Search Documents

By Felix M. Zhuravlev, Vitalij P. Lyalyuk

"While there is a large industry based on recovering platinum group metals (PGM) from complex sulphide ores containing Cu and Ni, there is little information on the thermodynamics of the matte phase formed on smelting the ore and in particular the behaviour of Pt in the matte.The equilibrium between metallic platinum and industrial mattes was studied in a series of experiments where a small quantity of matte (5g) and platinum foil of 150 µm thickness (0.5 g) were equilibrated in magnesia crucibles at 1400°C under conditions where the oxygen and sulphur partial pressures were controlled.The chemistries of the phases were determined from EDS analysis using a scanning electron microscope. Line scans were initially used to confirm that the foils had attained a uniform composition across their thickness. The composition of the matte was determined from analysis of an area of matte including the small metallic droplets and dendrites.The distribution of Pt in Cu-Fe-Ni mattes (Pt: 1.5 – 3%) and solid Pt alloys (Pt: 54 – 72%) was studied. For a matte of base composition 45%Fe, 11%Ni, 8% Cu, at a base temperature 1400°C, the following behaviour was observed for the Pt distribution ratio:Under the range of sulphur and oxygen pressure of this study (log pO2 from -15 to –12.2, log pS2 from – 4.6 to –2.2), the distribution ratio of Pt between matte and alloy was independent of the sulphur and oxygen pressure, having a value of 0.04.The Pt distribution ratio increased from 0.02 to 0.08 with temperature from 1340 to 1450 oC.For mattes where the proportions of Cu and Ni were kept constant, a slight decreasing trend was shown for the distribution ratio of Pt versus the iron content (37 to 53%) in the matte but was not confirmed because of experimental uncertainty."

Jan 1, 2003

By Steven Wright, Sharif Jahanshahi, Shouyi Sun

"While there is a large industry based on recovering platinum group metals (PGM) from complex sulphide ores containing Cu and Ni, there is little information on the thermodynamics of the matte phase formed on smelting the ore and in particular the behaviour of Pt in the matte.The equilibrium between metallic platinum and industrial mattes was studied in a series of experiments where a small quantity of matte (5g) and platinum foil of 150 µm thickness (0.5 g) were equilibrated in magnesia crucibles at 1400°C under conditions where the oxygen and sulphur partial pressures were controlled.The chemistries of the phases were determined from EDS analysis using a scanning electron microscope. Line scans were initially used to confirm that the foils had attained a uniform composition across their thickness. The composition of the matte was determined from analysis of an area of matte including the small metallic droplets and dendrites.The distribution of Pt in Cu-Fe-Ni mattes (Pt: 1.5 – 3%) and solid Pt alloys (Pt: 54 – 72%) was studied. For a matte of base composition 45%Fe, 11%Ni, 8% Cu, at a base temperature 1400°C, the following behaviour was observed for the Pt distribution ratio:Under the range of sulphur and oxygen pressure of this study (log pO2 from -15 to –12.2, log pS2 from – 4.6 to –2.2), the distribution ratio of Pt between matte and alloy was independent of the sulphur and oxygen pressure, having a value of 0.04.The Pt distribution ratio increased from 0.02 to 0.08 with temperature from 1340 to 1450 oC.For mattes where the proportions of Cu and Ni were kept constant, a slight decreasing trend was shown for the distribution ratio of Pt versus the iron content (37 to 53%) in the matte but was not confirmed because of experimental uncertainty."

Jan 1, 2003

By Herman D. Mendoza

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

Jan 1, 2003

By Jan Svoboda, Atsushi Shibayama, Toyohisa Fujita, Toshio Miyazaki, Kenji Murata

"The electrical disintegration (ED) method and disintegration by underwater explosion were investigated as new fragmentation techniques for the kimberlite ore and tungsten scrap of superalloy (carbide bit). The results of various breakage tests demonstrated that both fragmentation methods as new disintegration techniques for recycling were able to liberate various components of materials of different resistivity and elastic impedance.SUMMARYThe breakage of the kimberlite ore from South Africa (sample size: 15 mm x 15 mm x 6 mm) was achieved by the ED method. A relationship between the sample size, voltage, the electrical current and the energy needed for the disintegration was found from the experimental results. Kimberlite exhibited a high degree of breakage, when the high voltage of 60 kV and the pulse current of approximately 5000 A were applied to the sample by the generator during 0.2 microseconds (Fig. 1)."

Jan 1, 2003

By Hiroshi Sasaki, Herman D. Mendoza

"The surface-chemical property of nickel slag fiber was modified in aqueous media in order to act as an active and efficient collector of colloidal amorphous silica from geothermal effluent. Colloidal amorphous silica (0.05 µm) from geothermal wastewater causes operational and ecological problems. It clogs geothermal pipes and waterways, and when disposed to the environment poses hazardous effects on both the flora and fauna ecology. Because colloidal amorphous silica is extremely fine and remains in suspension in aqueous media, conventional solid/liquid processes to remove colloidal silica from its bulk liquid are rendered ineffective.A rapid separation technique is then developed to remove effectively amorphous colloidal silica from its bulk liquid by attaching the extremely fine particles of silica onto a nickel slag fiber coated with magnesium hydroxide, Mg(OH)2. The nickel slag fiber has the following characteristics: a high surface area and predominantly possesses negative charge in aqueous media with an iso-electric point at pH ~2.5. When soaked and saturated with Mg(OH)2 solution, the surface of the nickel slag fiber becomes coated with fine and compacted Mg(OH)2 precipitates. This modifies the surface- chemical property of nickel slag fiber transforming the negatively charged surface into a positively charged surface in aqueous media with an iso-electric point of around pH 11.8, which corresponds to the iso-electric point of Mg(OH)2. When geothermal effluent containing negatively charged colloidal amorphous silica particles (i.e.p. of silica is pH 2) is passed through a column-bed packed with surface-chemically modified nickel slag fibers under natural pH conditions, the colloidal amorphous silica particles are recovered by attachment onto the surface of the surface- modified nickel slag fiber. The recovery of colloidal amorphous silica by attachment is 97-99%, resulting in the production of a relatively clean and silica free output. Aluminum hydroxide, Al(OH)3, was also used to modify the surface –chemical property of the nickel slag fiber. In the same way, it also exhibited a positive response with the collection of colloidal amorphous silica from its bulk liquid, but yielding a lower recovery of 90%. This may be attributed to the lower surface potential and iso-electric point of pH 9.3 for Al(OH)3. The separation process was evaluated and analysed using existing potential energy theories. Figure 1 shows a micrograph of the surface of an unmodified and the Al(OH)3 –modified surface of nickel slag fiber."

Jan 1, 2003

By T. Mütze

The progress in comminution is often described by the specific surface area and the energy utilization or the brakeage ratio and breakage function. The principal mechanisms of size reduction are therefore, well known. All four parameters share a common disadvantage: They are limited in describing the behavior of a feed material with fine particles and a wide particle size distribution. An alternative is to describe the progress in comminution with ? The particle size distributions Q(x) before and after the stress or ? The particle concentration ratio pcr(x), which combines these two states. In the case of a self-similar fracture behavior, (in short: self-similarity) the progress in comminution can be simplistically described by the relationship between a reference particle size and the stress intensity. Self-similarity means self-similar breakage functions of geometrically similar particles of the same strength, a certain stress situation and stresses up to the breakage point. Self-similar breakage functions show the same curve at different stress intensities when the fragment size x is normalized with a reference particle size xNorm. This behavior was demonstrated for single particle comminution repeatedly and unequivocally. In the field of particle bed comminution, these results differ in literature. Therefore, the self-similarity in particle bed comminution was systematically investigated at the Institute of Mechanical Process Engineering and Mineral Processing of the TU Bergakademie Freiberg as part of a research project. Limestone, quartz, silicon carbide and glass beads, were examined with a hydraulic press. The particle sizes covered monodisperse fractions in the range of 1-1000 µm, bimodal mixtures of these fractions and polydisperse fractions. The results obviously show that there is no self-similar fracture behavior in the case of glass beads, quartz and silicon carbide. Statements on the self-similarity of limestone are much more difficult: Only some of the investigated limestone fractions show a similar clear, non-self-similar behavior. Self-similar curves appear particularly for fine monodisperse fractions as well as bimodal mixtures, which stressed with pressures above 200 MPa. Considering the measurement accuracy of particle size analysis the coincidence of the curves can be explained by the fact that progress in comminution almost ends above 200 MPa. At this point, a hydrostatic stress state is reached in the particle bed which is known from the literature. This stress state inhibits further fractures. Overall, it has been verified for the investigated materials that there?s no self-similar fracture behavior in particle bed comminution. One reason for the diverse opinions in literature is that there are three types of particles after the comminution in the particle bed: ? Particles which are stressed below their breaking point, ? Particles which are stressed right up to the break point, and ? Particles which are stressed beyond their breaking point.

Sep 1, 2012

By C. Dorfling, P. Albertyn

"Hydrometallurgical recovery of metals from printed circuit board waste typically involves multiple leaching stages; conventionally, base metals dissolution by mineral acids is followed by leaching of gold using cyanide or halide based lixiviants. Alternative lixiviants for gold leaching can reduce the potential hazardous nature of operation. The aim of this study was to investigate the feasibility of using thiosulfate as lixiviant for gold leaching from waste printed circuit boards. Batch leaching tests were performed to investigate the effects of thiosulfate concentration (0.1 – 0.2 M), ammonium concentration (0.2 – 0.4 M), pH (9 – 10.5), solids content (25 – 50 g/L), and copper content of the feed material on the gold leaching behavior. The concentrations of metals in liquid samples were determined by means of inductively coupled plasma atomic emission spectroscopy, and high pressure liquid chromatography was used to evaluate thiosulfate degradation/consumption. The ratio of thiosulfate to ammonia was found to have the most significant effect on the extent of gold leaching, with the best gold extraction observed at a ratio of 0.5 mole thiosulfate per mole ammonia. The lower pH range of 9 – 9.5 generally produced better Au leaching, while low copper extraction during the first stage base metal leaching stages has proven to be detrimental to gold leaching from the residue. Thiosulfate degradation/consumption increased and gold extraction decreased with higher residual copper in the feed. An optimum gold extraction of 78.5 % from base metal leach residue was achieved with 0.2 M thiosulfate, 0.4 M ammonia, and 0.02 M cupric ions at 25°C in the pH range 9 – 9.5.INTRODUCTION Background and Aim The recycling of e-waste is becoming increasingly important as it provides certain economic benefits apart from the obvious environmental benefits. Printed circuit boards (PCBs) are found in most forms of e-waste and contain especially high concentrations of base and precious metals. Hydrometallurgical processes are often considered for the recovery of valuable metals from PCBs; these processes normally implement several leaching stages to selectively recover certain metals."

Jan 1, 2018

By S. F. Kaplan, A. S. Dolotov, V. N. Kovalev

"At this point in time, autoclave recovery (POX) followed by sorption cyanidation (CIL) is one of the most promising technologies for the efficient processing of refractory and double refractory gold-bearing sulphide rock. This hydrometallurgical technology requires the use of a large amount of water (up to 5 tons per ton of used concentrate), resulting in highly saline water with high pH (pH 9-11), high content of cyanides (up to 300 mg/l), arsenic compounds, antimony and a number of heavy metals. In order to determine the possibility of using recycling industrial water in POX-CIL technology, several batches of experiments were conducted (using the double refractory flotation concentrate of the Mayskoye deposit). Urban water with a chloride ion content of 10 mg/l was used during the reference POX-CIL experiment as the liquid phase of the ??? feed pulp. At that, gold extraction rate amounted to 89.4 %. A full cycle of POX(200 ºC)-CIL was carried out during the first part of the experiments; at that, fresh concentrate of the untreated recycling water with cyanide ion content of 95.4 mg/l, left after the completion of the previous CIL cycle, was used for repulping. It has been established that direct use of untreated recycling industrial water is impossible due to a sharp decline in gold extraction (preg-robbing) down to 79.5%. The presence of chloride ions, cyanide ions and other ligands in the liquid phase provides effective complexion formation for the recovery gold; complex gold compounds are quickly and firmly sorbed by the natural carbonaceous substances of the concentrate at the stage of autoclave recovery; they are virtually not cyanized onwards, and gold is carried over to the tailings. During the the second batch of experiments, recycling industrial water with cyanide and chloride ions' concentration of 500 mg/l each was subjected to the reverse osmosis desalination (up to 6 times in a successive order). After each desalting cycle, such water was used as the liquid phase of the feed pulp in POX-CIL. 6 cycles of RO desalination enabled the extraction of 86.1% Au (which is close to reference recovery). During the third batch of experiments, recycling industrial water was diluted with distilled water (2, 4, 16, 32, 64 times); after each dilution cycle, this water was used as the liquid phase of the POX-CIL feed pulp. Gold extraction after 64-fold dilution amounted to 82.6%. The results of the conducted researches show that the use of untreated recycling industrial water in POX-CIL hydrometallurgical technology is absolutely impossible due to a sharp decline in gold extraction. Even the 64-fold dilution with pure water is inefficient both from the perspective of gold extraction and of rational use of water. Reverse osmosis desalination solves the problem of water treatment except for the removal of cyanide ion, which requires separate reagent treatment of water."

Jan 1, 2018

By Zoran M. Stevic, Mirjana M. Rajcic-Vujasinovic, Zvonimir D. Stankovic

"Covellite is one of the most important copper sulfide minerals. Its behavior during the anodic polarization, as well as its rest potential in various solutions are of importance for flotation processes and also define the mechanism of electrochemical extraction of copper from sulfide minerals.Measurements of the rest potential have been performed in water solutions in presence of different cations and anions. The existence of the Nernst type dependence of the rest potential on the activity of some of the investigated ions was established. The biggest influence on the rest potential of covellite exhibits the change of activities of the ions forming the crystal lattice of investigated mineral, i.e. Cu(II) and S(II). The dependence of the rest potential of covellite on logarithm of Cu(II) activity has a shape of a strait line with a positive slope of 60 mV, while when concentration of S(II) anions increases ten times, the rest potential of covellite decreases about 120 mV. The first dependence was established in acid medium, while the second was registered in alkali one.Different electrochemical techniques have been used for the investigation of the behavior of covellite during anodic polarization. In the basis of all these measurements is a three electrode system consisted of: 1) working electrode made of the sample of natural covellite from the Bor mine (Eastern Serbia, Yugoslavia); 2) reference electrode (saturated calomel electrode) and 3) counter electrode (flat platinum). Electrolyte solutions used in these investigations are of the same kind as those used in the rest potential measurements. Sulfate and chloride solutions, as well as the mixed sulfate/chloride solutions, are the most perspective from the application point of view.Cyclic voltammetry has been the first used method that confirmed the formation of passive layer on the electrode surface during anodic polarization. Elementary data about the reaction mechanism have been obtained too. More detailed data needed for more precise conclusions about the mechanism and rate of the anodic process on covellite have been obtained by galvanostatic method. It was established that the essential difference exists between the reaction mechanism in sulfate and chloride solutions. Also, the difference is exhibited in the nature of reaction product directly connected with the passivation of the electrode surface. Chronopotentiometric investigations have been performed in sulfate, nitrate, chloride perchlorate and mixed sulfate/chloride solutions applying different current densities. The shape of obtained curves indicate if there was a surface passivation and how intensive it was."

Jan 1, 2003

By V. Yu. Ivanov, A. I. Yuryev, L. S. Lesnikova, M. S. Datsiev

"The Talnakh Concentrator was put into operation in 1981 using a direct selective process to produce copper, nickel and pyrrhotite concentrates.In 2001, a novel selective-bulk-selective (SBS) technology for the concentration of massive sulfide ore was implemented, that allowed to increase the valuable components recovery into target concentrates with feedstock grinding coarser than before. In 2003-2005, the process upgrading included the differentiated processing of nickelpyrrhotite products recovered at the different concentration stages and made it possible to significantly enhance the contrast of the pentlandite and pyrrhotite flotation properties and to simultaneously increase separation of polymineral intergrowths.The next stage of the PD concentration complex development is the Project ""Talnakh Concentrator Reconstruction and Technical Upgrading"" to raise its capacity by half. The project has three Phases.• Phase 1. Pilot startup was carried out in January 2015 following the installation of new OK-100 TC flotation machines, manufactured by Outotec.• Phase 2. The introduction of a new technology developed by the Polar Division specialists launched in mid-2016. The new technology is based on maximization of iron and sulfur removal into tailings and maintenance of non-ferrous and precious metals losses at the same level.After the realization of Phase 3, the annual capacity of the Talnakh Concentrator will raise by another ~ 40% at the expense of processing cuprous and disseminated ores currently being processed by the Norilsk Concentrator.By 2023, due to upgrading of the PD production facilities the emissions to the atmosphere will reduce by 75%.Polar Division of PJSC MMC Norilsk Nickel (PD) searches for the ways to solve its environmental issues. Thus, in 2016, the Nickel Plant was shutdown and a new flowsheet was introduced at Talnakh Concentrator. It resulted in the reduction of sulfur specific amount sent to pyrometallurgical processing as a component of the concentrates.The Talnakh Concentrator was put into operation in 1981 using a direct selective process to produce copper, nickel and pyrrhotite concentrates.In 1981, its design capacity for processing of Cu-Ni ores from Oktyabrsky and Taimyrsky Mines did not exceed 3 million tons per year.In 2001, a novel selective-bulk-selective (SBS) technology for the concentration of massive sulfide ore was implemented at the Talnakh Concentrator, that allowed to increase the valuable components recovery into target concentrates with feedstock grinding coarser than before."

Jan 1, 2018

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

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

Jan 1, 2003

By G. M. Wolfaardt, B. C. Qi

"From a lifecycle point of view, biologically digested sludge is a valuable product that should be reused if possible. In contrast to landfill and incineration, application of digested sludge comprises the recycling nutrients and organic matter and reflects the policy of resource conservation and sustainability in waste management. Sludge may be applied to agricultural land for further sludge treatment and improvement of the characteristics of the soil. The characteristics of sludge that affect its applicability for soil conditioning include moisture content, pathogens, nutrients and metals. In this paper, the use of solid anaerobic digestion sludge of the lignocellulosic residues to condition soil was investigated in regards to nutrient composition, water content/retention capacity, hygienic quality (contamination of pathogenic bacteria) and heavy metal adsorbing/releasing characteristics of the sludge.The digested sludge had approximately 44.5% free and immobilized water and 46.9% bound water. The water-holding capacity of the sludge was high. This was the first merit to apply the sludge to agriculture in dry areas. Secondly, the sludge was rich in macro-nutrients (N, P, K) and a series of micro-nutrients (Ca, Mg, Mn, Na, Fe, Cu, Zn, B). The water-soluble fractions of macro-nutrients in the sludge were found considerably low. This implied that the loss or leaching of the macro-nutrients from the anaerobic sludge to the aqueous environment be slow. The nutrients loss owing to leaching could be reduced. However, because of a lower macro-nutrient content of the anaerobic sludge, comparable to the chemical fertilizer, it is advisable that the anaerobic sludge be applied together with commercial fertilizer to the agriculture land. Thirdly, the anaerobic sludge had strong selective adsorption affinity for heavy metal ions; non-heavy metal ions such as K(I) and Ca(II) could be released easily from the sludge in a wide pH range, such as pH from 1.0 to 8.0. Since the metal toxicity to the environment mainly comes from the mobile heavy metal ions, the selectivity to adsorb heavy metal ions of the anaerobic sludge can be utilized to reduce the availability of heavy metal ions to plants and humans in aquatic environment. Fourthly, the anaerobic digestion under thermophilic conditions showed a promising disinfecting effect to the sludge, in contrast to the landfill digestion."

Jan 1, 2003

By Claudio Bisso, Javier Vargas, Charles Lozano

"This paper summarizes the activities performed to predict SAG milling throughput at Minera Esperanza. Having highly reliable geometallurgical models aided by the new computer tools and specific metallurgical testing to rely on has become a major issue in recent times due to their impact on the business. Minera Esperanza’s SAG throughput model was constructed by characterizing the ore feed and operational parameters of the plant. Producing a robust model required satisfying number of requirements for the parameters used for both the ore and the plant as follow: 1) geologic and geometallurgical characterization parameters; and 2) parameters of a plant running at the rated production under optimum operating conditions. Obtaining these parameters required increasing the quantity of samples and metallurgical testing through drilling campaigns, mapping the drill holes and putting samples through various comminution tests while sampling all operational flows at the plant to obtain the metallurgical information for the operational parameters. The information was entered on the JKSimMet software to enable a throughput model to be developed to output SAG mill throughput predictions that are very close to actual throughput figures, i.e., accurate within a range of 2% on a monthly basis."

Jan 1, 2014

By Yi-jun Cao, Jin-cai Ran, Guo-sheng Li, Jiong-tian Liu, Li-jun Deng, Shu-lei Li

A fly ash sample from a power station in Guangdong province has been characterized by size separation, X-ray diffraction, contact angle measurements and X-ray Fluorescence. The ash sample had a loss-on-ignition (LOI) (%) of 15.54% and a large proportion of the unburned carbon was distributed in the +74 µm and -20 µm fractions. Contact angle measurements suggested that the removal of unburned carbon from the fly ash using flotation would present some challenges. The effects of various operating variables on the removal rate of unburned carbon (RUC) (%) and (LOI) (%) were determined using a cyclonic-static micro bubble flotation column (FCSMC). Within the range studied, the optimum diesel oil dosage was 1200 g/t, abies oil (frother) dosage was 600 g/t, pulp density was 20%, superficial gas velocity was 1.4 cm/s and circulating pressure was 0.20Mpa. Column flotation tests indicated that the FCSMC method a viable and effective separation process for removing the unburned carbon from the fly ash. Under the optimized conditions, a clean ash with 2.13% LOI and 94.21% RUC was obtained.

Jan 1, 2014

By Ljubisa Obradovic, Baisui Han, Batnasan Altansukh, Zoran Stevanovic, Atsushi Shibayama, Jonovic Radojka, Daizo Ishiyama, Kazutoshi Hag, Radmila Markovic, Ljiljana Avramovic

This paper discussed a combined froth flotation and high pressure leaching (HPL) process to recover copper from mine tailings. At first, froth flotation were carried out using the mine tailings sample under various conditions in the presence of different pH (3.5-10) and sulfurizing reagent value (0-1000 g/t), respectively. As a result, optimum condition (sulfurizing reagent of 1000 g/t, pH of 10, PAX (collector) of 100 g/t and MIBC (frother) of 200 g/t) was found for froth flotation of copper. The effects of sulphuric acid concentration, leaching temperature, total pressure (steam pressure plus oxygen pressure) on high pressure leaching (HPL) of mine tailing were investigated using sulphuric acid solution in the presence of oxygen. The best conditions for HPL were obtained to be sulphuric acid concentration: 0 mol/L (water), leaching temperature: 170 ?, total pressure: 2 MPa, leaching time: 1 hr and pulp density: 100 g/L. The leaching rate of copper reached nearly 99 % and the concentration of copper in pregnant leach solution (PLS) was 0.24 g/L. Finally, copper in the mine tailing was recovered using froth flotation followed by HPL. As a result of a combined froth flotation and HPL, the vast majority of copper (99 %) from the mine tailings was extracted.

Jan 1, 2014

By S. K. Kawatra, T. C. Eisele

"Coal-fired electrical power generation plants produce 56% of all the electricity in the U. S., burning approximately 860 million tons of coal annually. This results in production of 58 million tons of fly ash; 19 million tons of bottom ash and boiler slag, and 23 million tons of desulfurization sludges. These are a considerable disposal problem, and it would therefore be very valuable to be able to sell these residues and byproducts as useful materials, rather than paying to dispose of them in landfills. However, utilization rates are currently very low, below 32% even for the most thoroughly commercialized combustion byproduct (fly-ash). This poor utilization rate is a result of the high impurity content of combustion byproducts, which make them unsuitable for existing markets. The experiments described in this paper were therefore carried out to study the application of mineral processing techniques to upgrading combustion byproducts.Experiments were conducted to upgrade and purify low-grade combustion products to produce highgrade, readily saleable materials. A process was developed for removing unburned carbon from coal fly-ash by froth flotation, using reagents originally developed for flotation of oxidized coals. This technique could reduce the carbon content of fly-ashes containing up to 11.1% carbon to less than 2%. The froth product was sufficiently carbon-rich that the recovered carbon has value as fuel, or as low-grade activated charcoal.Fly ash was also examined as a binder for producing iron ore pellets, primarily as a substitute for bentonite. It was found that Class F fly ash in combination with an activator (calcium hydroxide) was an effective pellet binder at dosages comparable to the currently-used dosages of bentonite clay. While the flyash worked very well alone, tests in combination with bentonite showed that fly-ash/bentonite mixtures had significantly worse binding performance than either pure bentonite or pure fly-ash-based binder. This was determined to be due to the fact that bentonite and fly ash have profoundly different binding mechanisms, with fly-ash undergoing a pozzolanic chemical reaction while bentonite binds through a physical attachment process."

Jan 1, 2003

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

"INTRODUCTIONThe objective of this work was to deport silver into a residue suitable for subsequent extraction without compromising the high base metal extraction during the pressure oxidation.RESULTS AND DISCUSSIONThe complex sulphide from the tailing pond of Cannington mine, Australia, was re-grounded and a bulk concentrate was recovered by flotation. The major economic sulphides are galena and sphalerite. The chemical analyses revealed 17.22% Fe; 4.88% Zn; 8.22% Pb; 9.49% S2- and 422 g/t Ag. The baseline cyanidation yielded 36.60% of silver at high cyanide consumption (24.17 kg/t NaCN). Pressure oxidation at 150°C (in the absence of iodide), followed by cyanidation of the residue yielded 8% silver extraction. Pressure oxidation was carried out at 110-150°C in sulphate media with additions of potassium iodide and the resulting residue was subjected to cyanidation. High recovery of silver was achieved because its incorporation into lead-hydronium jarosite during the pressure oxidation was minimised, while its deportment into the cyanide-soluble silver iodide was favoured (Figure 1). Silver recovery decreased slightly at high temperatures (140° and 150°C) because of the enhanced stability and kinetics of leadhydronium jarosite, but this could be overcome by increasing acidity and iodide concentration of the slurry."

Jan 1, 2003

By Kazutoshi Haga, Atsushi Shibayama, Kosuke Nishioka, Yasushi Kihara, Altansukh Batnasan

The aim of this study was to investigate the feasibility studies of seawater in flotation of a 1.5 mass % of copper containing copper sulfide ore. Effects of seawater, various frothers, and different distilled water/seawater ratio on flotation of copper sulfide ore were studied. Surface properties of native copper ore and sea water soaked were examined by using FT-IR spectrometry. Chemical contents of the copper sulfide ore and after flotation were identified using XRD studies. XRD pattern showed that the main constituents of the copper sulfide ore are chalcopyrite (CuFeS2), bornite (Cu5FeS4). Flotation experiments were carried out using mechanical agitation flotation equipment with a mineral separation type flotation cell. Froth flotation experiment was conducted using calcium hydroxide was used as a pH adjuster, using methyl isobutyl carbinol (MIBC), dow froth 250 (DF 250) as frothers and potassium amyl xanthate (PAX) as collector for the copper recovery from the copper sulfide ore containing chalcopyrite and bornite. The results showed that the copper recovery by flotation was dependent on the flotation medium and frother. The main results showed that the copper recovery by flotation was dependent on the process water and frother. The copper recovery reached 97 % and 87 % in the distilled water flotation and seawater flotation with MIBC, respectively. For seawater flotation with DF 250, the Cu recovery achieved near 95 % and grade of Cu in the froth product after 5 min flotation was reached 29.8 mass %.

Jan 1, 2014

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

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

Jan 1, 2018

By Yuqing Feng, Ping Fang, Dezhou Wei, Zichuan Guan, Shuling Gao, Baoyu Cui

"Air core, as an inevitable phenomenon of rotational flow field inside hydrocyclone, has a major impact on hydrocyclone’s separation characteristics. The air core formation process, its structural features and its influencing factors were investigated systematically in this paper. Numerical study on the air core of a f50 mm hydrocyclone was carried out using a commercial computational fluid dynamics (CFD) modelling software package ANSYS/Fluent with the selection of Reynolds Stress Model (RSM) for the turbulence model and Volume of Fluid (VOF) for capturing the air-liquid interface. Physical experiments were conducted under the same geometry and operation conditions and the particle image velocimetry (PIV) technology was used to measure the liquid flow. The CFD results are compared with the experimental measurement with good agreement demonstrated. The air core was formed generally at around 0.7 s, air outsides of the air core was eliminated completely after 1.7 s. The centreline axial velocities and the water blocking hydrocyclone simulations showed that air of the air core entered the hydrocyclone from the underflow area and exhausted from the overflow area, but there was also some air flowing into the air core from the overflow area. The largest swing of the air core occurred on the top of the underflow pipe, and a protuberance similar to the shape of Adam’s apple appeared when the air entered the vortex finder. CFD simulation results also showed that the air core diameter was decided by axial position and parameters of hydrocyclone. The structural parameters have more effects on the air core diameter than the operational parameters."

Jan 1, 2014