Search Documents

By A. Mainza, H. Hill, L. Bbosa

"The main objective of ore breakage characterization is to quantify the relationship between the energy supplied for breakage and the size of the resultant progeny. In this study, the ore breakage characteristics of drill core particles and crushed particles were compared. Single impact breakage tests were conducted on a basalt ore sample using the JK rotary breakage tester. The size-dependent model by Shi and Kojovic was used to fit the breakage data because the effect of initial particle size is accounted for in the model. It was found that crushed particles consistently gave a higher degree of brokenness than drill core particles and the differences were attributed to the internal structure of the particle. Drill core particles contain a lower crack density (paths of weakness) than crushed particles which have been pre-weakened during the crushing process. An increase in crack density is directly related to a decrease in the minimum threshold energy. Upon application of a stressing force above the stress threshold, particles with a higher crack density will fracture more easily because of a higher rate of crack propagation due the numerous paths of weakness available. It was therefore concluded that under the same conditions, drill core particles are more competent than crushed particles. INTRODUCTION Comminution devices have commonly been reported as being the largest energy consumers in mineral processing circuits (Napier-Munn, 2015). Cleary (1998) reported that the inefficiency of comminution devices is a result of less than 5% of the supplied energy being utilized for breakage, with the majority being expended as heat and noise. With steadily declining fossil fuel supplies and a looming global energy crisis, the mineral processing industry has increased its focus on energy reduction measures. Ore breakage characterization plays a critical role in the design and optimization of comminution circuits (Tavares and King, 1998). The main objective of ore breakage characterization is to empirically quantify the relationship between the energy supplied for breakage (specific input energy) and the size of the resultant progeny through laboratory particle breakage tests (Napier-Munn, 1996). Commonly used laboratory breakage devices include the Drop Weight Tester (DWT), Split Hopkinson Pressure Bar (SHPB) and more recently, the JK Rotary Breakage Tester (JKRBT). Typically, a standard single impact breakage test procedure consists of a matrix of experiments on particles in four to five narrow size fractions each tested at low, medium and high energy levels. The resultant progeny of each test is sized to obtain a particle size distribution (PSD) from which the degree of breakage, usually in the form of a t10 value (percentage of the progeny material that passes 1/10 of the feed size) can be obtained for each size fraction and energy level (Napier-Munn, 1996)."

Jan 1, 2018

By A. N. Lazaridi, S. I. Afanasenko, S. A. Safonov

"This scientific work presents the results of comparative tests of centrifugal concentrators with a vertical and horizontal axis of rotor spinning. The article also describes a qualitative analysis of the behavior of mineral particles in a centrifugal field with fluidization for concentrators with a horizontal axis. The analysis was aimed at studying the effect of the quantity of fluidization holes in the basket on the degree of magnetite recovery and the effect of the basket rotation speed. The experiment demonstrated that concentrators with a horizontal axis are not inferior in terms of the qualitative recovery indicators to concentrators with a vertical axis. The researches also provided facts about the wide practical application of concentrators with a horizontal rotor arrangement. INTRODUCTION Methods of gravity concentration, attractive for their ecological purity, see their revival. First of all, this is due to the invention of centrifugal concentrators with the fluidization of the layer of mineral particles. The creation of a new generation of gravitational centrifugal concentrators made it possible to achieve a tangible breakthrough toward the direction of reducing the grain size of mineral particles effectively divided by density, up to 10 µm at a high production capacity of the vessels reaching 500-600 tons per hour. Such separators enable dressing a crushed ore of flotation grade and in some cases replace the flotation process, especially in primary operations, in joint cycles. In the long term, this method of concentration should significantly expand the scope of application, replace partial flotation processing, ensure the introduction of environmentally friendly energy-saving technologies at ore-processing plants. Summarizing the international experience for the last 30 years, it can be confidently asserted that centrifugal methods of concentration have become widely used in processing schemes for the separation of finely dispersed products in the mining industry. An interest in the creation of installations using centrifugal acceleration for the separation of mineral particles by density has particularly grown. Dozens of centrifugal concentrator designs have been created and continue to be improved. Centrifugal concentrators have a number of obvious advantages, which are especially relevant today. Losses of fine gold during the process of mining are well known. Gold washers screw, spiral separators, jigging machines and sluices traditionally used in the production processes do not solve the problem of fine gold recovery. Table 1 shows the results of comparative tests of a centrifugal concentrator (ITOMAK) and other concentration equipment. According to the research conducted by the company GMRS, Krugersdorp, South Africa (2000), the initial material was gold ore crushed to 150 micrometers (90%)."

Jan 1, 2018

By D. Whittles, S. Kingman, S. Bradshaw, N. Miles

The influence of microwave energy at high electric field strength on ore strength has been investigated. Two ores of varying mineralogy were treated in both multimode and single mode microwave cavities. After treatment, any change in strength relative to untreated ore was established. It was shown that in the multimode cavity reduction in strength was related to the applied power level. Results of tests carried out in the single mode cavities showed that, for both ores, reductions in ore strength > 50% could be achieved in very short periods (<1 second). Thermal stress analysis showed that the mechanism of microwave induced failure was the development of tangential stresses at the grain boundaries. It was concluded that operation at 915 MHz with high power and short residence times gives feasible scale-up possibilities.

Jan 1, 2003

By Tingting Li, Guodong Li, Dezhou Wei, Cong Han, Shuling Gao, Yanbai Shen

"The aim of this study is to solve the problem of tungsten leaching from scheelite concentrate using sodium hydroxide in the presence of phosphate by an autoclaving process, which is one of main development techniques for tungsten recovery. To understand the applicability and leaching performance of tungsten leaching from scheelite concentrate, the main work is to investigate the effects of leaching parameters such as initial concentrations of NaOH and Na3PO4, reaction temperature, stirring speed, liquid-to-solid ratio and reaction time on WO3 dissolution rate to obtain the optimal leaching conditions. The chemical composition and phase composition of the samples were evaluated by means of XRF and XRD. The content of WO3 in the leaching residue obtained via washing, filtering and drying was analyzed by thiocyanate spectrometric method. The optimal leaching conditions were found to be NaOH concentration of 6.76 mol·L-1, Na3PO4 concentration of 1.35 mol·L-1, reaction temperature of 180°C, stirring speed of 400 rpm, liquid-to-solid ratio of 2.0 and reaction time of 90 min. The results showed that tungsten leaching rate of 99.20% was reached under the optimal conditions. The technology exhibited superior leaching performance, which not only reduced the consumption of sodium hydroxide but also greatly decreased reaction time in comparison to a conventional autoclaving process. INTRODUCTION Tungsten is an important strategic metal, which is widely utilized in a variety of industries due to its high melting point, superior mechanical properties and heat-resistant performance [1-3]. The chief raw materials for the production of tungsten in nature are wolframite and scheelite [4, 5]. Among them, approximately two-thirds of the tungsten reserves contain scheelite deposits in the world [6, 7]. With the progressive exhaustion of high-quality wolframite resources, scheelite has become a predominant primary resource for the production of tungsten [8, 9]."

Jan 1, 2018

By M. Ghiani, P. Massacci, A. Manni, R. Peretti, Center for Mining and Mineralurgical Studies, S. Serranti

"The research study has been devoted to point out the better way to process a colemanite concentrate characterised by a high content of two kind of As sulphides (realgar and orpiment) finely disseminated in the matrix of the boron mineral.Scanning Electron Microscopy analyses have been carried out to determine all the mineralogical constituents and to evaluate their distribution, frequency and mutual relationships.There is a not neglectable part of realgar highly disseminated into the colemanite crystals.The orpiment recurs in small acicular crystals of 20÷30 µm in length and 5 µm in width, very difficult to liberate from the colemanite mineral particles.The colemanite concentrate, in fact, has been addressed to a chemical plant which products were boric acid and gypsum, the contaminated by As. It was unacceptable for further applications and very dangerous if rejected in the environment.Grinding tests under dry and wet conditions have been carried out in order to evaluate the liberation modality of the As sulphide minerals from the colemanite.Some difficulties have been overtaken in order to obtain a selective liberation of the As sulphides from the colemanite even if a certain decreasing in As grade in the higher size classes has been registered.All the comminution products have been analysed by SEM monitoring the liberation degree of the mineralogical constituents.Flotation tests have been carried out after different comminution flow sheet adopting different modulating conditions in order to evaluate the influence of the following parameters:• collector concentration versus contaminants concentration (As minerals) in the froths;• kind and quantity of the frothing agent to maximise the mineral recovery in the froths;• kind and quantity of the modulating agents to maximise the selectivity of the process;• conditioning modalities in order to obtain a suitable dispersion of fines and to avoid segregation of Asbearingparticles together with colemanite ones;• solid/liquid ratio of the pulp to minimise the water flow into the plant."

Jan 1, 2003

By M. Nonaka

An attempt to produce fuel from biomass waste by hydrothermal treatment has been made with expectation to effectively utilize a renewable source of energy and an unexploited energy resource. In this paper, we have used empty fruit bunch (EFB) that is a residue after recovery of crude palm oil from palm oil trees in a plantation in a tropic country. A hydrothermal treatment was conducted using a 500 mL autoclave, with alkali (K2CO3) catalyst. EFB was decomposed more by alkali catalyst of lower alkali content and rather polymerized in higher catalytic region. The functional group composition of the solid phase oil is similar independent of alkali content, however its burning characteristic became more incombustible with increasing alkali content.

Jan 1, 2014

By Patricio Berrios, David Barriga, Peter Amelunxen, Gerson Sandoval, Roger Amelunxen

Laboratory test equipment and methods have undergone relatively little change since their introduction in the 1950’s. The Denver sub-A cell is still by far the most commonly used laboratory batch cell, but bottom-driven and instrumented cells are beginning to appear. Test procedures have also changed little over the past 60 years, and most engineers continue to rely on empirical models for predicted plant scale performance from laboratory scale tests. This approach is problematic, because changes in the froth phase characteristics in the lab hamper the reliability and reproducibility of the empirical extrapolation to plant scale. In a previous work, the authors reported on the impact of froth recovery at the laboratory cell, showing that variability in the froth recovery can lead to significant error in the modeled kinetic rate constant unless it is properly accounted for in the scale-up model. In this work, we review the role of hydraulic entrainment in the lab cell, and how variability in particle specific gravity, shape factor, and froth residence time can impact the entrainment recovery.

Jan 1, 2014

By A. M. Raichur, S. P. Vijayalakshmi

Fine coal dewatering is a major problem in the coal industry because the fines hold significant amounts of water and account for two-thirds of the product surface moisture. Present methods include flocculation using polymeric flocculants followed by filtration. But the kinetics of filtration are not favorable with such flocculants. In this study, a novel method involving flocculation with microorganisms for improving the settling rates and dewatering kinetics has been investigated. Two Gram-positive microorganisms, Mycobacterium phlei and Bacillus subtilis were used as bioflocculants. The effect of time, pH and ionic strength of the electrolyte used on adhesion and flocculation were studied. Adhesion tests showed that maximum adsorption occurred in 20 minutes in the pH range of 2-7 and the total amount of bacteria adsorbed was almost the same with both the flocculants. Nearly 98% of the coal flocculated with both the microorganisms in about 5 minutes. The filtration rate of fine coal almost doubled with addition of bacteria and electrolyte and maximum filtration occurred near the point-of-zero-charge of the bacteria.

Jan 1, 2003

By S. V. Balikov

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

Jan 1, 2014

By ?. S. Yusupov, V. P. Isupov, N. Z. Lyakhov, A. G. Vladimirov, S. A. Kondratiev

"Considering the shortage of lithium reserves, the desirability has been shown of studying the possibility of processing into spodumene concentrates the concentration tailings and low-grade ore reserves at Transbaikalian Mining and Processing Project (ZabGOK). The mineral and chemical composition and the particle size distribution of the mineral feed, separability and recoverability of spodumene have been studied, forming the basis for the development of a concentration process and flowsheet for the studied mineral resource. Prospects have been established for impact ore grinding in a disintegrator, the efficiency of mechanochemical processes has been shown in the separation of quartz and feldspar and the recovery of lithium from concentrates.INTRODUCTION Lithium is widely used in various industries - nuclear power engineering, defense technology, power generation, ceramics, special-purpose glass and other technologies. The increasing demand for lithium and its compounds (lithium carbonate and hydroxide) for use in the manufacture of high-tech products (lithium-based materials for chemical sources of electric power and electrochemical energy) explains the need for establishing a sustainable raw material base (Mohr et al. 2012; Moores, 2007). The main sources of lithium are spodumene pegmatites (Vladimirov et al. 2012; Garret, 2004), brines (salars), and hydromineral materials (Garret, 2004; Zeng and Liu, 2009; Kotsupalo and Ryabtsev, 2008). However, in view of the absence in Russia of extractable hydromineral resources of lithium, the practical significance of spodumene ores remains decisive. Russia controls some of the world’s largest reserves of rare-metal pegmatites holding over 8 million tons of Li2O: approximately 7 million tons on the Kola peninsula and 1 million tons in the Sayan mountains. The main lithium mineral in pegmatites is spodumene LiAlSi2O6 containing 8.1% Li2O. It can be efficiently recovered by gravitational concentration and flotation. However, most of the deposits are located in remote areas and none is currently developed. In this connection, special research is focused on technogenic raw materials, which are increasingly seen as a solution to the resource base problem (Chanturia, 2008, Kurkov and Kotova, 2007)."

Jan 1, 2018

By Ana C. C. Pacheco, Maurício L. Torem

"The aim of this work was the study of the effect of the type (Fe3+, Al3+ and Fe3+ and Al3+ mixture) and of the coprecipitant concentration on As5+ removal by ACF. Removal of As5+ using Sodium Dodecylsulfate (SDS) as a collector was studied by Adsorbing Colloid Flotation (ACF) at laboratory scale. The morphology of the precipitate retained in the foam was analyzed by Scanning Electron Microscopy (SEM/EDS).By comparing the coprecipitant type and concentration studies it could be concluded that As5+ removal in the presence of Fe3+ (160 mg.L-1) alone, or a Fe3+ (80 mg.L-1) and Al3+ (40 mg.L-1) mixture, is equivalently effective (99,9%) and greater than in the presence of Al3+ (80 mg.L-1) alone (97,7%).Through SEM/EDS analyses, it was possible to observe that both morphology and composition of the precipitate produced in the solution and retained in the foam depend on the physico-chemical conditions of the system, that is, on the species present in the solution."

Jan 1, 2003

By Alexander V. Epiforov, Stanislav V. Balikov

This paper describes of processing tests of 8 flotation concentrates containing precious and base metals by pressure oxidation (POX) method. Concentrates 1 and 2 were gold-copper concentrates. Zinc and silver were by-products in these concentrates. These concentrates were processed by POX with following cyanidation of POX solid residues, copper solvent extraction and zinc carbonate precipitation. The recoveries of gold, copper, silver and zinc from these concentrate were 98, 99, 30 and 99%, respectively. Concentrates 3 and 4 were lead-zinc concentrates. Silver was by-product in concentrate 3. Gold, silver and copper were by-products in concentrate 4. Concentrate 3 was processed by POX with following reduction smelting of POX cakes for primary lead recovery and zinc carbonate precipitation from POX solution. The recoveries of lead and silver into ingot were 96 and 98%, respectively. The recovery of zinc was 99%. Concentrate 4 was processed by POX with following zinc carbonate precipitation and reduction smelting of POX cakes. Copper was precipitated from the POX solution using sodium sulphide prior zinc precipitation. The recovery of zinc into carbonate concentrate was 99%. The recoveries of lead, gold and silver into lead bullion were 95, 99 and 98% respectively. The recovery of copper into sulphide concentrate was 99%. Concentrate 5 was gold-pyrite concentrate, and concentrates 6-8 were double refractory gold-arsenopyrite concentrates. Silver was by-product in these concentrates. These concentrates processed by POX and CIL. The recovery of gold from concentrate 5 was 98%, and from concentrates 6-8 was 85-90. Low gold recovery from concentrates 6-8 was associated with preg-robbing factor. The recovery of silver from gold-sulphide concentrates did not exceed 10-20%. The low silver recovery was associated with argentojarosite formation that is resistant to cyanidation. The flowsheets for complex processing of precious and base metals sulphide flotation concentrates based on this test work results were proposed.

Jan 1, 2018

By Liu Xiaodong, Jiang Meifen, Kou Xiaoyan, Lu Xianjun

"Two low grade calcium bentonite (Ca-bentonite) from Hejia (HJ) and Boteng (BT), Shandong, China, mainly contain montmorillonite and quartz, and slightly feldspar, calcite, and montmorillonite content are 32.37% and 55.90%, respectively. In order to examine the feasibility to produce activated clay to meet the standard of Chinese chemical industry (HG/T2569-2007), three types of industrial production process are proposed according to the screen test. The first process is grinding the raw bentonite to -200 mesh and then activating. The second is to activate the raw bentonite directly, and then classify the production to -200 mesh. The last one is to activate the raw bentonite after purification. Three parameters, which are sulphuric acid strength, ratio of liquid to solid and treatment time, are optimized via orthogonal test for every process. The preliminary test indicates that it is feasible to produce acid bentonite of difference quality. For the sample of HJ, milling the raw to -200 mesh first and activating afterwards, the parameter is at 9% acid strength, the ratio of liquid to solid 2, the time 4 hours, temperature 90°C, the products can meet the qualified grade for animal and vegetable oils refining. For the sample of BT, milling the raw to -200 mesh first and activating afterwards, the parameter is at 15% acid strength, the ratio of liquid to solid 2.5, the time 2 hours, temperature 90°C, the products can meet the qualified grade of high decolorization for petroleum products refining. For the sample of BT, adopting purification to -325 mesh by stirring and rubbing at first, and then activating at 9% acidity, under the conditions that the ratio of liquid to solid is 2, the time is 2 hours and the temperature is 90°C, the products can meet the first grade of high activation degree for petroleum products refining."

Jan 1, 2014

By T. Fujita

The incinerated bottom ash includes titanium oxides and a small amount of chromium oxides. Initially, 89 % of chromium was removed from ash at magnetic flux density at several stages of wet magnetic separation from 0.05 T to 6 T. The bottom ash from an industrial waste incineration process includes about 10 % of TiO2. The minerals that contain titanium are rutile and perovskite. These minerals are paramagnetic and show low magnetic susceptibility, therefore, the high magnetic field strength, created by a superconducting magnet, is necessary to capture these minerals. At the first stage of separation, a magnetic field strength of 6T was applied in order to attract the particles that contain iron oxides to the higher magnetic field region. At the second stage of separation, the magnetic capture rate of weekly magnetic particles decreased with increasing the magnetic susceptibility of the aqueous solution (i.e. the medium for dispersing the particles) in order to separate the highly magnetic materials. The para-magnetic salt such as ferric chloride was dissolved in the water to increase the magnetic susceptibility of the medium. After a two-step wet HGMS with superconducting magnet, the grade of TiO2was increased from 10 % to 28 % with a recovery of 64 %.

Jan 1, 2014

By R. Batterham

"The history of mineral processing is indeed long and has always been tied to mining methods of the day. While many new processes and devices have been developed, the most significant trend in recent times has been the ever increasing scale of operations. There is serious doubt this ongoing trend is enough to match the global trend of falling grades and rising energy demands needed for mineral processing. As well we see that globally, the social licence to operate for mining (and hence mineral processing) is far from secure. Indeed, some argue that mining as we know it is dying and in the future, mining will be much more environmentally secure, even to the extent of insisting upon “in place” recovery and only bringing to the surface the valued product. In this paper we comment on the ramifications of broader uptake of more sustainable mining for mineral processing methods and address the specific question of whether recent progress in in place recovery will lead to a timely breakthrough. The answer proposed is that this is possible in the next 10 - 20 years but in the meantime, the focus is just as likely to be on removal of waste much earlier in circuits. Either way, mineral processing as we know it will change significantly. INTRODUCTION Mineral processing has a long history well covered in a recent publication (Habashi, 2006) who points out “the ancient Egyptians knew already that it would be easier to melt an earth rich in gold particles than another which is poor. As a result, all efforts were made to enrich the gold by washing away the light gangue minerals. Similarly, another ancient method for enriching gold particles from a river stream was by means of fleece, hence the Greek myth of the “golden fleece”. Since the middle ages, striving for improved economics has driven a string of innovative developments in milling and separation. The reader is again referred to the informative and illustrative text by Habashi."

Jan 1, 2018

By Stoyan N. Groudev, Ioannis Paspaliaris, Irena I. Spasova, Kostas Komnitsas

"Acid mine drainage waters generated in the uranium deposit Curilo, western Bulgaria, were treated by means of a constructed wetland located in the deposit. The waters had a pH in the range of about 2.5 - 4.1 and contained uranium, radium, some heavy metals (iron, manganese, copper, zinc, cadmium, cobalt, nickel, lead), arsenic and sulphates in concentrations much higher than the relevant permissible levels for waters intended for use in the agriculture and/or industry. The wetland covered an area of about 140 m2 and was characterized by an abundant water and emergent vegetation and a diverse microflora. Typha latifolia, Typha angustifolia and Phragmites australis were the prevalent plant species in the wetland but species of the genera Juncus, Eleocharis, Scirpus, Carex and Poa as well as different algae were also well present.The water treatment was started in July 2001 and until now (August 2002) the wetland effluents were characterized by pollutants concentrations decreased below the relevant permissible levels. The water flow rate through the wetland during this period varied in the range of about 0.2 - 1.0 l/s. The water clean-up markedly depended on the temperature but was efficient even in the cold winter months at air and water temperatures close to 0oC. The removal of pollutants was due to different processes but the microbial dissimilatory sulphate reduction and the sorption of pollutants on organic matter (viable and dead plant and microbial biomass) and clay minerals played the main role. Negative effects of the pollutants on the growth and activity of the indigenous plant and microbial communities were not observed."

Jan 1, 2003

By Lev Malytin, Alexander Dyachenko, Roman Kraydenko

"A new method of processing beryllium raw materials based on interaction of beryllium mineral with ammonium fluoride was developed. Possibility of extracting ammonium tetrafluoroberyllate from bertrandite and phenakite minerals is proved. The developed technology allows replacing the outdated sulfate technology developed by Materion Corporation. INTRODUCTION Beryllium is a strategic metal, which is produced only in three countries: the USA, China and Kazakhstan. The world production of beryllium amounts to only 300 tons per year. More than 75% of the world beryllium production is manufactured from oxide beryllium ores in the USA by Materion Corporation. In Russia there is a large-scale beryllium ore deposit «Ermakovskoye» containig over 1 million ton of ore. Concentration of beryllium in ore amounts to 1,2%. Since the mineral base of ore is calcium fluoride, the sulfate technology of Materion Corporation cannot be used. After alloying and sulphuric acid leaching the fluorine-ion found in fluorite and beryllium together evolve into the aqueous phase. At the stage of Be(OH) precipitation fluoride-ion prevents complete isolation of beryllium into the phase of hydroxide, which increases the loss of beryllium. A new ammonium-fluoride method for processing high-fluoride beryllium ores has been developed at Tomsk Polytechnic University. The ammonium-fluoride method involves fusion of ammonium fluoride (NH4HF2) with a beryllium concentrate. The concentrate containing 10% of beryllium oxide and 25% of calcium fluoride is produced by flotation from the ore of Ermakovskoe deposit. The main component of the ore concentrate is fluorite CaF2, which does not react with NH4HF2. Beryllium in this concentrate is represented by two minerals: phenakite Be2(SiO4)2 and bertrandite Be4[Si2O7](OH)2. Ammonium fluoride, unlike sulfuric acid, does not react with calcium fluoride and is not consumed for decomposing the mineral base of the concentrate. Ammonium fluoride is an accessible and environmentally safe reagent, worth up to $ 1,200 per ton. The total amount of ammonium fluoride is recycled and is used for decomposing a new batch of beryllium concentrate."

Jan 1, 2018

By Lev Malytin, Alexander Dyachenko, Roman Kraydenko

"A new method of processing beryllium raw materials based on interaction of beryllium mineral with ammonium fluoride was developed. Possibility of extracting ammonium tetrafluoroberyllate from bertrandite and phenakite minerals is proved. The developed technology allows replacing the outdated sulfate technology developed by Materion Corporation. INTRODUCTION Beryllium is a strategic metal, which is produced only in three countries: the USA, China and Kazakhstan. The world production of beryllium amounts to only 300 tons per year. More than 75% of the world beryllium production is manufactured from oxide beryllium ores in the USA by Materion Corporation. In Russia there is a large-scale beryllium ore deposit «Ermakovskoye» containig over 1 million ton of ore. Concentration of beryllium in ore amounts to 1,2%. Since the mineral base of ore is calcium fluoride, the sulfate technology of Materion Corporation cannot be used. After alloying and sulphuric acid leaching the fluorine-ion found in fluorite and beryllium together evolve into the aqueous phase. At the stage of Be(OH) precipitation fluoride-ion prevents complete isolation of beryllium into the phase of hydroxide, which increases the loss of beryllium. A new ammonium-fluoride method for processing high-fluoride beryllium ores has been developed at Tomsk Polytechnic University. The ammonium-fluoride method involves fusion of ammonium fluoride (NH4HF2) with a beryllium concentrate. The concentrate containing 10% of beryllium oxide and 25% of calcium fluoride is produced by flotation from the ore of Ermakovskoe deposit. The main component of the ore concentrate is fluorite CaF2, which does not react with NH4HF2. Beryllium in this concentrate is represented by two minerals: phenakite Be2(SiO4)2 and bertrandite Be4[Si2O7](OH)2. Ammonium fluoride, unlike sulfuric acid, does not react with calcium fluoride and is not consumed for decomposing the mineral base of the concentrate. Ammonium fluoride is an accessible and environmentally safe reagent, worth up to $ 1,200 per ton. The total amount of ammonium fluoride is recycled and is used for decomposing a new batch of beryllium concentrate."

Jan 1, 2018

By Jin Duan, Qi Zheng, Bo Yan, Li-cheng Ma, Xiao-kui Che

"The aim of this study was to develop an approach for the selectively removal of Zn(II) from heavy metal-bearing-wastewaters produced by gold chemical industry using magnetic chitosan based adsorbents. Chitosan beads were prepared by cross-linked modification of Fe3O4-chitosan bead using epichlorohydrin as the cross-linker. Both Zn(II) inprinted magnetic chitosan hydrogel bead (Zn-MCHB) and unprinted magnetic chitosan hydrogel bead (MCHB) were prepared and batch adsorption experiments were used to evaluate their adsorption capacities and selectivities. Single adsorption of Cu2+, Pb2+, Zn2+ and Ni2+ showed that both MCHB and Zn-MCHB adsorbs all the four metals and had an affinity in the order of Cu2+ > Pb2+ > Zn2+ > Ni2+. The binary adsorption revealed that the two adsorbents preferentially adsorb Cu2+/Pb2+ from Zn2+-Cu2+/ Zn2+-Pb2+ binary systems. However, MCHB does not show any selectivity in Zn2+-Ni2+ system. In contrast, Zn- MCHB preferentially adsorbs Zn2+ from Zn2+-Ni2+ binary system. Furthermore, the adsorption selectivity is predictable from stability constants of the metals with the amino group of chitosan. The information revealed in this work can be used for further applications in the treatment of wastewater containing a range of heavy metals from the gold chemical industry."

Jan 1, 2014

By S. M. Bradshaw, C. Dorfling, F. Ngandu

"Selenium removal and platinum group metal (PGM) recovery from copper sulfate leach solutions form an integral part of platinum producers’ base metal refinery operations. Recent studies found that substantially higher PGM recoveries can be achieved if excess thiourea instead of sulfurous acid is used as precipitation reagent; however, the impact of the proposed process on downstream processes was unknown. The objective of the current study was therefore to investigate the effects that excess thiourea and residual selenium present in the rich copper sulfate electrolyte have on copper electrowinning. This was achieved by performing cyclic voltammetry tests using a platinum rotating disk as working electrode, a graphite auxiliary electrode and a saturated calomel reference electrode. A synthetic copper sulfate electrolyte containing 0.01 M Cu2+, 1 M H2SO4 and varying amounts of selenium (20 – 150 mg/L) and thiourea (10 mg/L and 100 – 300 % excess) was used. The results obtained have shown that both thiourea and selenium interact with the cupric ion cathodic reduction reaction. Thiourea polarized the working electrode and lowered the efficiency of the cupric reduction reaction. Tests on the solutions with high selenium concentration showed simultaneous reduction of copper and selenium, which led to copper selenide formation. Nodule formation on the deposit surface also became significant at higher selenium concentrations. Based on these results, it was concluded that precipitation of PGMs from the copper sulfate solution with excess thiourea is not feasible due to the detrimental effect on downstream electrowinning operations.INTRODUCTION During the high pressure oxidative leaching stage in a typical base metal refinery (BMR) circuit, selenium, tellurium and some platinum group metals (PGMs) are dissolved along with the desired copper ions. The impurities (selenium and tellurium) may pass to the copper electrowinning circuit where it can result in the formation of nodular copper electrodeposits or reduced current efficiency. The presence of PGMs in solution, on the other hand, increases the precious metal inventory and potential losses of the operation. Currently, sulfurous acid is widely used for selenium and tellurium precipitation from copper sulfate leach solutions. Sulfurous acid is capable of reducing cupric ions to cuprous ions; the generated cuprous ions then react with Se4+ and Se6+ to precipitate selenium as a copper selenide (Bello, Dorfling, 2014). Recent studies have, however, proposed the use of excess thiourea as a precipitating agent for the removal of selenium and tellurium; it has the advantage of improved PGM recovery in the Se/Te precipitation step (Lottering, Eksteen et al., 2012; Mulwanda, Dorfling, 2015)."

Jan 1, 2018