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By Thi Hien Tran, Huy Hoan Nguyen, Thi Thu Hien-Dinh, Hong Son Nguyen, Duy Anh Dao

"Lithium discovered in La Vi deposit, Quang Ngai province in the Central of Vietnam was estimated with a reserve of approximately 1.5 million tones of lithium ore, which was characterized by mainly lepidolite with muscovite, albite, and quartz. This study investigated reasonable flowsheet and optimal conditions of enrichment technology for lepidolite from the mined ore by flotation process, and producing Li2CO3 and LiCl from lepidolite concentrate by hydrometallurgical process. The significant conditions for froth flotation of lithium ore including particle size, pH of ore slurry, pulp density, chemical spending (collectors, activators, depressants, and frothers), stirring and flotation time were investigated. The identified reasonable flowsheet and optimal conditions for froth flotation the La Vi lithium ore processed lepidolite concentrate containing up to 4 wt.% Li2O with a recovery of 87.61 wt.%. The processes of sulfation roasting lepidolite concentrate using FeSO4.7H2O, then leaching in H2SO4 solution, and removing impurities were carried out to prepare a pure lithium-bearing solution for producing Li2CO3 and LiCl. The products containing 99.05 wt.% of Li2CO3 and 99.02 wt.% of LiCl were obtained with recoveries of 82.97 wt.% and 85.77 wt.%, respectively. The obtained results indicated that lithium ore in La Vi deposit can be processed to produce commercial lithium products and lead the study to develop technology for a new industry of lithium processing in Vietnam.INTRODUCTION Lithium is the lightest metal, which has many superior properties in both metal and compound (such as Li2CO3, LiCl...) states since its discovery in 1817 by Swedish chemist Johan August Arfwedson. Lithium is a favourable material for many high technology fields such as aerospace industry and chemical industry. Some industrial fields which present the main consumptions of lithium include ceramic, glass, aluminum alloy, lubricants, rechargeable Li-ion batteries, synthetic rubber, and pharmaceuticals. The worldwide production of lithium in 2014 was 31,700 tones, continually increased in 2015 of 5% presented 32,500 tones (USGS, 2016). The demand for lithium tends to increase because of the raising in consumption of lithium batteries. Lithium is mined from salar brines, mainly in Chile and Bolivia, but also from granites and pegmatites, which contain minerals with high lithium content such as spodumene - LiAlSi2O6 (Heinrich et al., 1978; Linnen et al., 2012); petalite - LiAlSi4O10 (Sitando & Crouse, 2012); zinnwaldite - KLiFe2+Al(AlSi3O10)(F,OH)2 (Paukov et al., 2010; Kondás J. & Jandová J., 2006); lepidolite - K(Li,Al)3(Si,Al)4O10(F,OH)2 (Luong et al., 2013); and amblygonite - LiAlPO4F (Garrett, 2004)."

Jan 1, 2018

By S. H. Shahcheraghi

The four operational parameters including collector dosage, frother dosage, conditioning time and pH, are the most important parameters effective on the flotation kinetic. The present study has been done at the Miduk Copper Complex in the province of Kerman. The effect of operational parameters on the flotation kinetic of copper minerals were studied and analyzed, so the optimum conditions for flotation kinetic of copper sulfide minerals were obtained. Under the obtained optimum operational conditions, the recovery of copper sulfide flotation at the 95% confidence level was estimated to be between 92.76% and 95.48% and the copper sulfide flotation kinetic constant was estimated to be between 1.40 and 1.45 min-1. Results showed that the consumption of the X231 collector (O-isopropyl-N-ethylthionocarbamate) and the A65 frother (Polypropylene glycol) could decrease up to 35% and 50% respectively while the copper sulfide recovery increases about 1 up to 2% at the laboratory scale. Keywords: flotation kinetics, sulfide copper, optimization, the Miduk Copper Complex, response surface method

Sep 1, 2012

By M. F. He

The depression of pyrite in marmatite flotation by a new depressant Sodium Glycerine-xanthate (SGX) has been investigated through microflotation, zeta potential, and adsorption measurements in systems open to the atmosphere. The flotation tests of single mineral show that in the presence of SGX, marmatite can be activated by Cu2+ and shows good flotability, while pyrite can?t be activated and therefore shows poor flotability, as a result these two minerals can be separated using butyl xanthate as the collector. The depression mechanism of SGX on sulfide minerals is discussed based on Zeta potential and adsorption isotherm. Zeta potential experiments demonstrate that at the presence of Cu2+, SGX can strongly adsorb on the surface of pyrite, while it can?t adsorb on the surface of marmatite, thus the results are in accordance with those of flotation experiments; the results of adsorption isotherms show that the adsorption density of SGX on pyrite is greater.

Jan 1, 2014

By Z. Y. Ma, M. D. Liu, Z. Z. Liu, J. K. Zhou

"An optimization method of microwave assisted leaching of copper anode slime was proposed based on Taguchi method. The results showed that the solid-liquid ratio was the most important factor affecting SNR(signal noise ratio) of the copper, tellurium and selenium leaching efficiencies, and its contributions to the copper, tellurium and selenium leaching efficiencies were 58.34%, 55.54% and 60.74%, respectively. Solid-liquid ratio was a significant factor to SNR(copper, tellurium and selenium). Reaction period was a less significant factor to SNR(copper and tellurium) and concentration of sulphuric acid was a less significant factor to SNR(selenium) while microwave power was an non-significant factor, its contributions to the copper, tellurium and selenium efficiencies were 3.22%, 11.73% and 12.73%, respectively. The optimized parameters of the microwave assisted leaching of copper were as follow: microwave power of 500 W, reaction period of 10 min, solid-liquid ratio of 0.10 g·mL-1 and H2SO4 concentration of 1.0 mol·L-1. The optimized parameters of the microwave assisted leaching of tellurium and selenium were as follow: microwave power of 800 W, reaction period of 15 min, solid-liquid ratio of 0.10 g·mL-1 and H2SO4 concentration of 1.0 mol·L-1. The validation experiments under the optimized conditions has been carried out, the copper, tellurium and selenium leaching efficiencies were 98.66%, 84.32% and 64.54%, respectively. INTRODUCTION As important valuable by-products, copper anode slimes are collected from the bottom of the electrolytic bath in the electrolytic copper refining process(Kilic, 2013). The value of the anode slime is determined by the content of rare metals and precious metals(Liu, 2014). Generally, pretreatment of copper anode slime is the first stage of the whole process. The aim of the pretreatment process is to remove Cu, Te and Se etc and enrich the valuable metals. The methods of pretreatment process can be differentiated into: pyrometallurgical processing and hydrometallurgical processing(Amer, 2003). The hydrometallurgical processes are more promising compared to pyrometallurgical processes in terms of cost effectiveness, low energy consumption and environmental concerns(Ranjbar, 2014). However, the disadvantages of the hydrometallurgy process are evident, such as lengthy processing times and low efficiency. Therefore, it is necessary to develop a highly efficient and environmentally friendly pretreatment method."

Jan 1, 2018

By D. W. Feng

Thiosulfate leaching has been considered as the most promising alternative to cyanidation for the recovery of gold from ores. The chemistry of the ammonia-thiosulfate-copper system is complicated due to the simultaneous presence of complexing ligands such as ammonia and thiosulfate, the Cu(II)-CU(I) redox couple and the stability of thiosulfate in solution. However, acceptable gold leaching rates can be achieved by maintaining suitable concentrations of thiosulfate, ammonia, cop-per and oxygen in the leach solution. The leaching of gold in thiosulfate solution is an electrochemical reaction involving the oxidation of gold to gold thiosulfate and the reduction of Cu(II) to Cu(I). Previous studies have shown that the reaction rate of gold was much higher in the presence of am-monia thiosulfate than corresponding sodium thiosulfate. This indicates that cations may affect the anodic oxidation of gold in thiosulfate solution. The present study utilises electrochemical tech-niques to study the anodic oxidation of gold in thiosulfate medium in the presence of the different alkali metal and alkaline earth metal ions. It has been shown that the presence of the heavier cations such as potassium, calcium, and caesium enhance the oxidation of gold in thiosulfate solution. Leaching tests also indicated the improved recovery of gold from a sulfide ore with the addition of heavier cations. XPS results demonstrated that the binding energy for gold shifted to lower values in the presence of the heavier cations, indicating the ready formation of ion-pair with gold thiosulfate at the gold surface during adsorption.

Jan 1, 2014

By Nicolás Caussade

"The IMPC 2014 slogan "Optimizing Water and Energy Usage for Sustainable Mining" refers to the ever decreasing availability of fresh water and energy, which has turned out to be a crucial aspect in assuring the competitiveness of the mining sector in Chile and the Andes region in general. The concern for sustainability and the supply of water and energy is nowadays part of the top priorities of mining countries and communities. This is also the case for Antofagasta Minerals. Several years ago its mining group implemented an energy policy based on intelligent, strategic and sustainable management, in order to successfully face the challenges imposed by the market. It is called an intelligent policy because it intends to provide energy through a diversified range of sources, which includes the development of in-house alternatives. It is also a policy focused on efficient energy consumption and one that is strategic. Because it is in line with the characteristics of the mining activity, it is developed for the long term. Finally, what makes the policy sustainable is that it gives priority to energy usage from renewable resources and as such, the required capacities have been developed to generate innovative solutions that respond to the restrictions that characterize the industry today."

Jan 1, 2014

By M. Lundström, J. Aromaa, S. Seisko

"Gold dissolution was investigated in ferric chloride media, promoted as one of the most promising alternatives for cyanide-free gold leaching. The characteristics of gold open circuit potential (OCP) and its relation to gold dissolution rate was investigated. Furthermore, the reaction mechanism of gold dissolution in ferric chloride leaching was discussed. The effect of solution variables, [Fe3+] = 0.02–1.0 M, [Cl–] = 2–5 M, T = 25–95 °C and pH = 0–1.0, on OCP and dissolution rate were studied electrochemically using rotating disk electrode (RDE) and linear polarization resistance (LPR) method with ?cyc = 1000 RPM. The investigated gold OCPs in ferric chloride solutions were found to vary between 597 and 688 mV vs. saturated calomel electrode (SCE) and gold dissolution rates between 1.2·10–5–5.0·10–4 mol s–1 m–2. Increase in temperature (25–55 °C) and ferric concentration (0.02–1.0 M) changed OCP into more positive values, while increase in chloride concentration (2–5 M) decreased it. The logarithm of gold dissolution rate was shown to increase linearly with increasing OCP at the temperature range 25–95 °C, ferric concentration range 0.02–0.5 M and chloride concentration range 2–4 M. A notable increase in the gold dissolution reaction rate was evident at T > 55 °C. INTRODUCTION Currently cyanide leaching is the predominant method used in gold production from primary raw materials (Marsden and House, 2006). However, since the Baia Mare disaster in Romania in 2000, the use of cyanide has been the subject of international concern due to the toxic nature of the chemical posing a significant health threat if exposed to the ecological entities (Hilson and Monhenius, 2006; UNEP/OCHA, 2000). Therefore, several countries e.g., Costa Rica, Czech Republic, Germany, Hungary, Turkey, many states of the USA and provinces of Argentina, have started to ban or limit cyanidation via legislation (Laitos, 2012). Thus, alternative solutions, such as thiourea, thiosulphate, oil-coal agglomerates as well as halides have been investigated in order to replace cyanide as lixiviant (Adams, 2016; Aromaa et al., 2014; Aylmore, 2005; Hilson and Monhemius, 2006; Lampinen et al., 2015). Ferric and cupric chloride leaching have been reported to have advantage over cyanidation being capable to dissolve refractory gold minerals, without pre-treatment such as pressure oxidation or roasting (Angelidis et al., 1993; Lundström et al., 2014; Marsden and House, 2006; van Meersbergen et al., 1993). According to Aylmore (2005), 4% of publications for alternative lixiviants to cyanide in gold leaching were within category oxidative chloride processes including aqua regia and acid ferric chloride leaching. Further, some patents have been granted for ferric chloride leaching of gold (Abe and Hosaka, 2010; Lundström et al., 2016)"

Jan 1, 2018

By Hernán Menares, Diego Hernández

"Over the last few years, the copper mining industry has been facing a changing scenario with respect to both the market and society which has resulted in taking a closer look at the way in which mining is being done. The cycle of high metal commodity prices of the last decade left the industry with high costs. Together with increased environmental and social requirements, a decrease in ore body quality, scarcity of resources such as human capital, water and energy, has resulted in an ever-increasing challenge to make a profit out of new ore bodies but also to maintain the current operations in the longer term competitive. The major challenge is to create value for all stakeholders, exploiting ore bodies that nowadays are not considered economically attractive, which applies to both new ore bodies and existing operations, but in a sustainable way. In order to achieve this objective, it is necessary to approach the mining business differently in five main areas: (1) operational efficiency through better use of installed equipment and skilled labor force; (2) efficient use of energy; using opportunities for alternative energy sources and efficient design; (3) efficient use of water; diminishing losses and use of seawater as an alternative source; (4) minimizing environmental impact by eliminating effluents and controlling dust emission; and (5) a more closer, proactive and collaborative relationship with the communities allowing them to be an active player in the solution to assure a sustainable business in the long term. This presentation will show how Antofagasta Minerals is facing these challenges with a long term strategy that includes all of the above mentioned aspects and that is allowing the company to exploit its mineral properties in an efficient and successful way."

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 P. Braga, A. Chaves, S. França, A. Luz

"In Brazil, there are no official reserves of molybdenum, and the unofficial reserves are modest. Therefore, the molybdenum ore (molybdenite) produced as a co-product or by-product of emerald mining in the region of Campo Formoso, BA, has arisen as the only economic option. The flotation process is commonly used in Brazil for the recovery of molybdenite from emerald mining tailings, but it produces low-content molybdenite concentrates, thereby resulting in a product with low economic value. The molybdenite concentrate produced in Campo Formoso contains 83.5% MoS2 (50% Mo), which is less than the international standard product that contains > 90% MoS2 (54% Mo).This low value is due to the presence of talc, which is a naturally hydrophobic mineral that is also floated with the molybdenite during the flotation process. Separability studies were done in a Partridge-Smith cell using dextrin and quebracho as a molybdenite depressor in solution at different concentrations and pH values to produce a technical-grade concentrate (i.e., 90% MoS2). These basic studies were accompanied by measurements of the zeta potential and contact angle, which was determined by the capillary rise method. The results of these floatability studies were used in bench flotation tests, which were performed in a Denver laboratory flotation cell. A molybdenite concentrate of 93.4% MoS2 was obtained when dextrin was used at a concentration of 100 g/t."

Jan 1, 2014

By E. A. Brocchi

This work is dedicated to the chemical treatment of a typical EAF dust by chlorine content agents. The residue contains 48% Fe2O3 and 28% ZnO mainly in the franklinite structural form. The others less compounds are: 4.00% CaO, 2.60% MnO, 2.20% MgO, 1.08% PbO, 0.36% Cr2O3, 0.21% B2O3, 0.03% CdO and 0.01% NiO. The experiments purpose is selectively to remove zinc from the residue leaving the iron behind. Thermodynamics studies have indicated some roasting possibilities. The experimental results have shown that both HCl leaching and carbochlorination will remove iron and zinc without any selectivity. However, either the chlorine or the calcium chloride roasting seems to form, preferentially, the zinc chloride. So, it will be presented how the EAF Dust can be treated by the using of chlorine roastings process such as carbochlorination in order to separate the zinc and iron present in the frankilinite, main EAF dust content.

Jan 1, 2014

By Y. S. Sun, P. Gao, Y. F. Li, J. W. Yu

"Oolitic hematite ore is one of the most refractory iron ores in the world. It is not feasible to separate iron minerals from raw ore by conventional mineral processing methods. In recent years, a new technology called coal-based reduction followed by magnetic separation has been employed to produce metallic iron powder from oolitic hematite ore effectively. In order to find the nucleation mechanism, a thermodynamic analysis of the carbothermic reduction of oolitic hematite ore was conducted by the HSC 6.0 thermochemical software. The phase transition and microstructure changes during the coal-based reduction were also investigated by X-ray diffraction (XRD), scanning electron microscopy (SEM) and energy dispersive spectroscopy (EDS). The results showed that the iron oxides in the oolitic ore were reduced to metallic iron following the route of Fe2O3–Fe3O4–FeO–Fe in reduction process. The reduction reactions preferentially occur in the high activity regions to form a higher concentration of Fe elements, and then more supersaturated Fe separated out to form new phase nuclei. Under the action of homo-aggregation and principle of minimum free energy, the iron cores continuously absorbed migratory Fe around and grew into quasi-spherical iron particles embedded in slag phase. INTRODUCTION With the rapid increase of steel industry, premium grade iron ore in China has been gradually exhausted, and its supply has become bottleneck for the development of iron and steel industry (Guo et al., 2015). Therefore, it is urgent to exploit the refractory iron ores, such as oolitic hematite ore. Oolitic hematite ore is widely distributes in France, Germany, America, Canada, Pakistan, China, and among other countries (Maynard and Van Houten, 2004; Li et al., 2011). In China, the explored reserve of oolitic hematite ore is 3.72 billion tons, which accounts for approximately 1/9 of total iron ore resources of China (Gao et al, 2016). Oolitic hematite ore possesses complex composition,"

Jan 1, 2018

By N. Schreithofer

Varying flow regimes, particle populations and consequently different types of bubble-particle interactions characterize the different zones of a flotation cell. The understanding of the role of the different zones of turbulent intensity on the particle-bubble interactions is of primary importance to optimization and scale-up of flotation processes and machinery. In order to study particle-bubble interaction in turbulent flow a completely new approach was taken. A laboratory scale stationary turbulence device was developed and built. The set-up allows the observation of particle-bubble interactions in various turbulence intensities, particle and bubble sizes and surfactant concentrations. The experimental set-up consists of a transparent, specially designed tank in which liquid flows between carefully designed and placed obstacles that create different turbulent flow fields. Bubble-particle aggregates are introduced into this turbulent zone. The behaviour of the aggregates, such as attachment /detachment of particles to/from air bubbles is recorded by a High Speed Imaging (HSI) system (5400 fps). The images are then analysed and the dependence of attachment/detachment process is described as a function of turbulent intensity in different conditions for different surfactant concentrations. This paper introduces this method to study particle-bubble interactions in turbulent flow and correlation to Computational Fluid Dynamics (CFD) simulation. CFD modelling is validated for two and three phase systems; liquid-gas with liquid and solid tracers respectively. The CFD simulations using the RANS k-S turbulence model showed good correlation to the experimental data and the experimental method is a promising tool for investigation of particle-bubble interaction in turbulent flow.

Jan 1, 2014

By Ventsislav Stoilov, Angel Angelov, Ivan Korolev, Antti Remes, Todor Pukov, Stoyan Gaydardzhiev

"The process modeling and simulation studies aim for better process operation and favorable environmental impact. The objective of this work is to set up a simulation model of the Assarel copper flotation circuit for planning and optimization of the process with the aim of increased metal recovery. The model is parameterized based on laboratory flotation tests of ore samples from the Assarel mine together with mineralogical studies by X-ray diffraction (XRD) and Mineral Liberation Analysis (MLA). A combination of several techniques allows reliable identification of mineralogical composition of the selected samples. The changes in the proportion between sulfide minerals in various products along the circuit have been also documented. Further, this information is used in the simulation model and in data reconciliation procedures for establishing mass balance. The simulation accuracy is studied by comparing the simulations with the plant survey based mass balance. Constructed simulation model can be used to run alternative process scenarios. As an example, the flotation circuit performance with different reagent regimes is simulated based on data from batch flotation tests. Two reagents were considered for this purpose, dithiocarbamate-based one and xanthogen formate/mercaptan blend. It has been revealed that these reagents do not have the same influence on the selective flotation. The one improves flotation of gold and overall recovery but without notable selective action. In contrast, the other showed better selectivity towards copper, rejecting more gangue and lowering gold recovery. The model is also applicable for process optimization studies with different feed compositions, flowrates and circuit configurations. INTRODUCTION In the pursuit for elevated technological indicators, finding an optimal set of process parameters fully utilizing the potential of the equipment is a key factor to success. Over the past decades, there has been a substantial development in flotation circuit optimization through performance benchmarking using metallurgical modelling and steady-state computer simulation. Practical simulation engineering tools are based on experimental data obtained from laboratory tests and from plant surveys and operation data. With the help of flotation models, it is possible to study grade-recovery response with different feed compositions and throughputs, and with various cell-operating parameters, as well as to evaluate and design different flowsheet configurations. Outotec’s HSC Chemistry® software, which incorporates simulation and mass balancing modules alongside with several other tools, can assist with this optimization (Mattsson et al., 2015)."

Jan 1, 2018

By Shobhana Dey

A low volatile medium coking coal from Jharia coal field, India was used for this investigation. The proximate analysis of the sample shows that it contains about 25% ash, 21.4% volatile matter and 52.4% fixed carbon. The sizewise ash analysis of -0.5 mm coal indicates that -0.5+0.15 mm fraction contains lower ash (21.9%) than -0.15 mm fraction (29.7%). The flotation characteristics of the -0.5 mm coal were determined by release analysis. The study reveals that recovery of combustible at 15% ash is about 27% and 43% at 17% ash level. The low recovery is due to the presence of high ash in finer fraction. In order to improve the recovery, the -0.5 mm feed was classified into -0.5+0.15 mm and -0.15 mm fractions, and subjected to flotation separately. The recoverable combustible obtained from the release analysis of -0.5+0.15 mm at 15% ash level is 46%. The effect of collector dosage, frother dosage and aeration rate on flotation was studied. The responses of these variables in collectorless flotation of the -0.5+0.15 mm fraction containing low ash were also studied. A lower aeration and frother dosage favored the generation of clean coal with low ash. The ash content in the tailing stream from -0.5+0.15 mm flotation circuit is close to that of the original -0.15 mm fraction. The former is ground to -0.15 mm and the two fractions are processed together. As -0.15 mm fraction contains relatively higher ash, collector aided flotation using sodium silicate was performed to recover the clean coal with 17% ash. The products of the overall flotation circuit having different ash levels were recommended for use in different applications. The reject from -0.15 mm flotation circuit contained 56.4% ash which can be used for fluidized bed combustion (FBC). This eventually leads to complete utilization of coal.

Sep 1, 2012

By Mingfei He, Yude Gao

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

Jan 1, 2018

By Walter W. Focke, Zola Kwela, Willem van der Merwe, Arao Manhique

"The advantages of the de Wet zirconium process for the manufacture of AZST from zircon sand include: Fewer processing steps; removal of radioactivity by leaching rather than by precipitation; alkali reagent recovery in the form of a sodium silicate solution that is essentially free of radioactivity; fewer waste streams and less waste to discard.Here we report on our efforts at maximizing the alkali recovery in the sodium silicate product stream: Milled zircon, d50 ˜ 9 µm, was fused with caustic soda pearls in open reaction vessels at temperatures between 650°C and 850°C. Fusion times of 1, 2, 4, 24 and 336 hours (with intermittent milling) were investigated.For prolonged fusion times the fusion products approached equilibrium compositions. However, the phases Na2ZrSiO5 and Na4Zr2Si3O12 observed at 850°C did not form at 650°C. Since they do not hydrolyse in water they limit the recovery of alkali. For 2-hour fusions a broad spectrum of non-equilibrium reaction-products, that are rich in sodium silicates, is formed. Optimum process conditions, with respect to NaOH recovery, correspond to a 2:1 stoichiometry and fusing for 2-hour at 850°C. At these conditions the alkali recovery is ca. 78% and the zirconia yield is ca. 57%."

Jan 1, 2003

By G. K. Nimal Subasinghe

The performance of Knelson Concentrators (KC) used in gravity gold circuits often rely on manufacturers recommendations for selecting operating parameters such as, fluidising water flow rate and cycle time. These parameters are not only dependent on the machine parameters but also on the feed characteristics such as, size distribution of the gold and gangue particles and the composition of its constituent minerals. Particles entering a KC experience a centrifugal acceleration due to the high bowl rotational speeds. The required centripetal force to maintain a circular orbit is provided by the fluid drag force created by inwardly moving fluidising water and the Bagnold?s force arising from the particle/particle collisions. When the fluidising water flow rate is low particles move towards the outer rim of the bowl forming a packed bed within the grooves. On the other hand, if the fluidizing water flowrate is high the bed expands and some particles would move towards the centre of the bowl and gets carried to the tailings stream, thereby reducing the amount of material retained within the grooves of the bowl. Based on a force balance on a particle captured within the KC, a criterion which can be used to predict the amount of material recovered in the bowl has been derived and validated by experimental data. It has been shown that this criterion may be used to demarcate the transition of the bed from a packed-bed state to a fluidised state depending on the fluidizing water flowrate. By identifying the conditions for the optimal expansion of the bed comprising different constituent mineral particles, it is possible to estimate the optimal fluidising water flow rate that would maximise gold recovery while minimising gangue recovery. It was also observed that the accumulation of valuable mineral in the bowl takes place by the displacement of the gangue particles in the bed by the incoming heavy mineral particles. The kinetics of this segregation process was also studied using synthetic mineral mixtures. The experimental data suggests that this process follows first order kinetics and that the build up of heavy mineral reaches a maximum which could be used to identify the optimal cycle time. In this work, the suitability of the above model in optimizing the performance of a KC have been reviewed. Also the necessity and merits of testing an unaltered feed samples in a larger KC as op-posed to the commonly adopted GRG test has been discussed.

Jan 1, 2014

By Samira Rashidi, Robert Dunne, Raj Rajamani

"High Pressure Grinding Rolls (HPGR) are known for more energy saving, better particle weakening and higher tonnage compared to semi-autogenous/ball mill circuits. These machines are scaled up based on the roll diameter. Therefore, studies in lab-scale HPGRs are always preferred because less weight of feed sample is required to run batch tests. The 50-hp laboratory-scale machine at the University of Utah is the first-ever lab-scale machine designed by late Prof Schönert, the inventor of HPGR. This 200x100 mm [DxL] HPGR is equipped with a robust data acquisition system, which enables running steady state tests using a feed sample as small as 6 kg to 12 kg. This paper presents the HPGR grinding study on two different categories of ores, one hard ore and the other soft. A test series has been performed to evaluate the energy consumption and the relation with size reduction on open multi-pass and locked-cycle HPGR circuits. The effect of feed size distribution and operating conditions has been investigated. Locked-cycle tests result in higher reduction ratio and higher rate of fines (100 mesh, 48 mesh and 28 mesh) generation when compared to open multi-pass tests. Data shows that scale-up trend lines stay consistent even when the circuit configuration and operating conditions are varied."

Jan 1, 2014

By Zhongxue Li, Cuiping Li, Xiaoyun Hu, Yiqing Zhao, Huaying Wang

"Under the auspices of World Federation of Engineering Organizations’ Committee on Engineering and the Environment (WFEO-CEE) and Society for Mining, Metallurgy and Exploration (SME) Task Force on Sustainability and Mining (TFSM), and aimed at helping raise global understanding on and promote capacity building of risk management, this paper presents a framework and prototype of toolbox or knowledge warehouse for facilitating the application of risk analysis, mitigation and management techniques (RAMT) in the mining and minerals sector. The conception, design and implementation of the intended system of knowledge warehouse of risk analysis, mitigation and management techniques (KWoRAMT) are outlined. The system architecture of KWoRAMT is based on a browser/server (B/S) structured platform with web oriented interfaces and composed of data base, knowledge warehouse, searching mechanisms, and risk assessment methodologies. The knowledge sources for KWoRAMT will come from the TFSM expertise and Internet, including laws and regulations, codes of practice, standards, guidelines and best practices. The functions and outcomes of KWoRAMT consist of guided navigations with web links, categorized risk tools, annotated best practices with cases, and integrated risk approaches and solutions based on typical risk methods. Risk categories include those impacting on mine safety, mine occupational health, and mine community environment throughout the whole of mine life cycle from exploration, construction, and exploitation and processing to post-mining closure. Mining associated hazards will be identified and linked by vulnerability onto risk receptors through pathways. Risk management techniques for assessing impacts and consequences of mining induced hazards will be categorized with their pros and cons or strengths and weaknesses. Best practices and significant cases including experiences and lessons with problems, solutions, and recommendations on how solutions would be derived will be illustrated. Integrated risk approaches based upon typical methods will be developed for alternatives and actions."

Jan 1, 2014