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By Mario Gerards

Maturing populations create an increase in the demand for the type of non-ferrous metals, which are used in the manufacturing of electronic products. The depletion of existing high-grade deposits, make it necessary to boost the efficiency of beneficiation processes and most importantly the yield (recovery rate of the value commodities). Increasing demand now makes it worthwhile to tap into resources, which were previously considered waste, for example secondary deposits such as tailing ponds and electronic waste. In order to achieve increased yields, increasingly finer, or even ultra-fine fractions have to be treated. In most cases, mineral characteristics such as size and density call for non-mechanical, high-capacity technologies to be used. There are various technologies that allow for the separation of fines and ultra-fines. This paper outlines the dynamics and efficiency of pneumatic flotation with emphasis on its application to the recovery of non-ferrous metals, from electronic waste. The conventional agitator flotation technology of an electronic waste processing plant was not able to recover ultra-fine non-ferrous metals satisfactorily, due to losses within the tailings. Because of its adaptability and its inherent suitability for non-ferrous metals, pneumatic flotation has been incorporated to the existing process achieving excellent results. The merit of the process lies in the state of the art technology, which does not utilize compressed air and any rotating parts, which results in a reduction in wear and energy requirement, both features of this innovative technology. In addition, the use of pneumatic flotation resulted in a significant reduction in residence time, when compared to mechanical flotation. Pneumatic technology is able to influence air bubble size in a flotation cell, by changing the speed and pressure of the slurry stream. Higher slurry velocities lead to the formation of smaller air bubbles, which is required for the successful flotation of ultra-fine particles. Conversely, decreasing the slurry velocity leads to bigger air bubbles. The optimum bubble size can be created according to different feed characteristics. This is important in case of feed fluctuations.

Sep 1, 2012

By G. H. Lind

The Minerals Tertiary Education Council (MTEC), over the past decade, has worked closely with Australian higher education partners in metallurgical education to build capacity in this discipline to deliver high quality graduates for the Australian minerals-industry. This paper explores the benefits of industry engagement through a national innovative collaboration, the Metallurgical Education Partnership (MEP), and specifically the final year Process Design Project (PDP), over the four year period from 2008-2011. Some of the benefits discussed in this paper include outcomes of the student experience during an industry-intensive workshop, collaboration between higher education providers and aligning graduate attributes to meet industry requirements.

Sep 1, 2012

By H. Breuer

Mineral processing is characterized by a constant adaptation to changing raw materials and market conditions. It is the link between the mined raw material and a marketable product. As a lot of high grade reserves are exploited, a steady deterioration of raw material quality can be observed. At the same time, the customers´ requirements for product purity and consistent quality increase. Indian iron ores are characteristically friable in nature that results in generation of sizeable quantity of fines (0-10mm) every year during various stages of mining and processing in the country with grade varying from 50-60% Fe(T) and are high in silica/alumina (Al2O3 5-8%). The fines generated cannot be directly used for iron making (blast furnace/DRI units) in the country due its physical and chemical properties; therefore, large portion of it gets exported or dumped. Value addition to iron ore fines through various activities such as beneficiation & palletization is the need of the hour, which will not only be economically beneficial to the mining sector in the long run but also generate value addition & employment opportunities. Incentives may be given for adopting latest technologies for direct use of fines in iron making by agglomeration of fines in the form of Pellets and Sinters, as fines forms considerable part of iron ore resources. Due to selective mining, although the iron content of the ore is high but the alumina(Al2O3) content in iron ore is also significantly high which demands for beneficiation as high alumina will create problem during iron making like higher fuel consumption, poor quality of slag produced, reduced productivity of B.F. all mineral has been engaged in hematite iron ore beneficiation with its gravity and magnetic separators since the mid-nineties. The delivery of jigs started for the upgrading of iron ore for its utilization in a Direct Reduction Plant in Australia. Since then, various other installations with jigs for lumps and fines as well as upstream separators for fines are in operation in Brazil, Australia, India and South Africa. Low grade run of mine and/or dump ores are being processed with alljig®-, allflux®- and gaustec®-WHIMS as the core equipment. The article describes the technology in use, the characteristics of various iron ores and the product qualities achieved. The data presented shows the specific advantages of jig application on iron ore upgrading due to the possible high gravity cuts and the easy and low cost operation. The allflux®- technology provides the big advantage of high capacities and a two-step process, both in a single unit. The gaustec®-wet high intensity magnetic separators offer high capacity in combination with unsurpassed process flexibility. Keywords: iron ore, gravity separation, jig, upstream classifier, wet high intensity magnetic separation (whims), pelletization, sinteringMineral processing is characterized by a constant adaptation to changing raw materials and market conditions. It is the link between the mined raw material and a marketable product. As a lot of high grade reserves are exploited, a steady deterioration of raw material quality can be observed. At the same time, the customers´ requirements for product purity and consistent quality increase. Indian iron ores are characteristically friable in nature that results in generation of sizeable quantity of fines (0-10mm) every year during various stages of mining and processing in the country with grade varying from 50-60% Fe(T) and are high in silica/alumina (Al2O3 5-8%). The fines generated cannot be directly used for iron making (blast furnace/DRI units) in the country due its physical and chemical properties; therefore, large portion of it gets exported or dumped. Value addition to iron ore fines through various activities such as beneficiation & palletization is the need of the hour, which will not only be economically beneficial to the mining sector in the long run but also generate value addition & employment opportunities. Incentives may be given for adopting latest technologies for direct use of fines in iron making by agglomeration of fines in the form of Pellets and Sinters, as fines forms considerable part of iron ore resources. Due to selective mining, although the iron content of the ore is high but the alumina(Al2O3) content in iron ore is also significantly high which demands for beneficiation as high alumina will create problem during iron making like higher fuel consumption, poor quality of slag produced, reduced productivity of B.F. all mineral has been engaged in hematite iron ore beneficiation with its gravity and magnetic separators since the mid-nineties. The delivery of jigs started for the upgrading of iron ore for its utilization in a Direct Reduction Plant in Australia. Since then, various other installations with jigs for lumps and fines as well as upstream separators for fines are in operation in Brazil, Australia, India and South Africa. Low grade run of mine and/or dump ores are being processed with alljig®-, allflux®- and gaustec®-WHIMS as the core equipment. The article describes the technology in use, the characteristics of various iron ores and the product qualities achieved. The data presented shows the specific advantages of jig application on iron ore upgrading due to the possible high gravity cuts and the easy and low cost operation. The allflux®- technology provides the big advantage of high capacities and a two-step process, both in a single unit. The gaustec®-wet high intensity magnetic separators offer high capacity in combination with unsurpassed process flexibility. Keywords: iron ore, gravity separation, jig, upstream classifier, wet high intensity magnetic separation (whims), pelletization, sintering

Sep 1, 2012

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 A. Ekrem Yüce, Gülay Bulut, Feridun Boylu, Vecihi Gürkan, Istanbul Technical University, H. MustafaTarkan

"Nickel sulphide ores subjected to this investigation were taken from Orhaneli district where widespread chromite mineralisation can be occurred. Significant amounts of magnetite and chromite have been existed in the ore together with sulphide and oxide type nickel minerals. Ore sample is constituted about 70% magnetite, 15% sulphide minerals, 5% chromite and iron oxides as well as 10% gangue minerals.According to the chemical analysis results, the representative ore sample contains 1.32% Ni, 10.8% SiO2, 78.4% Fe2O3, 1.3 g/t Ag and 1.0 g/t Au. Mineralogical examinations showed that pentlandite (Fe, Ni)9S8, violarite (Ni, Fe)3S4, millerite NiS and asbolane (Co, Ni)MnO2.(OH)2.1-2H2O were the main nickel minerals and magnetite, chromite, chromium spinals, pyrite, pyrhotite, chalcopyrite, cubanite, goethite, hematite, limonite, rutile, chlorite, talc and serpentine group minerals as associated minerals.The aim of this investigation was to produce valuable nickel concentrates by using gravity separation, magnetic separation and flotation methods. As a results of laboratory scale studies, alternative flow sheets for the production of nickel sulphide concentrate were developed. The related experimental results are given in Table 1."

Jan 1, 2003

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

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

Jan 1, 2014

By Alim Gül, Güven Önal

"Recently, there have been some attemps on the treatment of low grade ores and old plant tailings due to the decrease in high grade ores and the increase in the metal consumption with increasing world population. Cost of mining and beneficiation of low grade ores become prohibitive since the costs of grinding before flotation make the process unfeasable. Therefore, preconcentration of these ores prior to flotation becomes more important. The concentration of ores preconcentrated in coarse sizes provides a decrease in investment costs as well as grinding and flotation costs. Additionally, the requirements of less space for tailing disposal and less reagent consumption are also important from the viewpoint of environmental considerations.In this study, ore sample containing 0.6 % Cu and taken from Küre-Asiköy ore body was used. Ores contained chalcopyrite and pyrite as main minerals and diabaz, quartz, chlorite and sericite as gang minerals.Samples were all crushed under 9 mm by considering the preliminary studies and float and sink experiments were carried out with the sample in the size range of –9+1 mm using different densities of heavy liquids. As a result of these tests, it was shown that 60 % of the material should be discarded as tailings with 0.1 % Cu content. By considering float and sink test results, the following experiments were performed by using heavy medium cyclone and jigs.As a result of heavy medium cyclone tests, which are conducted on the sample in size range of –9+1 mm, 57 % of the feed material was discarded as tailing with 0.08 % Cu content and 7 % Cu loss, while a concentrate containing 1.43 % Cu was produced with 93 % efficiency (Table 1). Jig tests were done with the samples in the size range. As a result of jig tests with –9+1 mm size group sample, approximately 46 % of feed material was discarded as tailings with 0.13 % Cu content and 10.6 % Cu loss, while a concentrate with 0.94 % Cu content was obtained (Table 2)."

Jan 1, 2003

By S. Fiore, M. C. Zanetti, C. Clerici

The processes and fluxes of a cast iron foundry plant located in the North of Italy were analyzed in order to determine the amount and kind of wastes produced from the different activities. The main fractions are the molding sands, furnace muds and powders coming from the dust abatement plant, the exhaust lime, the furnace refractory, the metallic shotting and the sludges and powders coming from the dust abatement plant of molding lines for a total production equal to 800 t/d. All wastes were sampled and particle-size and chemically characterized to evaluate the best recycling or reuse solutions. Wastes were sieved and divided in three categories according to the particle-size dimensions: below 0.1 mm, between 0.1 and 0.6 mm and above 0.6 mm. The fraction above 0.6 mm dimension underwent magnetic separation to recover and quantify the potentially available iron fraction. The fraction between 0.1 mm and 0.6 mm was chemically characterized in order to understand if it could be fed to the sand regeneration plant. The below 0.1 mm fraction was characterized to evaluate the following recycling and reuse solutions: additive for molding sands and/or raw material for the concrete industry.

Jan 1, 2003

By Nirmal S. Weerasekara, Malcolm S. Powell, Mohsen Yahyaei

"An understanding of the fundamentals of different breakage mechanisms is critical for successfully modelling comminution processes. Due to the high frequency of low energy collisions in the grinding process, superficial breakage has a significant role in size reduction in mills. To develop a scaling up methodology for characterising low energy surface damage behaviour of ores, a series of surface damage tests have been conducted using three grinding mill sizes with diameters of 1.8, 1.2, and 0.8 m. Two types of gold-copper ore with different hardness were used: C-ore and T-ore. All tests were conducted using wet mode to remove the products during grinding. To ensure low energy surface damage was the dominant mechanism, in all tests the mill speed was reduced to 40% of critical speed. A surface damage index normalised with product mass and surface based specific energy is determined from the results enabling a study of the grinding performance regardless of mill size, energy level or charge mass. Incremental breakage tests using the JKRBT indicate that the A×b values of the T-ore were 60% higher than C-ore, suggesting the T-ore is considerably softer. The fractional mass loss rate for the T-ore in the 0.8 m and 1.8 m mills started from 16 and 7 (1/kWh/m2) while for the C-ore it was 4 and 2 (1/kWh/m2). After 6 minutes the surface damage index in all tests tends towards steady state a constant value, an indication of steady state surface damage condition. It has been demonstrated that this methodology can be a useful basis for characterising superficial breakage behaviour of ores."

Jan 1, 2014

By A. Üçyildiz

Magnesium borates of varying compositions and properties are synthesized using different reagents and methods for the possible potential uses as termoluminescent material, anti-wear and anti-friction additive, fire retardant, catalyst for the conversion of hydrocarbons, whisker material and ferroelastic material. In the current work, the conditions for synthesizing MgB4O7 compound by high-temperature solid-state reactions were investigated using boric acid and magnesium oxide as the starting materials. The MgO:B2O3 mole ratios applied in the experiments were 1:1, 1:2, 1:3 and the thermal behaviours of each mixture were studied between 25~1200°C by TG/DTA analyses under static air atmosphere to find out the probable formation temperature of MgB4O7 phase. Considering TG/DTA results, freshly prepared boric acid and magnesium oxide solid mixtures were heated at predeter-mined temperatures and the phases were identified by XRD analyses. Accordingly, MgO:B2O3 mole ratio of 1:2, heating temperature of 950°C and a reaction time of 2 hours were selected as the optimum conditions for the formation of maximum amounts of MgB4O7. After obtaining solids prepared at the pre-determined optimum conditions, their chemical composition, IR spectrum, TG/DTA pattern, B.E.T. surface area value, particle size distribution and SEM images were deter-mined to point out the possible use areas. Experimental results showed that it was not possible to obtain pure product at the optimum conditions due to formation of minor amounts of t-Mg2B2O5 and the presence of some unreacted B2O3. The product synthesized has the potential for use as a corrosion inhibitor on some metal surfaces.

Jan 1, 2014

By Victor V. Karmazin

"The appearance of the precise equipment for analysis of results of the laboratory tests with using of the computer processing allows raising accuracy of the estimation of the expected indexes of the mineral processing. Certainly, laboratory tests will not be able to substitute completely industrial ones, but allow reducing the amounts of the industrial tests, that vastly reduces the cost of technological forecast and preparation of data for projection.Calculation of the potential indexes of mineral technology (mass yields and the metal content in products) is fulfilled on the base of results of measurements of the mineralogical liberation parameters using special formulas.The quality of the concentrate ßc = a + ?l • ?s • ?i (ßt - a) can be achieved on a level of the theoretical metal content of a mineral - ßt only if coefficients of mineral processing, of liberation ?l and selectivity ?s of grinding, effectiveness of separation, will reach 1 (a - metal content of the row ore).Increase ?l by overgrinding automatically decrease ?s, and only by destroying of particles of ore along surfaces of accreting of minerals (?i = 1) and when values of all three coefficients will aim to 1, then ßc ? ßt.The degree of liberation of the useful mineral is the ratio of mass of free grains of the useful mineral in the product to general mass of the same mineral in this product. The factor of the liberation, which is defined by curves of dressability by angle of slop of the curve of the elementary faction - ?, is Kr=1 – tga. Kr changes from 0 to 1 when a changes from 45o to 0o. The liberation is reached by grinding of the material, although degree of the liberation, as a rule, is lower than the degree of the grinding."

Jan 1, 2003

By A. Güney

Currently, recycling technologies use relatively homogeneous plastics because hand-sorting waste feed is too expensive. Many promising technologies for separating mixed thermoplastics are being investigated, but they are still uneconomical and unreliable. Burning plastics for recycling is not used because of poisonous gases released when their burning. PET and PVC cannot be separated using conventional gravity separation techniques because their specific gravities overlap; PET density changes from 1.33 g/cc to 1.37 g/cc and PVC density is between 1.32 g/cc and 1.37 g/cc. For this reason, plastic recycling needs new techniques. Among these techniques froth flotation, also used in mineral processing, can be useful for its cost and simplicity. In this method, a plastic is selected from mixture, tried to regain its natural property, hydrophobicity. In this study the experiments were applied by using plasticizer reagents to make the selected plastics hydrophobic and us-ing NaOH to render the other hydrophilic. Post-consumed PET and PVC are evaluated by concerning the effects of selected parameters (particle size, concentration of plasticizer reagent, amount of NaOH, conditioning time and the effect of conditioning temperature). At the end of the experiments, optimum parameters for selective flotation of PET from PVC was found as particle size of -4.76 + 3.36 mm, NaOH amount of 30 kg/t, conditioning temperature of 75°C, conditioning time of 30 minutes, reagent amount of 300 mg/L. Consequently, PET was re-covered with 99.7% purity and 57% efficiency.

Jan 1, 2014

By Kobus (J. J. ) Rossouw, P. Chris Pistorius, Johannes Sono

"Different coal properties are important in SL/RN (coal-based rotary kiln) direct reduction of iron ore, and in electrical smelting of prereduced iron ore. Laboratory simulation was used to evaluate two candidate coals, which were similar in volatile content and rank, but differed in reactive maceral content. The tests served to evaluate both the reduction rate (during direct reduction) and the dissolution rate of carbon from the coal into an iron bath.Reduction rates of mixtures of magnetite and devolatilised coal were assessed isothermally, using samples of 55 grams. The off-gas analysis was used to follow the extent of reduction. Results are shown in Figure 1 a). Coal B was expected to be more reactive, based on its higher content of reactive macerals (77%, compared with 61% for Coal A). In line with this, Figure 1 a) does show a higher reduction rate with this coal. However, increasing the reduction temperature increases the reduction rate substantially,especially for the less-reactive Coal A.For the dissolution experiments, samples (mass: 250 g) were melted in an alumina crucible (under a nitrogen atmosphere), to give a starting composition around 1%C. When the desired temperature was reached, 25 g of char was added on top of the melt, through a feeding port. Results are shown in Figure 1 b). The more-reactive Coal B was less crystalline (the graphite crystallite size estimated from X-ray diffraction was 4.6Å, compared with 11.4Å for Coal A). This is associated with a lower dissolution rate in iron."

Jan 1, 2003

By G. H. Li

Preparation strontium carbonate with celestite concentrate by flotation is studied by agglomeration-reduction roasting process. It is shown that, the drop strength of green pellet is greater than 4 times?(0.5m) -1 under the balling conditions of moisture of 9% and time of 18~20min. Lignin xanthate obviously improves the ballability of the celestite concentrate due to the increase of surface hydrophilicity, therefore, the balling rate is increased, as well as green ball strength. The drop strength of 7 times?(0.5m) -1 can be obtained when the lignin xanthate dosage is 0.5%~1%. Reduction roasting with anthracite shows that the SrSO4 conversion is 82%~83% for celestite concentrate pellet, under the optimal conditions of roasting temperature of 1100~1150? and time of 30min, and it is 10% higher than conversion of celestite lump under the same conditions. The strontium carbonate are prepared after roasted pellets are leached, purified and carbonated, and the SrCO3 content is higher than 97% , while the content of CaCO3 is less than 0.3%.

Jan 1, 2014

By K. Saberyan

Zirconium and hafnium are two important elements in nuclear industries. Zirconium is used in fuel rod for its low neutron absorption. Hafnium should be separated from zirconium, because it has higher neutron absorption cross section. In this research, separation process of the mentioned elements from (Zr+Hf)Cl4 mixture produced in the Zirconium production plant (ZPP) of Iran (Esfahan) was studied employing solvent extraction method in the nitric acid medium by using tributylphosphate (TBP) as solvent, and kerosene as diluent. The effects of nitric acid, sodium nitrate and amount of extractant, zirconium concentration in aqueous phase, phase ratio, agitation time, and mixer agitation rate were investigated on the separation process. The process was optimized at 5 M nitric acid, 1 M sodium nitrate, 60% (v/v) TBP. The optimum values for phase ratio, agitation time and agitation rate were 1.5 min and 500 rpm respectively. In this condition, the zirconium extraction efficiency, and separation factor were 65.1% and 205.59, respectively. Finally, we accessed to nuclear grade in this research.

Jan 1, 2014

By I. A. Makalin, V. N. Yakovlev, S. V. Imangulov, I. V. Zyryanov

"The presented studies aimed at solving the problem of recovery of low-luminescent diamonds, which are not recovered using X-ray luminescence method of processing of diamond-bearing ore. The purpose of the studies is the determination of possibility and feasibility of recovery of such crystals using dual-energy X-ray transmission separation technology. In preparation for studies methods were developed for testing technical and process parameters of XRT-400 Aliud GmbH (Germany) sorter, manufactured according to design specifications of Yakutniproalmaz Institute, and XRT/D TOMRA Sorting Solutions GmbH (Germany) sorter, designed for processing of -50+5 mm ore, as well as RGS-1M, RGS-2 X-ray sorters, developed by JSC NPP Bourevestnik (St. Petersburg) in cooperation with Yakutniproalmaz Institute, for material size of -6+3 mm. In 2018, in-process tests of RGS-OD-3 X-ray sorter prototype on material size of -3+1 mm are scheduled. Studies to determine the content of low-luminescent diamonds and feasibility of their recovery using XRT-400 and XRT/D sorters at the Processing Plant #3 of Mirny Mining and Processing Division (MPD) were carried out. The results of the studies confirmed presence of such diamonds and possibility of their recovery using the dual-energy X-ray transmission method. In-process tests of RGS-2 X-ray graphic sorter were carried out at the Processing Plant #16 of Nyurba MPD. According to the results of sampling of RGS-2 sorter separation products, the recovery of -6+3 mm sized diamonds from tailings of X-ray luminescence sorters was achieved at the level of 96%, while the diamonds recovery was 98.5% in value terms, the concentrate yield was 0.1%. In-process tests of RGS-1M sorters at final recovery of the grease separation concentrate at the Processing Plant #3 of Mirny Division were carried out. Recovery of low-luminescent -6+2 mm sized diamonds by RGS-1M sorter amounted to 92.4% for single operation, and with the double control operation reached 99%. As a result, -6+2 mm sized base product recovery increased by 0.2%."

Jan 1, 2018

By Alan Bax, John D. Lewins

"Platinum Australia Limited (“PLA”) is currently undertaking a Feasibility Study into the development of the Panton Platinum Palladium Project located in the Kimberley region of Western Australia. The Panton Intrusive is a layered, differentiated mafic to ultramafic structurally complex body that has undergone numerous folding and faulting events. This has resulted in a southwesterly plunging synclinal structure some 10 kilometres in length and 3 kilometres wide.High grade PGM mineralisation is hosted within a number of stratiform chromitite reefs within the ultramafic sequence. The focus of the Feasibility Study is on the Top Reef which is situated approximately 150 metres below the mafic/ultramafic interface. The average true width of this reef is approximately 1.2 metres and the reef has a defined resource is 10.1 million tonnes at an average grade of 6.1 g/t PGM + Au.The Palladium to Platinum ratio of the Top Reef is approximately 1.15:1 and the reef also contains an average of 0.4 g/t Au and 0.3 g/t other PGM’s. The Panton ore is considered refractory in nature and standard flotation techniques achieve a PGM + Au recovery of less than 65% with a concentrate grade of less than 100 g/t PGM + Au. As part of the Feasibility Study Platinum Australia in conjunction with their major shareholder Lonmin Plc developed a new metallurgical process which has achieved an overall PGM + Au recovery of approximately 80% into a high grade PGM concentrate which is suitable for feeding direct to a PGM refinery.THE PANTON PROCESSThe Panton Process incorporates a two-stage Milling and Flotation circuit with no cleaning stages to produce a low grade, high recovery flotation concentrate. Typically the concentrate mass constitutes 25% of the feed and has a PGM + Au grade of 25 g/t.This low-grade concentrate is subjected to calcination in a rotary kiln or fluid bed at a temperature of approximately 400 0C for a period of 60 minutes. This step is designed to render the platinum minerals soluble in cyanide and both the temperature and the duration of calcination are critical to the efficiency of the leaching process."

Jan 1, 2003

By Gus Van Weert, Samuel A. Bolorunduro, David B. Dreisinger

"INTRODUCTIONAcid pressure oxidation of high-grade natural acanthite was investigated in the range of 110-150°C and in the sulphate media. High acidity and low pulp density were used to prevent the precipitation of silver.RESULTS AND DISCUSSIONTwo sets of acanthite specimens were obtained from La Valenciana and Ingleterra mines, Mexico. Pressure oxidation tests were conducted on 1 litre scale in a standard 2-litre Parr titanium autoclave.EFFECT OF TEMPERATURE ON THE REACTION OF ACANTHITEThe effect of temperature on the rate of oxidation is presented in Figure 1. The rate of reaction increased with increasing temperature."

Jan 1, 2003

By Özcan Y. Gülsoy

Coal beneficiation with optical sorting has not yet been studied in detail. Optical separation of coal from gangue particles as a dry method can be used as an alternative to the other enrichment methods. Developments in optical and pneumatic technology put the optical sorting one step forward by means of applicability of the mechanical sorting to the mineral industry. Because of the inefficient separation performance, dry separation of coal is very limited in industrial applications. Therefore, optical sorting depending on color differences between two minerals can be introduced to coal beneficiation systems. In this study, coal samples obtained from Konya-Ilgin region (Central Anatolia-Turkey) were separated by optical sorter and the results were evaluated. The conventional floatability tests for Ilgin coal are conducted by collecting most of the feed material at 1.3-1.4 g/cm3 density range. Separation of Ilgin coals with heavy medium methods will not be possible because of poor density difference between coal and its impurities. There are important color differences among the floated particles at 1.3-1.4 g/cm3 density range. Therefore, Ilgin coal is subjected to optical sorting in a pilot scale optical sorter. Different size fractions were fed to the optical sorter and then rejected and accepted particles by equipment were analyzed. Ash content of the accepted particles was reduced to 25-35 % and ash in tail increases up to 55-60 % after sorting. Although most of the material floated at low densities, optical sorter separated the feed into two different quality products. This study has the potential to clarify applicability of optical sorting even to non-floatable coals and to provide an important contribution to both industrial applications and literature.

Sep 1, 2012

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