulfate alone, increasing the zinc and indium recoveries of 10% and 6%, respectively, while the concentrate grades were unchanged. Also, the new activator acted well around pH 10, allowing large savings on lime consumption in the marmatite flotation. In addition, it has been found that a sufficient activated time of activator with ore slur-ry in the flotation is needed to achieve good beneficiation of the marmatite ore.

Jan 1, 2014

By N. R. Mandre

In the present work an attempt has been made to study beneficiation of low-grade iron ore associated with manganese clay by selective flocculation. The representative samples obtained for this purpose were initially subjected to chemical and mineralogical studies. The chemical analyses indicated 42.56 % Fe, 6.25% Mn, 8.72 % SiO2 and 13.60 % Al2O3 in the representative sample. The mineralogical analyses carried out under optical microscope showed that the ore mineral predominantly comprised of iron oxides and hydroxides along with manganese and silicate as gangue minerals. Prior to the beneficiation studies, the samples were subjected to settling tests. These tests were carried out by using poly acrylic acid (paa), poly acryl amide (pam) and guar gum (gg) as the flocculants. From these studies it was observed that settling rate of particles increases with increase in pH and a maximum settling rate of 0.23 cm/secwas obtained at pH 9 in presence of poly acryl amide. Further, the samples were subjected to selective flocculation studies using above-mentioned reagents. From these studies better results were obtained in the experiments carried out with poly acryl amide as the flocculant. Thus, it was possible to upgrade the iron content from 42.5 to 53%, with a recovery of 92%. However, 51% Fe with recovery of 87% and 50% Fe with recovery of 94 % was obtained using poly acrylic acid and guar gum respectively.

Sep 1, 2012

By X. Tan

In this paper, the application of the metallurgical residue produced from oceanic cobalt-rich crusts leaching process is studied in the area of wastewater treatment. The modification of the sulfuric acid leached residue and the adsorption behavior behaviour of heavy metal ions such as (Pb(II), Cu(II), Zn(II), Cd(II), Ni(II), Mn(II), Co(II), As(III)) in the waste-water on the surface of the residue are investigated. The rate of adsorption efficiency of heavy metal ions in the wastewater on the modified sulfuric acid leached residue is investigated. It is showed that when the initial pH value of wastewater containing heavy metal ions including Pb(II), Cu(II), Co(II), Ni(II), Mn(II), Zn(II), Cd(II) and As(III) range from 1.4 to 1.9, with the use of the modified leached residue, As(III) ion can be selectively separated from wastewater, and the separation efficiency can reach 90%~100%. The adsorption efficiency is affected by pH, reaction time, amount of adsorbent and the initial concentration of heavy metal ions in wastewater. The results showed that the optimum pH was natural pH value and the absorption equilibrium time was 2 hours. The adsorption efficiency of heavy metal ions increases with the addition of adsorbent and de- creases with the increase of the initial concentrations of heavy metal ions in wastewater. The heavy metal ions polluted wastewater, of which concentration is about 22.5~793.6 mg/L, can meet the national disposal standard for industrial wastewaters in China very well, after 2 hours treated by 3.75~7.50 g/L modified adsorbent at room temperature. Standard methods were used to investaged investigate the absorption behaviour between abosorbentabsorbent and metal ions in the wastewater. It is concluded that the adsorption behaviour of Co(II), Mn(II) and As(III) ion fits Freundlich iso-therm, while the adsorption behaviour of Pb(II), Cu(II), Zn(II), Cd(II) and Ni(II) ions fits Langmuir isotherm, and the saturated adsorption amount of the modified adsorbent is about 54.0~336.7mg metal/g?adsorbentabsorbent. It is suggested that the modified sulfuric acid leached residue produced from oceanic cobalt-rich crusts is environmentally attractive and meaningful as absorbent used in heavy metal ions polluted wastewater.

Jan 1, 2014

By X. Q. Cui

The surficial collapse pits of the mining areas are continuously spreading in Beiminghe iron mine of Hebei Province as a result of the exploitation, which has done a lot of damage to the circumjacent ecological environment of the mine. It is a new path to handle the mine tailings at the present time that the thickened tailings are used to backfill the collapse pits, in which way the problems of the surficial collapse and the nowhere discharge of the tailings are both solved, and the rational utilization of the mine tailings and the clean production of the mine enterprise are realized. The?200 mesh content of the tailings of Beiminghe iron mine is 76.87%, and?325 mesh content is 63.50%. Semi-industrial dehydration experiments to the tailings have been done using the hydro-cyclone classified thickening process. The results show that when FX125 hydrocyclone is adopted, the concentration of underflow can achieve 73.31%, and the yield ratio is 61.78%; when FX150 hydrocyclone is adopted, the concentration of underflow is 70.57%, and the yield ratio is 54.31%, which indicate that the minor diameter hydrocyclone is propitious to improve the yield ratio of the underflow. The semi-industrial dehydration system of hydrocyclone classified thickening process is com-posed of FX125 hydrocyclone, D3600 high-efficient thickener and D2000 thickening bucket. The capacity to tailings disposal can achieve 3.07t/h and the tailings can be thickened from 28.86% to 62.45%. Compared with the traditional thickening-filtrating process and the total tailings directly filtrating process, the hydrocyclone classified thickening process has the merits of simple flow and high thickening efficiency, so it is easy to be realized in the mine.

Jan 1, 2014

By G. J. Mei

Separation of hematite and iron-containing silicate gangue - aegirine by Flotation using fatty acid collector and (NH4)2SiF6 depressant was investigated in this paper. The flotation tests with single minerals show Ammonium hexafluorosilicate (AHFS) is a selective depressant for aegirine in hematite flotation in weakly acidic medium, which can strongly reduce the flotation recovery of aegirine in the pH range of 4~5 while has no evident effect on the flotation of hematite. The results of flotation tests with actual ores show the separation of hematite from aegirine gangue by forth flotation was possible, and good selectivity was achieved. Chemisorptions were observed on aegirine mineral surface treated by AHFS solution by using SEM and IR spectrum measurements. On the basis of coordination chemistry, a comprehensive mechanism, which reasonably explains the selective depression of AHFS for iron-containing silicate, is also proposed.

Jan 1, 2014

By Herman D. Mendoza

"The removal of finely dispersed mercury from the effluent of small-scale gold mining operations and the eventual recycling of the recovered mercury was investigated. The conventional method of amalgamation, or “ball milling”, involves the mixing of a substantial amount of mercury with ore to recover free gold during the grinding stage. During grinding, a portion of the mercury is transformed into minute droplets, which readily float and disperse in the effluent. A kilo of gold ore containing 20-30 g Au/t, when ground with 3 g of mercury, yields 0.005 to 0.050 mg/l of mercury in the effluent, exceeding the concentration of 0.002 mg/l allowed by law for surface water.A rapid recovery/ recycling process was introduced using the surface-chemical properties of finely dispersed mercury in aqueous suspension. It was observed that fine droplets of mercury in an aqueous environment readily attached onto a collector having positive surface potentials for pH 3 to 10. An example of a suitable collector is alumina (Al2O3), which has an iso-electric point of pH 9.8. When an effluent, coming from a settling procedure after an amalgamation-grinding stage, was passed through a column-bed packed with alumina fiber, removal and subsequent collection of finely dispersed mercury occurred under acidic conditions, pH 3-6. Recovery was around 99.5%, yielding an acceptable quality of water safe for disposal into the aquatic environment. The collected droplets of mercury, which are left in the column-bed of alumina fiber, were subsequently treated with a cleansing solution at alkaline conditions pH 10 -12. At these conditions, both mercury and alumina fiber carry negative surface potentials, which results in the eventual detachment of attached mercury from the alumina collector. The effective recovery of attached mercury from the alumina collector by the detachment procedure is only 30-35%. The collected mercury from the cleansing stage can be collected, cleaned and can be used for subsequent amalgamation processes. The attachment and detachment behaviour of finely dispersed mercury onto and from the microscopic collector alumina is shown in Figure 1. The result of attachment and detachment were analyzed using existing interaction energy theories like the Darguin-Landau- Verwey- Overbeek (DLVO) Theory."

Jan 1, 2003

By U. A. Peuker, T. Mütze, F. Shayo, A. Wikedzi

"Buzwagi Gold Mine (BGM) is operated by Acacia Mining and located in the Lake Victoria Goldfields of central Tanzania. The mine treats a sulphide copper-gold ore to produce gold in form of doré bars and a concentrate containing gold, copper and silver. The BGM comminution circuit includes a primary crushing stage with a gyratory crusher and two grinding circuits using a Semi-Autogenous Grinding (SAG) mill and a ball mill. Currently, the grinding circuits are inefficient in achieving the aspired product fineness of xP,80 = 125 µm even at low to normal throughputs (450-600 t/h). An optimization study of the circuit performance was conducted to improve product fineness through circuit surveys, experimental lab work and simulations. The analysis of survey data revealed inefficient grinding operation characterized by ball mill product (xP,80 > 400 µm) coarser than designed and higher hydrocyclone cut sizes (xT > 200 µm), too. In MODSIM several options were simulated such as reduction of vortex finder and apex diameters, reduction of SAG screen aperture, and modification of cyclone feed density. An improvement of product fineness could be achieved by simulation of the cumulative changes but it was also concluded that the current equipment cannot achieve the target. Further simulations showed that a second ball mill might help to achieve the desired fineness as well as an increase of grinding throughput by up to 26 %. INTRODUCTION Simulation of comminution, particularly grinding and classification, has received great attention for many years, due to the fact that the unit operation is by far the most important both in terms of energy consumption as well as overall plant performance (King, 2001; Wills et al., 2006). The use of simulation is advantageous due to low cost as well as its ability to consider many variables simultaneously. Moreover, there are some benefits for optimizing grinding operations by simulation like decreased unit operating costs, increased capacity, and improved performance of downstream processes due to improved feed size distribution. This paper presents an optimization study for the grinding efficiency at Buzwagi Gold Mine. This was achieved through three full scale circuit surveys, lab work and simulations with MODSIM."

Jan 1, 2018

By T. Sun, J. Kou, C. Sun, D. Gong

"HPGRs (high pressure grinding rollers) have demonstrated higher energy efficiency than conventional comminution equipment due to their application of particle-bed breakage mechanism. In the past decades, a large number of studies have been conducted to understand the characteristics of particle-bed breakage, however, few of them concerned the difference between single-particle and particle-bed breakage in compression comminution. This comparison is able to demonstrate possible advantages of particle-bed against single-particle in a breakage event due to the isolation of the effect of force loading mechanism. The present study investigated progeny fragment size and energy efficiency of these two breakage methods at variable energy inputs and particle sizes. The results show that at low energy inputs of around 0.25 kWh/t, the numbers of fines resulting from the two breakage methods are very similar. However, at higher energy inputs of around 1.0 or 2.0 kWh/t, particle-bed breakage generated significantly more fines. From the perspective of newly generated surface energy, particle-bed breakage is obviously more energy efficient than single-particle breakage at energy inputs between 1.0 and 2.0 kWh/t. It was also found that HPGRs demonstrate high energy utilization rate in generating new surface energy within their typical range of specific energy input. INTRODUCTION Rock comminution is an important and energy-intensive process in mining operations. Nowadays, decreasing head grade and mineral liberation size of ore deposits greatly increase the energy cost of comminution process. HPGRs have been successfully applied in the mining industry to improve the energy performance of comminution circuits. Typically, 10 to 30% of energy saving can be achieved due to the application of HPGR (Wang et al., 2013; Casteel, 2005; Günter et al., 1996)."

Jan 1, 2018

By S. N. Titkov

Flotation of halite?new, recently begun to be developed method of processing water-soluble minerals. In the paper the results of studies of the new directions of the application of halite flotation are represented. Studies on the flotation separation of halite and sylvite in the salt solutions of different composition, showed the possibility of the halite flotation realization not only in the carnallite brine (content of salts: MgCl2 25.5% ~ 26.0 %, chlorides of potassium and sodium 5.5% ~ 6.0%), but in the sylvinite brine, which contains 10.5% ~ 10.8% KCl and 20.5% ~ 21% NaCl. On the basis obtained results are carried out experiences on the halite flotation from samples of the converted products of sylvinite ores by flotation method and by the method of the electrostatic separation. It is shown that in all cases it is possible to accomplish effective and selective recovery of halite. The halite recovery into the foam product of flotation makes 80% ~ 90% with an increase in the content of KCl in the cell product of flotation in comparison with a composition of flotation feed: - from the product after the electrostatic separation of ore ? from 76% KCl to 95% ~ 97% KCl; - from the small fraction of ore (- 0.25 + 0 mm) ? from 30% KCl to 60% ~ 64% KCl; - from middlings (intermediate products) of sylvite flotation from 16% ~ 31% KCl to 45% ~ 80% KCl. An increase in the molecular weight of alkylmorpholines increases the selectivity of the flotation separation of halite and sylvite. Halite recovery from small size fraction of ore or/and middling of sylvite flotation with the sub-sequent joint filtration halite product and tails of sylvite flotation allows to reduce considerably expenses for manufacture potash chloride from Pilot-industrial tests for the first time confirmed the possibility of the effective halite flotation in the saturated solutions of chlorides of potassium and sodium (sylvinite brine).

Jan 1, 2014