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By N. D. Dementieva

Based on lab and pilot-plant tests to recover gold, silver and lead of a Russian deposit, a collective and flotation concentration flow sheet with lead and sulfide concentrates was proposed. The type of the deposit is gold-sulfide-quartz-bearing. The recovery of gold, silver, and lead to a lead concentrate was 47.4%, 68.1% and 70.7%, respectively. The grade of gold and silver grade was 170 g/t and 372 g/t, respectively. The mass percent of lead, zinc and arsenic mass percent was 41.7%, 0.34% and 0.63%, respectively. The recovery of gold and silver recovery to a sulfide concentrate was 50.1% and 28.8% at the grade of gold 20.5 g/t and silver 66.2 g/t. Sulfide concentrate with arsenic mass percent ~3% was biooxidised followed by CIL and obtaining metal at the site as well as zinc concentrate from bioleaching solutions with zinc mass percent 45% at its 64% recovery. The total gold and silver recovery using a developed technology with the account of their treatment at a polymetallic plant was 92.2% and 93.7%, respectively.

Sep 1, 2012

By Hongchun Lou, Yan Zhang, Xianjun Lu

"The potential of Fe(NO3)3 modified attapulgite was assessed for the adsorption of phosphorous removal from aqueous solutions. Natural attapulgite was modified by Fe(NO3)3, and thus the composite adsorption materials could be obtained. Also phosphorous adsorption characterization on the iron-coated attapulgite was investigated in this research. Taken overall the results suggested that phosphorous adsorption capacity could be improved markedly by adopting ferric modified-attapulgite. The optimum phosphorous adsorption were developed in the experimental conditions of the solid-to-liquid Ratio 1g/L, the initial phosphorous concentration 5 mol•L-1, reaction time 120 min, temperature 25 ? and pH=4.0, the removal efficiency is 92.28% for the modified attapulgite. In comparison of batch experimental results of phosphorous adsorption onto natural and iron-coated attapulgite, it has provided much more likely the same change trend. On the other hand, the outcomes of this analysis indicated that ferric modified adsorbent was notably influenced by adsorption element over natural attapulgite. The Freundlich isotherm adequately described the adsorption course of phosphorous onto ferric modified-attapulgite with the correlation coefficient R2 goodness of fit of 0.9690."

Jan 1, 2014

By Ljudmil S. Grigorov, Eric Forssberg, Hanumantha Rao Kota, Robert J. Pugh, Lidia B. Alexandrova

"The effects of mixed collectors on some physicochemical parameters (thin film lifetime, kinetics of expansion of the three-phase contact, equilibrium three-phase contact angle and surface tension) involved in the elementary act of flotation were investigated experimentally.We have used the microscopic thin balance developed by Scheludko et al (1970) to study the stability and interactions within the thin film between the silica surface and air bubble with (i) a long chain C10 amine and (ii) a cationic/anionic surfactant systems consisting of mixtures of a long straight C10 chain amine with long straight chain C8, C10 and C12 sulfonates at 1:1 molar ratio. The experiments were performed at a range of concentrations at pH 6.5. The equilibrium contact angle was measured with a microscope technique using the modified Mingins-Sheludko method (Mingins and Scheludko, 1979). The surface tension was measured by the method of Wilhelmy (Adamson, 1960). It was found that the mixed surfactants caused lower contact angles, lower rates of the thin film rupture and longer film lifetime, compared to the pure amine system."

Jan 1, 2003

By James A. Finch

"Developed countries are in the midst of re-structuring education in minerals-related disciplines as raw materials, once vital to wealth generation, contribute progressively less to the economy. The changing circumstances mean institutions must justify mineral processing by new criteria. After considering what a mineral processor needs to know the ways to deliver are considered. Taking twenty graduates per year to illustrate, one institute could justify this number as a program in a department or school (of whatever grouping of disciplines). A strong research component would make the package additionally attractive. Divide among four institutions and mineral processing is best served in a research centre with delivery taking advantage of multi-centre and distance learning opportunities.INTRODUCTIONThis is one of five plenary lectures at the Congress. I may have preferred a technical subject, as that is what occupies most of my thinking hours, but it is reasonable as an academic that this topic should come to one of us. The fact the topic has not been high on my agenda may provide one advantage: that I can suggest directions free of pre-conceived notions, just as when reviewing the background to a technical article the temptation to select in support of one?s interpretations is resisted. The trick will be to add something new.One reading of the title seems to query whether mineral processing education has a future. (I will refrain at this juncture from defining mineral processing and assume for now we share an understanding of the term.) The proximity of natural and human resources led to the Industrial Revolution and the interaction continues to stimulate and sustain industrial and economic development in many parts of the world. The search for raw materials has historically played a key role in opening frontier areas, with undersea and outer space next. The reason is obvious: raw materials along with food and energy underpin society. Thus mining will remain part of human activity along with the need to pass on the knowledge and skills.What happens over time is resource relocation. Raw materials become exhausted in one area to be replaced by sources elsewhere; the demand for some materials lessens while growing for others. Educational institutions can find themselves with the wrong geography or with staff focussing on the wrong subjects. When this happens the restructuring produces much discussion. This is where we are in parts of the world. Japan, Western Europe and the United States are the prominent examples (Dowd, 1998; Flemings and Suresh, 2001; McDivitt, 2002). Mineral processing is becoming a series of courses in related disciplines: resource engineering, environment engineering or materials engineering. The start of the transition is evident in Canada and Australia, perhaps earlier than expected given the contribution of raw materials to their economies. A common backdrop in the developed countries is social pressure to reduce the impact of the ?untidy? aspects of life, and that includes mining. (This is not new: in classical times the mine face was the preserve of slaves and convicts.) These values are exportable, potentially affecting recruitment into mining schools everywhere."

Jan 1, 2003

By John A. Rumball, Martin R. Houchin

"For gold leaching of oxide ores, the optimum cyanide : oxygen ratio is approximately 106 mg/L NaCN : 10 mg/L O2 (6.9:1 by mole). In sulphide ores, galvanic coupling of the gold with the electrically conductive sulphides significantly increases this ratio such that leaching may occur in a relatively low dissolved oxygen (DO) environment.RESULTS AND DISCUSSIONGold leaching in a carbon in pulp (CIP) and carbon in leach (CIL) circuit is driven by the supply of cyanide and oxygen, as described by Elsner’s equation (1):• 4Au + 8CN- + O2 + 2H2O ? 4Au(CN)2 - + 4OH- (1)Equation 1 suggests that the ideal molar cyanide : oxygen ratio should be 8:1, or 122.5 mg/L NaCN to 10 mg/L of DO. It is generally accepted however, that oxygen is not completely reduced to hydroxide at the gold surface, and that equation (2) better represents the gold leaching reaction (Marsden and House, 1992):• 4Au + 8CN- + 2O2 + H2O ? 4Au(CN)2 - + 2OH- + H2O2 (2)Additionally, since gold leaching is limited by either the diffusion of cyanide or oxygen to the gold surface (Habashi, 1993), then the relative diffusion rates of cyanide and oxygen must also be taken into account. Marsden and House (1992) suggest that the ideal molar cyanide : oxygen ratio is 6:1, while a more recent investigation using laboratory and plant observations (Heath and Rumball, 1998) suggested a molar ratio of 6.9:1.In practice, although a cyanide : oxygen mole ratio of 6.9:1 applies to some circuits, there are many instances when this is not the case.In sulphide ores, galvanic coupling of the gold with the electrically conductive sulphides significantly increases this ratio. Galvanically connected sulphides, and possibly carbon, provide a large surface area for oxygen reduction. As the cathode surface area increases, the oxygen concentration required to sustain the same rate of gold oxidation decreases proportionately.Galvanically connected particles may be composite particles, however they may simply be liberated sulphides in the same suspension as the gold. Particle collisions within a slurry are sufficient to provide momentary galvanic connection (Greet, 1995)."

Jan 1, 2003

By Michel Brissette, Shaheer Sanuri, Veroljub Mihajlovic

"The main purpose of geometallurgy is to improve the mine planning by mapping hardness and throughput into the mine block model using geostatistics to eliminate the large production fluctuations caused by the ore hardness variability coming from different parts of the pit. However, not all grinding testwork available commercially can be applied to geometallurgy because of large amount of sample required and the associated cost per test. Geometallurgy is now being used in designing grinding mill but as the mining project progress, the accuracy of engineering must improve from 30-35% during the scoping study to 10-15% at the bankable feasibility stage and less than 5% for Engineering Procurement and Construction Management. For SAG milling, the SAG Variability Test (SVT) is an abbreviated version of the standard SAGDesign test developed for this geometallurgical need. The same amount of ore than the original test is ground for three cycles into an open batch SAG test and then the SAG pinion energy is predicted to the standard end point of 80% passing 1.7 mm. For ball milling, a Bond Variability Test (BVT) was also developed using the SAG ground ore from the SVT. The Bond ball mill grindability (Gpb), corresponding to the average of the net grams per revolution from the last three cycles of the Bond test, is also predicted from the third cycle. The relative error for both the predicted SAG ore hardness (WSVT) and Bond Work Index on SAG Variability ground ore (Sv-BVI), in kWh/t, averaged 5.5% and 3.8% respectively compared to the SAGDesign testwork values and are within 7% and 10% respectively from plant benchmark and within 7% from the samples sent to all partner laboratories. Both tests can be continued to measure the true hardness necessary for mill sizing. This article discusses the development of these two new tests, SVT and BVT, their potential use and accuracies in any geometallurgical program."

Jan 1, 2014

By P. Saha, D. M. Chakrabarti, National Metallurgical Laboratory, P. Bhattacharyya, B. R. Sarkhel

The estimated reserve of the phosphatic soils at the Sukulu hills region of Uganda upto a depth of 60 meters is of the order of 200 million tonnes and analyses 10-12% P2O5. A 5-tonne sample of the phosphatic soil was received at the National Metallurgical Laboratory (NML), India for beneficiation studies. The ore occurred as a naturally degraded soil deposit and the slime contained earthy lateritic minerals which made it extremely sticky. Microscopic examination revealed the presence of limonite, magnetite (martitised), hematite and goethite as major minerals along with apatite and quartz. Presence of clay was indicated by earthy smell on moistening the sample. It was observed that getting rid of the slime was as important for adequate beneficiation of the ore, as it was difficult. Earlier studies on samples of the same ore at the NML and abroad established that it was necessary to pass the ore through a series of three stages of desliming, each consisting of a step of rigorous attrition scrubbing followed by hydrocycloning, for effective elimination of the sticky slimy coating which stubbornly adhered to mineral grains. Studies were carried out at the NML in two phases to develop a flowsheet for beneficiation of the phosphatic soil. In the first phase, reproducibility of the results obtained by employing a flowsheet developed abroad was confirmed along with suggesting some modifications of the flowsheet. Then 200 kg. of concentrate was produced following the recommended flowsheet for evaluation of its suitability for production of fertilisers. To remove the limonitic slimy coating, which stubbornly adheres to the other mineral grains, it was necessary to subject the ROM ore, after screening out the +840 micron size fraction which contained only 1.75% P2O5 , to three stages of hydrocycloning each preceded by attrition scrubbing. Then only conventional beneficiation techniques could be effectively employed to upgrade the ore.

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 Z. B. Pan

The innovative point of this new technology can be described that under normal pressure and 70°C, using a new combined activating and oxidizing agent, the minerals bearing arsenic and organic carbon are oxidized effectively. This technology can be applied in commercial scale. The high carbon and arsenic bearing gold ore is taken from Gansu, where hundreds of tons gold deposit. The ore has 0.51% of arsenic and 1.75% of carbon (1.07% of organic carbon). The gold occurs mainly in native with particle size 0.02~0.6 mm and a few are enwrapped by pyrite and arsenopyrite. The traditional cyanide leaching process gains a very low gold leaching rate with only 5%. Through systemic experimental study, the technological flowsheet of flotation process for concentrating gold, then pre-treating the gold concentrate followed by cyanide leaching the gold concentrate are used. With 4.92 g/t gold of head sample, the gold concentrate with 46.52g/t gold grade and 96.53% gold recovery is gained by flotation process. Followed by pre-treatment for 6 hours, the gold concentrate is leached with cyanide after filtration and washing. The total gold leaching rate is up to 95.44%.

Jan 1, 2014

By Mingfei He, Yude Gao

"The leaching kinetics was studied during leaching of indium-bearing zinc oxide to recover indium. The effects of temperature, sulfuric acidity, particle size, stirring rate on dissolution rate of indium were investigated. Results showed that the leaching process was controlled by chemical reaction at the condition of low sulfuric acid concentration. While the acidity increased to 1.8 mol/L, the leaching process was transformed into an internal diffusion controlling model . The corresponding value of reaction orders under low acidity and high acidity were 1.194, 0.156, respectively. Further calculation results indicated that the leaching reaction activation energy was only 10.97kJ/mol under the condition of high acid. The leaching rate constant was inversely proportional to the square of particle size, while the effect of stirring rate on leaching process was insignificant. The leaching process can be enhanced by increasing the temperature and acid concentration, as well as the decrease of particle size. And a final kinetics eqution was proposed. KEYWORDS Indium, extration, leaching kinetics, zinc oxide INTRODUCTION Indium[1,2] plays an important role in many aspects of living and industrial production owning to its special physical and chemical properties. The reserve of indium is rare on the earth, and most of indium is distributed in some other minerals[3]. Some researchers [4,5] have proposed that indium mainly substitutes for the lattice of other minerals by isomorphous. Therefore, the extraction of indium should take into account of the production technologies of those minerals to comprehensively recover the valuable metals."

Jan 1, 2018

By W. H. Cheung

Nickel and cobalt production relies heavily on the processing of nickel laterite ores; these ores contain the majority of known Ni and Co reserves, 80% and 90% ~ 95%, respectively. The Caron process is used to extract Ni and Co metals from low-grade nickel laterite ores (1% Ni and 0.1% Co). It involves selective reduction of NiO and CoO minerals followed by ammoniacal leaching. Despite low metal recoveries, this process continues to be technically and economically reliable and thus commercially important globally. However, keeping the Caron process cost-effective has been a challenge for industries because of increasing economic and environmental pressures. Although studies of the Caron process have been extensive, significant improvements in recoveries of industrial processes have not been forthcoming. Reported industrial metal recovery from high iron content ore is only 80% ~ 85% for Ni and only 40% ~ 45% for Co; recovery is even lower for ores rich in magnesium. The real difficulty here is that there is no complete understanding of and control over the complex series of reactions in the ore and in the leaching solution during processing. Achieving effective metal recoveries from the Caron will require control of the mineral transformations, which affect the activation of minerals by reduction. In this study, the effect of over-reduction was investigated, which superseded the effect of activation by lowering the activity of the minerals and thus metal dissolution. Reduction was carried out in a batch reduction apparatus with 1:1 and 1:2.5 CO/CO2 gas mixtures from temperatures of 500 ~ 700°C and residence times up to 60 minutes. The pre-reduced ores were leached using a standard ammoniacal leaching test. Thermodynamic analysis and possible effects of stable phase field on leaching was examined. The results suggest that control of over reduction effects are directly related to the mineral transformation and reduction conditions.

Jan 1, 2014

By R. ?. Sherembayeva, ?. ?. ?uktybayeva, ?. B. Rakhimbekova, N. ?. ?m?rova

"Studying the interaction of the phosphorus-containing flotation reagent with the components of the flotation system ""liquid-gas-solid"" during flotation of non-ferrous metal ores allows regulating the enrichment process. The phosphorus-containing reagent-collector anabasinium-O,O-dimethyl-thiophosphate has beensynthesized under laboratory conditions and the concentrated aqueous solution of anabasin hydrochloride and the 40% solution of sodium hydroxide have been used as the raw material. The phosphorus-containing reagent has been obtained by interaction of the anabasine base, sulfur and dimethyl phosphite, due to sulfur and nitrogen atoms it can form complex compounds with heavy metal ions.Studying flotation of copper-sulfide ore of the Sayak field with the use of the phosphorus-containing reagent-collector anabasinium has been performed in the open and closed cycles. The degree of ore grinding was 60- 62% of the class -0.074 mm. There has been established the dependence of the ore flotation parameters on: a) pH of the pulp; b) the consumption of the reagent-collector anabasinium.The effect of the pulp pH on the technological parameters of copper ore flotation has been carried out with a reagent regulator that increases the difference in the surface properties of the minerals to be separated.There has been used lime as the regulator of the medium, the consumption of which varied from 750 to 1500 g/t; the consumption of the reagent-collector anabasinium was 10, 15, 20 and 25 g/t.It has been found that when flotation of copper ore without the use of lime at pH equal to 7.8 and at the reagent consumption, g/t: anabasinium - 25, T-92 - 55, copper recovery in the concentrate is 92.72%, the quality of concentrate is 7, 1%.In flotation according to the factory mode (the consumption of xanthate is 44 g/t, T-92 is 55 g/t, pH = 7.8 (without lime)) of copper ore, copper recovery in the concentrate is 87.38%, the concentrate quality is 7, 97%, with reducing the xanthate consumption to 25 g/t copper recovery is reduced to 83.39%, the quality of the concentrate is up to 6.69%.In the experiments using the factory reagent mode in the closed cycle, with the xanthate mass consumption of 44 g/t, T-92 blowing agent of 55 g/t, the third recleaning concentrate contained 19.13% copper with recovery of 86.89%.Using a phosphorus-containing reagent instead of xanthate with the same flow rate, the third recleaning concentrate with the mass fraction of copper of 19.82% has beenobtained with recovery of 87.77%."

Jan 1, 2018

By Igor Kuzymin, Isaak Maximov, Iuriy Akkerman, Galina Aksenova, Valentina Sentemova, Nataliya Trubechkova

"The Russian Federation possesses large reserves of oxidized ferruginous quartzites. These ores are currently mined together with magnetite ores and are stockpiled on account of unavailability of an efficient processing technology. This paper presents the results of investigations performed on the ores from the largest oxidized ferruginous quartzite deposit located in the Kursk Region iron-ore basin of Russia. The objective is to produce concentrate with iron content at least 66% and silica content within 5%. A processing technology was, developed under laboratory conditions, which included ore preparation, wet magnetic separation in weak and strong fields and reverse cationic flotation with addition of caustic soda, dextrin and Lilaflot 811M (alkoxy propylamine) collector from Akzo Nobel. The developed technology produced iron recovery concentrate with iron content of nearly 67%, and that of silica - 4.1%, iron recovery into concentrate being 72.6%."

Jan 1, 2014

By Kazutoshi Haga, Batnasan Altansukh, Shirchinnamjil Nyamdelger, Atsushi Shibayama, Narankhuu Ariunbolor, Gunchin Burmaa

In this study, the authors investigated the capability of an acidic thiourea solution and silicon-organic polymers (PSTM-3c, PSTM-3T) as ion-exchange resins to recover of gold from gold-bearing ore concentrates. Firstly, elemental gold present in flotation concentrates was dissolved by leaching with a solution of nitric acid and thiourea under the optimized conditions. The best result on gold thiourea leaching was obtained under a nitric acid concentration of 2 N, thiourea concentration of 2 %, and oxidant (Fe2(SO4)3) concentration of 3 %, solid:liquid ratio (S:L) of 1:20, leaching time of 6 hours at 20 °C with a gold recovery of 93.7 % for a gold flotation (unroasted) concentrate. For the roasted gold concentrate, the dissolution of gold was 94.8% under the optimized leaching condition as nitric acid concentration of 0.5 N, thiourea concentration of 2 %, and the oxidant concentration of 1 %, S:L ratio of 1:10, leaching time of 4 hours at 50 °C. Secondly, a series of sorption tests was carried out on a pregnant leach solution containing gold complex Au[S=C(NH2)2]2+ using silicon-organic polymers (PSTM-3c and PSTM-3T), respectively. The effects of solution pH, adsorption time and temperature in the sorption process are also studied. As the results, the PSTM-3c and PSTM-3T polymers were adsorbed about 94.4 % and 95.1 % gold from the pregnant leach solutions under the optimum sorption conditions, respectively. It was found that the both PSTM-3c and PSTM-3T polymers were effective for the recovery of gold from the pregnant leach solution of thiourea. Adsorption activity of the PSTM-3c and PSTM-3T polymers was confirmed by new peaks near 1700 cm-1 in the FT-IR spectrum of the gold-adsorbed polymers, respectively.

Jan 1, 2014

By I. A. Podkamennyi, O. E. Kovalchuk, G. P. Dvoychenkova

"Using the methods of electron microscopy and IR spectroscopy, mineral components in the run-of-mine kimberlites and on the surface of diamonds were identified. The mechanism and sequence of mineral formation on the surface of diamond crystals were established. Systematization and classification thereof were undertaken for the first time ever.INTRODUCTION The high content of diamonds sized -5 mm in the kimberlite ores of Western Yakutia determines their processing depth in the size class 0.5 mm and finer [1, 2]. These diamonds are recovered by grease and froth separation based on the natural hydrophobic properties of the crystals. When refractory altered kimberlite ores are fed into the industrial process, the performance of grease and froth separation of the diamond ore greatly deteriorates. [4, 10, 15, 17]. Altered kimberlites are known to contain up to 90% of secondary minerals. This reduces by up to 40% the recovery of diamonds by grease and froth separation. [6].Diamond crystals from the grease and froth separation tailings were investigated. On the surface of the examined diamonds, hydrophilic mineral formations were identified, responsible for reduced recovery by physicochemical separation methods [3, 12, 18]. The aim of this study was to identify and describe the relationship between the composition of kimberlite ores and mineral formations on the surface of diamond crystals in order to systematize, classify, and then develop methods of removal in grease and froth separation processes. MATERIALS Following materials were examined in this study: - kimberlites of various material composition, - resulting slimes, - diamond crystals from kimberlite ore processing products."

Jan 1, 2018

By V. F. Bannikov, A. P. Kozlov

"The purpose of this study was to develop a theoretical and technology framework for designing new combined technologies for processing tailings containing non-ferrous and noble metals using dry separation. The main goals of the study were: identifying the distribution patterns of valuable components in various fractions in dry separation processes; examining the patterns identified for different types of non-ferrous and noble metal tailings; developing a theoretical and technology foundation for the development of innovative technologies for the efficient processing of tailings with minimal process water consumption. To examine the possibility of efficient recovery of valuable components from refractory tailings using dry separation, an experimental laboratory unit was designed and built, composed of novel or modified lab equipment: plasma unit for increasing the magnetic susceptibility of Fe - containing minerals; magnetic separator supporting isodynamic and polygradient fields; pneumatic separator for the dry separation of finely dispersed mineral mixtures; dielectric separator using the resonant frequency of the electric field. The laboratory unit supports various modifications and combinations of the lab equipment. The experimental studies led to the development of a unified process scheme for lab tests of mature and non-ferrous smelter tailings using dry separation methods.INTRODUCTION The huge amount of waste from mining and processing of mineral resources Russia inherited from the extensive use of mineral resources in the Soviet Union and amount of waste continue to grow. Often the content of valuable non - ferrous, rare and noble metals in waste exceeds the balance contents in natural ores of the developed deposits. The availability of a rich scientific reserve for the creation highly effective innovative technologies for waste processing determines the necessity of the acceleration in the resolution of problems of the complex assessment, capitalization and involvement in the processing of technogenic resources in Russia (Chanturiya and Kozlov, 2014.)."

Jan 1, 2018

By Vasile Murariu, William Pate, John Herbst, Dilek Alkac

"Multiphysics models are essentially first principles models capable of providing very accurate micro scale predictions. The computational requirements for Multiphysics Models are extensive, precluding their direct use for flowsheet simulations. Population balance models get their form from theory but the model parameters have to be provided from experiments. The prediction errors for the population balance models are typically significantly higher than those for the Multiphysics Models but the computational effort for Population Balance Models is typically much less than that for the Multiphysics models. The hybrid model described here takes advantage of the strengths of each of the two primary model types and minimizing the weaknesses of each. The Multiphysics Model is used to perform high accuracy "experiments" that are difficult or impossible to conduct in laboratories or plants. The results of these experiments are combined with standard laboratory or plant data to parameterize and further enhance the Population Balance portion of the hybrid model. The creation of a useful Hybrid Model is demonstrated for a MP1000 secondary crusher. The validation of the hybrid crusher model is also presented. Finally the use of the Hybrid Model is illustrated in a full comminution flowsheet simulator."

Jan 1, 2014

By Jiongtian Liu, Jianjun Wu, Yongzhou Wan, Li Feng

"Mechanical Thermal Expression (MTE) is a developing non-evaporative lignite dewatering technology. It has been shown to effectively dewater high moisture content in low rank coals via the application of mechanical force and thermal energy at elevated temperatures. In this paper, the dewatering behavior of Yunnan Xiaolongtang lignite during the MTE-process was studied. Furthermore, how variations in temperature and pressure during the MTE process affect the surface oxygen functional groups and pore structure were also discussed using Fourier transform infrared spectrometer (FTIR) and mercury intrusion porosimetry (MIP). Increases in MTE temperature and pressure resulted in significant reductions in residual moisture content and moisture holding capacity, along with the increase of fixed carbon content and reductions of other elements content, especially oxygen content, which have been largely attributed to the destruction of the surface oxygen functional groups and porosity in lignite. The optimized technology conditions are temperature 150~220°C and pressure 7~10MPa. The residual moisture content of the lignite treated by MTE under the conditions of 200°C and 10MPa is decreased to 8%, while the dewatering rate exceeds 76 % and the calorific value approaches 22 MJ/kg. The experimental results offer a deeper understanding of the influence of process parameters on the lignite stuctures and properties. The reduction in residual moisture content, surface oxygen functional groups and pore volume under increasingly severe MTE conditions are suggestive of the structure changes that accompany increased coalification (rank) within the lignitic range."

Jan 1, 2014

Goethite ore are the third most abundant source of iron. The overwhelming part of these ore is represented by low-grade varieties which cannot be used in iron and steel industry without preliminary beneficiation. Biomass is a renewable solid fuel, which is feasible for the substitution of coke and coal in the manufacturing due to its lower contents of harmful elements and abundant raw material sources. Utilization of biomass in iron ore sintering process as heating agent and reducing agent contributes to energy conservation and emission reduction. The biomass was mixed together with goethite ore for roasting process to investigate the effects of reduction roasting. The results show that the goethite ore is dramatically reduced and magnetized into maghemite by biomass with the biomass dosage less than 20% and the roasting temperature below 650 oC. When the goethite ore roasted with biomass, the magnetization values could be enhanced more than two orders of magnitude, meets the standard of magnetic separation, which could significantly reduce the iron beneficiation cost. Keywords: biomass, goethite, reduction, magnetization, roasting

Sep 1, 2012

By A. N. Zagorodnyaya, T. N. Bukurov, Z. S. Abisheva, A. S. Sharipova

"This work presents the results of the development of a flowchart for the processing of copper production dusts resulting in the production of of lead sulphate, tribasic lead sulfate (TBLS), ammonium perrhenate, zinc, copper-cadmium cake.Dust having the following composition (wt. %): Pb 39.3, Zn 6.5, Cu 5.5, Stotal 10.8, As 1.05, SiO2 0.6, Cd 0.8, Bi 0.04, Ag 0.02, Re 0.03 was used in the work. Chemical and X-ray analyses indicated that lead, zinc and copper are generally present in dusts in the form of sulfates, various amounts of oxides, sulfides, silicates; sulfur (according to chemical analysis data) is present in the form of sulfates (~ 90%) and in the form of sulfides (~ 10%).Studies have been made on two-stage stripping of copper production dusts containing lead, zinc, copper, cadmium, rhenium in the form of various compounds, at first with soda solution, then with nitric acid solution. It has been indicated that the combination of the operations allows the recovery of 98% rhenium, 82% lead, 92% copper, and 96% zinc and cadmium each in the solution. Furthermore, at the stage of dust leaching with soda, 80% of rhenium and 0.01% of nonferrous metals each are recovered in the solution. Remaining quantities of all metals are recovered in nitric acid solution up to maximum values. Such behaviour of metals is caused by metal phase content in dusts and properties being formed during leaching of compounds. At the stage of dust leaching with soda solutions, carbonates with various compositions are formed due to the exchange reaction between nonferrous metal sulfates and sodium carbonate. The conversion of PbSO4 into PbCO3 is apparently occurs in the solid phase and the conversion of readily watersoluble zinc, copper, and cadmium sulfates into that of carbonate salts occurs in the liquid phase. Rhenium passes into solution due to the presence of rhenium greater quantity in dust in the form of readily soluble perrhenate. Some portion of the perrhenate is, probable, blocked by lead sulfate and liberated in the conversion of PbSO4 into PbCO3."

Jan 1, 2003