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Much has been published in recent years on the challenges facing the engineer of the 21st century and the consequent challenges facing the education providers and industry. The world is changing rapidly on the social, economic, political, technical and environmental fronts and our metallurgists/mineral processing engineers of the future will arguably need to have a different set of skills than those which were traditionally regarded as adequate during the latter half of the 20th century. This paper attempts to explore some of these issues and their ramifications.

Jan 1, 2014

"A sulfide ore containing 2.19 g/t Au and 13.13 g/t Ag was studied for the deportment and extractive metallurgy of pay metals. The ore is composed mainly of non-sulfide minerals with approximately 10% pyrite and minor amounts of marcasite, chalcopyrite, galena, sphalerite, and arsenopyrite. The deportment study using a comprehensive approach showed that gold and silver minerals in the ore are mainly native gold (Au), electrum (Ag, Au), petzite (Ag3AuTe2), hessite (Ag2Te) and volynskite (AgBiTe2) with minor amounts of acanthite (Ag2S) and argyrodite (Ag8GeS6). At current grinding fineness (P70@150µm), the majority of liberated native gold and electrum particles are larger than 20µm with 85% >40µm, some of them are even larger than 150µm, indicating that extended leach time may be required to dissolve these coarse gold grains to completion and gravity concentration needs to be considered. Petzite is relatively fine with all liberated particles being less than 80µm in size. Gold distribution and leach testwork results showed that liberated gold accounts for 50% of the head assay, exposed gold associated with sulfide and non-sulfide minerals accounts for 30% and 11% of the head assay, respectively. Gold locked in sulfide and non-sulfide accounts for 4.8% and 4.1%, respectively. Thus, it is recommended that cyanide leaching be considered for extracting gold in this ore. This paper presents the detailed results of the gold and silver deportment study and extractive metallurgy testwork, and discusses the impact on process selection and flowsheet development.INTRODUCTION Gold and silver have been found in various geological environments and ore types. In terms of the gold deportment and extractability, gold ores can be classified into two categories and 12 types. Their extractive metallurgy is largely driven by mineralogical factors such as particle size, association with other minerals, coatings and rimmings, presence of cyanicides, oxygen consumers and preg-robbers, presence of refractory gold minerals and locking of submicroscopic gold in sulfide and sulfarsenide mineral structures (Zhou et al., 2004; Marsden & House, 2006; Zhou, 2013a, 2013b). Compared to gold, silver mineralogy is more complex and metal extraction is more challenging due to its presence in over 200 Ag-bearing minerals (a silver-rich ore body would typically contain 5 to 10 different silver minerals) (Gasparrini, 1984), presence of refractory silver minerals (such as pyrargyrite, prustite and stephanite) and submicroscopic silver in sulphide minerals (such as galena and copper minerals) (Chryssoulis et al., 1985; Di Prisco et al., 2002). The presence of these issues often causes problems in Ag-bearing ore processing. Therefore, knowing the deportment of gold and silver in the ores and metallurgical products is important for flowsheet development and process optimization."

Jan 1, 2018

By I. A. Khabarova

The paper presents the results of experimental investigations of the effect by high-power nano-second electromagnetic pulses (HPEMP) on change of surface composition, physical-chemical and flotation properties of pyrrhotite, and search of optimal combination of flotation reactants providing increase of metal extraction from disseminated copper-nickel ores. Preliminary electropulse effect on products before flotation allows to create optimal conditions for flotation separation of pentlandite and pyrrhotite owing to forming the new phases and defects on the surface of sulphides. It was established, that the yield of monomineral fractions of pyrite, pyrrhotite and chalkopyrite increases by 5%~15 %, compared with flotation with one butyl xanthenes (BX), during usage of combination of thionic carbamate (isopropylmethylthionocarbamate (ITC) or isopropylethyl- thionocarbamate (Z-200) with butyl xanthate (BX). Sorption of investigated reactants and strength of their fixing on the surface of minerals are examined; it helped to reveal the cause of rise of floatability of sulphide minerals for the case of combination of reactants. Preliminary high-pulse processing and reactant conditions with usage of combination of thionic carbamate and xanthate have been tested on the sample of disseminated copper-nickel ore from Norilsk region.

Jan 1, 2014

By Chanki Jeong, Heechan Cho

In this study, batch flotation experiments were conducted to recover rare earth elements (REE) from an REE ore deposit in the Mushgai Khudag area, Mongolia. Apatite was found to be the major REE mineral while monazite and xenomite were observed as minor minerals. The main purpose was to float apatite to increase the REE content. Oleic acid was used as collector and MIBC was used as frother. Sodium silicate was used as a depressant for iron compounds. NaOH was used to adjust the pH of the flotation to about 9.5~10.0. The particle size was 90% less than 152 ?m (- 100 mesh). The content of REE elements was analysed by ICP - OES to determine the contents of REE such as Ce, La, Nd, Pr and Y. When sodium silicate was used, the best result was obtained using 0.2 mL of sodium silicate, 2 mL of oleic acid and 2 mL of MIBC. Under these conditions, the recovery of REE was about 78% and the grade of the first froth product was 42,995 ppm Ce, 24,088 ppm La, 14,236 ppm Nd, 4,388 ppm Pr and 1,570 ppm Y. When conducting flotation after iron minerals were removed by magnetic separation, a better result was obtained; with 2 mL of oleic acid, 2 mL of MIBC and no sodium silicate, the recovery was increased to 90% and the grade was also increased to 58,600 ppm Ce, 28,624 ppm La, 17,160 ppm Nd, 5,348 ppm Pr and 1,971 ppm Y.

Jan 1, 2014

By Vitaly Kretov, Valery Golikov, Vladimir Ryaboy, Elena Shepeta

"The aim of the study was to develop and implement more efficient dialkyldithiophosphate collectors. A hypothesis has been formulated on the action of dialkyldithiophosphates in the flotation of sulphide ores based on the assessment of their complexing properties and surface activity at the liquid-gas border. Dialkyldithiophosphates form weaker complexes with cations of heavy metals compared to xanthates, and are characterized by a large difference in the strength of their complexes with heavy metal cations. It provides higher selectivity of dialkyldithiophosphates in the separation of non-ferrous metal sulphides and iron sulphides from them. To make a quality assessment of the impact of surface activity of the synthesized collectors on their selectivity of action and collecting power, an assessment of the surface tension of their aqueous solutions has been performed using the stalagmometer method. It is well known that a stronger collector is characterized by a greater surface activity, in contrast to a more selective collector which is characterized by a lower surface activity. Based on studies, new dialkyldithiophosphate collectors BTF 163 and BTF 1552 are suggested for use in the flotation of sulphides to increase the recovery of copper, gold and silver. The collector BTF 163 was introduced at a gold processing plant in Kazakhstan and has been undergoing industrial testing in the sulphide cycle of scheelite-skarn-sulphide ores in a factory in the far east of Russia. The collector BTF 1552, which increases the recovery of silver and gold by 3%, was approved for industrial testing in another far-eastern factory. A modification of the reagent BTF 1552, the new reagent BTF 1522, was introduced in the flotation of copper-zinc ore at a factory in Kazakhstan."

Jan 1, 2014

By Andrey V. Yevdokimov

The results of the test work on gold leaching using high strength of cyanide solution in the presence of activators are presented. The major physical and chemical relationships of gold leaching were determined. Intensive cyanidation results of gravity concentrates with various material compositions using an activator are given. Process design criteria for intensive cyanidation unit were developed on the base of pilot plant test results. Keywords: gold, activators, intensive cyanidation, gravity concentrates

Sep 1, 2012

By J. K. Afidenyo

Microbial pretreatment of a double refractory gold ore was investigated using a white rot fungus, Coriolus versicolor (ATTC 20869) to improve gold extraction. A series of preliminary tests were conducted to determine the amenability of the ore to fungus degradation. Pulp density, temperature, pH and retention times were the process variables considered. Both a single. Stage and two-step processes were studied to establish process conditions. During the single stage process, the ore was contacted with the fungus culture (biotic) for 3weeks at 30°C and pH 10.0. For the two-step process, the first step involves ore contact with the culture medium (abiotic) at 45°C and pH 10.0 followed by fungal contact (biotic) at 30°C. Results indicated that carbonaceous matter is not degraded but passivated by the fungus; as preg-robbing activities dropped significantly from 18.1% to below 1.0%. The single stage process resulted in higher gold recovery than the two-step process Gold ex-traction by cyanidation of the pretreated ore was 78.9% for the single stage compared with 15.0% for the untreated ore. In the two-step process, only 73.3% gold extraction was achieved. Application of the two step process conditions to a high grade flotation concentrate, resulted in 93.3% gold ex-traction as against untreated concentrate of 30.5%. The results show that the microbial pretreatment process degraded sulfide sulfur, passivated carbonaceous matter and enhanced gold extraction from double refractory gold ore and concentrate.

Jan 1, 2014

By S. S. Lu

In sodium oleate and salicylhydroxamic acid systems, sequence of floatability of three sulfate minerals is: barite>celestite>gypsum, in lauryl amine system is: gypsum>celestite>barite. XPS and infrared spectrum analysis are used to find that it is chemical reaction between sodium oleate and three minerals, physical adsorption on the surface of minerals for lauryl amine, and it is chelating reaction between salicylhydroxamic acid and three minerals possibly. Theoretical calculation on chemical bonds in crystal structures of three minerals is carried out, finding that breaking degree of Mn+?O2- in crystal structures of three sulfate minerals is closed related with floatability.

Jan 1, 2014

By C. Hazotte, M. -O. Simonnot, V. Houzelot, B. Laubie, M. Guilpain, B. Jally

"More than 400 plants in the world are able to accumulate nickel (Ni) in their aerial parts at concentrations higher than 10 g per kg of dry biomass. Then, Ni can be recovered from the biomass of these hyperaccumulator plants. Agromining is a chain consisting in growing such plants and producing metal. It has been developed thanks to the combined know-how of agronomy and hydrometallurgy. Hydrometallurgical processes have been designed; two different pathways were developed: i) acid leaching of ashes after biomass combustion and ii) direct Ni water extraction after biomass grinding. Experiments with 2 species of Ni hyperaccumulators (A. murale from the Balkans and R. bengalensis from Malaysia) have shown that the first process is very robust and transposable. For the second one, the way Ni is stored in the plant has a strong influence on extraction. A. murale transformation was explored in more details from the lab scale to the pilot scale. Results proved that Ni is totally transferred into the aqueous phase after 2 hours at 90 °C. A life cycle assessment has clearly shown the environmental interest of this approach (especially by detoxifying agricultural soils). INTRODUCTION Agromining is an emerging phytotechnology aiming at recovering metals thanks to hyperaccumulating plants (van der Ent et al., 2018). These plants are able to accumulate very high concentrations of metal in their aerial parts. For example, for nickel, the hyperaccumulation threshold is 0.1%, i.e. 10 grams of metal per kilogram of dry biomass. There are some advantages to use this technology instead of conventional mining techniques: •,Treating unconventional metal resources: they can be (i) natural soils with low concentrations but also low fertility because of the metal presence (like serpentine soils for Ni) (Bani et al., 2015), (ii) mine tailings or (iii) industrial wastes (like surface treatment sludge) after soil construction (Rue, 2017);"

Jan 1, 2018

By B. McFadzean

Thiol collector mixtures are commonly used in the flotation of base metal sulfides and platinum-bearing ores. There are synergistic benefits claimed for these collector mixtures, with mixtures lowering total dosage requirements, improving coarse particle recovery and the rate of flotation. Some of the secondary collectors are known to have frothing effects. Therefore, a previous study aimed to decouple froth and pulp effects and considered only the bubble-particle attachments in the pulp. This was achieved through microflotation experiments, where various ratios of the ethyl and isobutyl chain lengths of xanthates, dithiocarbamates and dithiophosphates were used as collectors. Results from these experiments showed a marked increase in galena recovery using mixtures of xanthate with either dithiophosphate or dithiocarbamate. In the current study, the froth phase was introduced in the form of batch flotation studies to determine whether benefits noted during microflotation experiments would be carried through to a more realistic froth flotation scenario. The findings show that there was no significant difference in the flotation recovery between the best single collectors and mixtures of these collectors. In addition, single collectors that behaved poorly in microflotation experiments showed good flotation performance in the batch flotation experiments. The differences between the microflotation results and the batch flotation results are attributed to differences in particle froth stabilizing properties and/or differences in energy input between the microflotation and batch flotation cells. In addition, galena and pyrite showed marked differences in frothing effects and in their preferences for collector type. Keywords: froth flotation, collector mixtures, sulphide minerals

Sep 1, 2012

By E. V. Zelinskaya, A. E. Burdonov

"In the raw aluminum production processon electrolyzers with self-baking anodes, during technological operations, flow-regulated treatment and adjustment of the electrolyte chemical composition produced by diesel and floor-rail technology, alumina and fluoride salts are spilled through the ventilation grilles of the electrolysis body aeration channels, leakage in the equipment, machinery and motorized machinery construction elements. As a result of this, a part of the bus station and the level of ± 0 (the basement room with supporting curbstones and columns, on which electrolyzers are installed) are covered with technological raw material. This leads to an increase in the conductor (bus) resistance, contact connections deterioration, cell insulation failure and raw materials loss. For example, at the enterprise ""Rusal-Bratsk"", as a result of the activity, up to 420 tons per month of alumina-containing wastes that are unfit for use in the electrolysis process are formed due to the alumina substantial contamination with salts of iron and silicon, which significantly reduces the technical and economic parameters of the electrolysis process. The main scientific problem is the processing (cleaning) of alumina-containing waste having a complex mineralogical composition of the material, its contrast and structure, from oxides of iron and silicon, using ""dry"" separation methods. In the performing the work process, qualitative indicators of the influence of external factors (air humidity, precipitation, sunlight, temperature difference) on the physicochemical processes components interaction during the raw materials storage are established. On the basis of laboratory tests, it is established that alumina-containing raw materials can be separated and photometric and gravitational separation methods used. A mathematical model of the motion of particles of primary and prepared alumina-containing raw materials in a cascade-gravity classifier was developed. The criteria and factors characterizing the alumina-containing material, as well as influencing and determining the maximum material enrichment, are established."

Jan 1, 2018

By V. E. Vigdergauz

Possibilities of the analysis of potential energy of bubble-particle interaction on the basis of extended DLVO theory are discussed. Interaction energies of mineral silica, coal and pyrite particles with contact angles from 20 to 100 degrees and air bubbles in neutral and alkaline solutions are analyzed. It is shown that for hydrophilic silica, the formation of particle-bubble complexes (either in the area of primary or secondary potential minima) is impossible due to the high potential barriers (from 6000 to 140000 kT)and low values of the attraction force. There is shown that for coal and pyrite particles the energy of interaction vs. particle-bubble distance strongly depends on pH value. Improvement of floatability of minerals by the hydrophobization of their surface by collectors leads to the eliminating of potential barriers and the formation of the stable complexes of minerals and air bubbles in the area of primary potential minimum. In addition, due to hydrophobization of coal and pyrite changing pH has practically no influence on the energy of interaction. There are discussed results of the numerical calculations of normalized interaction forces in particle-bubble systems on the nanometer distances for pyrite, sphalerite, and chalcopyrite at contact angles up to 85° in comparison with the experimental data. The values of adhesion force for hydrophobized particles of coal and pyrite to air bubbles are about 10-5 N, that indicates stability of the formed flotation complexes. Examples of using the long-range and high-energy effect of hydrophobic interactions for improving minerals floatability are discussed. There are illustrated minerals wettability changes induced by the adsorption of thiol collector and hydrophobic polymer. Possibilities of flocculation and flotation of mineral slimes due the polymer-collector hydrophobic interactions are shown. Keywords: bubble-particle interaction, contact angle, extended DLVO theory, floatability

Sep 1, 2012

By J. Zou

With the increasing applications of the Chinese metal magnetic liners, the requirement for their technical specifications has become stricter and stricter. It is necessary to deeply analyse the metal magnetic liners in order to improve their reliability and economics. The key issue of metal magnetic liners is magnetic forces: their attractions to mill shell and protection layer. Mathematical models were developed using finite-element analysis software and systematic simulation analyses of the two most critical indexes of magnetic liners were performed to find out the most important parameters and their effect on the liner?s characteristics. The relationships between magnetic forces and installation gaps, thickness of permanent magnets and thickness of protection layer may be used to guide the design of metal magnetic liners for ball mills in order to achieve good economic indexes.

Jan 1, 2014

Large quantities of electrical and electronic equipments are discarded at the end-of-life (EOL) in different industries and house hold applications. Their disposal is creating a challenging environmental problem due to the presence of hazardous metals and non-metallic materials such as epoxy resins and plastics including brominated flame retardants etc particularly in printed circuit boards (PCBs). The recovery valuable metals and separation of hazardous materials from discarded PCBs will not only conserve the resources but also mitigate the environmental pollution. From a recycling point of view for a sustainable society, the studies have been carried out for the recovery of metal values from EOL-PCBs of DVD video players (DVD-PCBs) and vacuum cleaners (VC-PCBs) using physical beneficiation techniques. The concentration of the metals was found to be 43 wt.% and 30 wt.% in DVD-PCBs and VC-PCBs, respectively. Both the PCBs samples were ground in a hammer mill after cutting and their size analysis was carried out. The liberation of the metallic and non-metallic fractions was investigated under the microscope for different size fractions. Subsequently, froth flotation and pneumatic air separation studied were carried out to enrich the metallic constituents. The studies showed the enrichment of metals in coarser particles (-1000 +150 µm) and non-metals in the finer particles -150 µm following mechanical recycling process. A grade of 90 % with 75% recovery was achieved by froth flotation, but lower grade and recovery of 75% and 65% respectively were obtained by pneumatic separation with -1500 µm powders for DVD-PCBs. A grade of 86% metal was obtained by froth flotation with 76% recovery in the case of VC-PCBs. The lower grade of 65% with almost same recovery was achieved during pneumatic separation studies due to lower concentration of metals in VC-PCBs. The studies also showed that the froth flotation is an efficient process for enrichment of metals from both the samples. The process is environmental friendly as no chemicals are employed during entire processing. The higher recovery and grade of metal is expected to be obtained on a large scale beneficiation process.

Sep 1, 2012

By Y. Foucaud, I. Filippova, L. Filippov

Flotation is the most common beneficiation route for tungsten ores, but the predominance of minerals with similar surface properties poses a real challenge. This study investigates the use of Falcon centrifugal separators for gangue minerals removal and desliming prior to flotation. A typical tungsten-bearing skarn ore (Tabuaço, Portugal) is used, containing scheelite (CaWO4) associated with dense calcium-bearing silicates (vesuvianite), fluorite, apatite and light silicates (feldspars) as gangue minerals. Two types of Falcon bowls (Falcon UF and Falcon SB) were used, prior to flotation tests with fatty acids as collectors and a combination of sodium carbonate and sodium silicate as gangue minerals depressants. This flowsheet was compared with a reverse configuration, using flotation prior to Falcon concentrator, and with a classical flowsheet, using hydrocyclone and flotation. Gravity pre-concentration and desliming with the Falcon concentrator were optimized through the design of experiments (DoE) methodology. The influences of pulp density, rotational speed, feed flowrate and fluidisation pressure (for the Falcon SB) were studied for both bowls. Results showed that the Falcon can be used as a desliming and pre-concentrating apparatus. Separation performances depend mainly on rotary speed and pulp density for the Falcon UF, whereas rotary speed and fluidisation pressure are the main parameters for the Falcon SB. With only one stage of Falcon SB, an enrichment ratio of 7.9 is obtained with a 65% WO3 recovery while 89% of apatite and fluorite are rejected, which constitutes a very good selectivity. In comparison, Falcon UF shows lower enrichment ratios, around 2.7. In terms of desliming, Falcon UF is as efficient as the hydrocyclone, while the Falcon SB is significantly more efficient. Hence, the use of Falcon SB prior to flotation improves the overall performance of the tungsten flotation route by desliming the flotation feed while pre-concentrating the scheelite. It impacts very advantageously the flotation, as most of the calcium-bearing minerals are eliminated. A Falcon SB stage prior to flotation yields a final concentrate grade of 29.7% WO3 with a recovery of only 36.8 % that should be optimized as the reagents dosages were not adapted to the pre-concentrated feed.

Jan 1, 2018

By A. S. Choudhary

India is endowed with a reasonably large reserve of chrome ore, amounting to about 100 million tons in different grades and is one of the leading producers of ore/concentrate along with South Africa, Kazakhstan and others. Chromite production in India has registered a 10 times increase during the last 30 years. Till late 1940s, the practice was to mine the high-grade ore for making ferrochrome and charge chrome and export the same leaving the low and medium grade ores stacked around the mine as waste material. This practice was continued for quite some time till the increase in world wide de-mand of chromite initiated an increase in production of chromite ore in India and the aspects of ?economy of operation? and ?conservation of mineral? dawned on the producers and associated authorities. Recovery of chrome values from low and medium grade ore was initiated to produce high-grade saleable concentrate. Suitable legislation was made by Government of India to induce and provide incentives to the operators to do so. First large scale chromite beneficiation plant in India was commissioned about three decades back at Byrapore in the Southern state of Karnataka. The ore, rather simple in nature and relatively free from clay, was obtained from underground. The processing consequently was simple washing and wet tabling. Further more, some test work, on the samples from eastern coast particularly from Orissa, was carried out at an internationally reputed laboratory by a large private operator and the plant was set up. This consisted of a combination of crushing-grinding-de-sliming-classification followed by concentration in multi stage spiral concentrators and shaking table. The plant was commissioned and operated successfully. Soon a state owned company followed suit on the basis of test work carried out in a well equipped Indian laboratory. The plant worked satisfactorily after some initial teething trouble and the company plans to double the existing plant capacity by installing balancing facilities. This development prompted a number of operators, including industry leaders to join the fray and the chrome ore processing industry flourished in India. More and more operators are now in the process of adding capacities to their existing plants. This paper attempts to chart the evolution of chrome ore beneficiation industry in India through a case study with reference to an existing plant. The process technology adopted and modified to suit the altered requirements as well as the future trends are outlined.

Jan 1, 2014

By V. A. Luganov

The processing of pyrite concentrates may be accomplished with a preliminary enrichment due to withdrawal of iron sulfides from initial concentrate. The withdrawal of iron sulfides is possible with the help of magnetic separation when the magnetic iron sulfides (pyrrhotites) separate from non-magnetic fraction?the sulfides of non-ferrous metals. The conversion of non-magnetic pyrite to magnetic pyrrhotites of adjusted structure occurs because of roasting with the withdrawal a part of sulfur. The resultant pyrrhotites are strong ferromagnetic with hexagonal structure. This paper contains results of studying magnetic properties of pyrrhotites, findings of the de-composition of initial pyrite with the consequence homogenizing of product and hardening of product of decomposition in liquid nitrogen. Define the type of pyrrhotites super lattice structure, was established that at increasing of sulfur contention the magnetic properties of pyrrhotites are change ? pyrrhotite begin evince the properties of antiferromagnetic material and magnetic susceptibility of material is interruptive changing from 0 to 1.7×10-3~1.9×10-3 SI units/g system. The paper contains results of technological researching of pyrrhotizing-magnetizing roasting of pyrite concentrates in different condition: in unoxidizing, with the decreasing contention of oxygen and in autogenous regime in air atmosphere. Established that the magnetic susceptibility of products depend of pyrite degree dissociation and achieve 2×10-3~3×10-3 SI units/g system at 40%~45% desulphurization. Defined the conditions of steady regime autogenous roasting, at resultant products possess coefficient of magnetization, and this can be used magnetic separation for separating ferromagnetic pyrrhotites from gangues and non-magnetic sulfides of non-ferrous metals. This technology of pyrrhotizing roasting is recommended for effecting in industrial condition. The initial materials and products of roasting subjected to chemical, thermogravimetric, Moss-bauer spectroscopy and X-ray analysis, defined the magnetic properties.

Jan 1, 2014

By M. A. Naimanbayev, Zh. A. Baltabekova, N. G. Lokhova

"Integrated utilization of natural resources using energy-saving and waste-free technologies is relevant around the world. This applies to such type of raw material as titanium-magnetite ore. Shortage of ilmenite raw materials in the Republic of Kazakhstan for titanium slag melting challenges the search of alternative titanium-containing raw materials, titanomagnetites in particular. At the same time there are deposits of indigenous titanomagnetite ores of huge reserves, which can become the complex raw material for production of iron, titanium and vanadium. Tymlay field in the South-Eastern part of Chu-Ili watershed is of particular interest, within the Karasay ore unit with iron content of 30,4 and titanium dioxide - 10%, and after beneficiation titano-magnetite concentrate contains Fe-52,4; TiO2-of 16,06%. Technology of titanium-magnetite concentrate ore-thermal melting was developed at Tymlay deposit producing hot metal and low titanium slag that can become the raw material for titanium production using chloride method. Slag produced in course of titanomagnetite ore-smelting consists mainly of alumosilicates and sodium silico titanates, magnesium and calcium titanates, titanium dioxide and lower titanium oxides. Following operations were carried out to improve titanium slag quality: slag sintering with sodium hydroxide; cake leaching with water to remove water-soluble impurities; to reduce the number of operations impurities leaching with hydrochloric acid was combined with pyrohydrolysis of titanic acid; rutile concentrate production (85% TiO2, 7,66% SiO2), silica removal by means of sodium hydroxide solution treatment, finally sellable rutile concentrate was produced with titanium dioxide 91-92% and silica 1,7 to 2,0% content. Based on research results technological scheme of titano-magnetite concentrate processing at Tymlay deposit was proposed producing hot metal and certified rutile concentrate, suitable for chlorination."

Jan 1, 2018

By V. A. Chanturiya

There is an obvious disproportion in Russia gold-mining industry between the structure of gold reserves and the gold-mining techniques. The bulk of gold (over 70%) is extracted from alluvial deposit, 20%~25% from bedrock occurrences and 5%~7% from complex copper-zinc and polymetallic ores. The proved gold-bearing reserves are distributed as follows: alluvial ?19%, bedrock ?80%, the latter being sufficient to supply gold at the current production level for 100 years to come. In Russia, like elsewhere in the world, development of primary gold deposits is considered a first-priority line of development for gold-mining industry. Most of the gold-containing ores characteristic of Russian gold deposits are resistant ores with gold content varying between 3 ppm and 5 ppm usually showing quite low gold and silver recovery by cyanidation. Processing resistance of gold-containing mineral complexes is related to the presence of gold particles of submicrometric (<1.0 µm) and nanometric (10~100 nm) size, mostly asso-ciated with pyrite and arsenopyrite. Therefore switching over to processing bedrock entails a num-ber of technological problems that stem from chemical and physical resistivity of ores. The chemical resistivity can be overcome through flotation, oxidation roasting, autoclave leaching (biohydro-metallurgical oxidation), etc. The physical resistivity can be dealt with by resorting to traditional power-intensive methods,e.g. grinding to a ?40µm size, thermal treatments of sulphide concentrates and non-traditional high-energy methods of liberating the sulphides by means of power attack, e.g. electro-chemical oxidation, exposure to electron beam radiation and other methods. The modern non-traditional processes for treatment of rebellious noble metal-bearing ore, aimed at improving disintegration efficiency for finely dispersed mineral complexes are analysed. Advantages and perspectives of a new high-efficient, energy-saving and ecologically safe process of the nanosecond high-power electromagnetic pulses (HPEMP) effect are proved. The said process provides a stable increase in the valuable component recovery at lower energy consumption and production costs of the finite product at the stage of dressing rebellious gold-bearing ores and concentrated products. The experimental data prove high efficiency of the preliminary impulse treatment, thus, the increase in the gold recovery at the cyanidation stage amounted to 4%~12% from rebel-lious ores, 10%~30% from gravity concentrates, 5%~45% from flotation concentrates, 30%~80% from oil flotation tailings.

Jan 1, 2014

The present study investigates the nature of gold mineralogical distribution and association in a Cu-Au flotation circuit (1.6 g/t Au, feed head grade). The approach used in this study involves a combination of bulk gold assay and quantitative X-ray diffraction (QXRD), on a size-by-size basis, followed by QEMSCAN analysis on the -75+53 µm, -53+38 µm and -38+20 µm size fractions. The QXRD and QEMSCAN analysis showed that gold mineralogy is complex and fine-grained in the flotation feed, particularly in the -38+20 µm size fraction. Gold and pyrite concentration follow each other closely, suggesting gold-pyrite association. Quartz as the major gangue mineral shows a bi-modal distribution with respect to particle size. The identified gold particles were completely locked (~94 %) in pyrite as fine inclusions and 6 % locked as complex aggregates (-75+53 µm size fraction) whereas in the -38+20 µm size fraction, the identified gold grains occur as fine inclusions (~1-3 µm) within pyrite and silicates. However, in the intermediate size fraction (-53+38 µm), 75 % of the identified gold grains (10-22 µm) were over 90 % liberated and 25 % associated and locked in pyrite and complex aggregates (~2-4 µm). The identified gold grains in the -53+38 µm size fraction occurred as native and electrum with varying silver concentration between 8 and 20 % by weight. In the -38+20 µm size fraction, a gold telluride was also identified. Based on the mineralogical analysis of the feed, possible flotation strategies are discussed. Optimizing the flotation of pyrite is suggested as the primary Au recovery strategy. Keywords: flotation, pyrite, gold, liberation, QXRD, QEMSCAN

Sep 1, 2012