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By Gülay Bulut, Mustafa Özer, Yunus Emre Benkli, Olgaç Kangal, Istanbul Technical University, Mustafa Tarkan

"Zeolites are naturally occurring as hydrated aluminosilicate minerals. They belong to the class of minerals known as “tectosilicates” and are related to the common minerals feldspar and quartz. Seven natural zeolites that occur in large, commercially exploitable deposits, are: clinoptilolite, chabazite, mordenite, phillipsite, erionite, analcite, and ferrierite. Among them, clinoptilolite is the most abundant one and is not the best known, but is one of the more useful natural zeolite. Clinoptilolite [(Na3K3) (Al6Si30O72) 24 H2O] is very closely related to heulandite and is currently being considered for disuse it self as it may just be a variety of heulandite. It differs from heulandite significantly only in its enrichment in potassium and slightly more silica and it is argued that a separate mineral is not needed.There are large zeolite deposits in western part of Turkey. The sample used in this study was taken from the Gördes region of these deposits. Screen, chemical, XRD and mineralogical analyses were used to determine properties of the natural zeolite sample used in this work.According to the size distribution of the sample, it is found that the zeolite sample used in this study is below 1 mm in size and almost 80% of it is below 0.5 mm. X-Ray Diffraction (XRD) and mineralogical analyses show that the natural zeolite is found to be in the forms of clinoptilolite and heulandite. Clinoptilolite forms as a devitrification product of volcanic glass in tuffs. Quartz, clay minerals, orthoclase, plagioclase, biotite, seruzite-muskovite and epidote were observed as impurities. XRD analysis also shows that the sample contains approximately 58-60% clinoptilolite, 5% quartz, 5% feldspar, 10% smectite, 5- 10% illite and mica and remaining volcanic tuffs. Several types of mineral processing methods were performed in order to remove impurities from the natural zeolite sample to produce high purity zeolite. Gravity separation methods such as shaking table and superpanner were carried out for separating heavy minerals, while high intensity dry magnetic separation was used for discarding magnetic minerals."

Jan 1, 2003

By Peter J. Coghill, Michael J. Millen, Brian D. Sowerby, Kirby D. Cunningham

"CSIRO Minerals has developed an integrated particle size analysis system for deployment in the grinding circuits of minerals processing plants. The system includes a slurry sampling station, a de-aeration tank and an ultrasonic particle size analyser (UltraPS). The UltraPS analyser is based on ultrasonic attenuation and velocity spectrometry combined with gamma-ray transmission. The technology is applicable to the measurement of 0.1 to 1000 µm diameter particles directly in highly concentrated slurries without dilution. The analyser incorporates a gamma-ray transmission gauge utilising an ultra-low activity radioisotope source that is exempt from International Atomic Energy Agency (IAEA) and Australian licensing regulations.The UltraPS analyser consists of a submersible probe and an electronics package. The probe can be immersed directly in a minerals processing plant stream if the aeration level in the stream is low. For applications where a de-aerated slurry stream is not available, CSIRO has developed an integrated system that includes a sampling station, a de-aeration tank as well as the UltraPS analyser itself. A schematic drawing of the continuous flow gravity fed system is shown in the figure. A commercial sampling system is used to supply slurry at a rate of 2.5m3/hr to the deaeration tank. The aim of the de-aeration tank is to remove air in the slurry to a level where meaningful particle size measurements can be performed. The tank volume is about 0.2m3, thus giving the slurry flowing through the tank a mean residence time in the tank of approximately five minutes. Tests at a number of concentrators have shown that five minutes is sufficient time for slurry to de-aerate to a level where ultrasonic measurements can be made. The tank incorporates an ultrasonic level detection sensor and flow to the deaeration tank is adjusted based on the level of slurry within the tank. The UltraPS probe is mounted through a flanged section of duct, which is bolted to the bottom of the de-aeration tank."

Jan 1, 2003

By Olga D. Khmelnitskaya

The paper presents the results of research on the development of gold recovery technology from oxidized gold-copper ore of one of deposits in Russian Federation. Gold and copper grades in the ore are 9.5 g/t and 0.4%, respectively. Various options of flowsheets, such as direct cyanidation, thiourea leaching, flotation and gravity concentration followed by subsequent cyanidation of con-centration products, were tested. It was established that gravity-cyanidation technology with filtra-tion leaching tails and the use of solutions in the recirculation is the optimal option for ore treatment. The proposed flowsheet provides gold recovery for 91.0% at NaCN consumption 2.5 kg/t ac-cording to the results of pilot testing. Ammonia-cyanide gold leaching was also examined. In comparison with cyanidation, gold leaching solution in system NaCN-NH4OH-(NH4)2CO3 allows NaCN consumption to be reduced three times from 4.5 kg to 1.5 kg per 1 ton of the ore and provides the same gold recovery. Based on the laboratory and pilot plant tests results of gravity-cyanidation technology of gold dissolution, process design criteria were developed.

Jan 1, 2014

By Sung. Hyejin, Kyoungkeun Yoo, Oh-hyeok Kown

The sediments near coasts and ports in Korea have been found to be contaminated with heavy metals and organic matters. So far the sediments have been dumped to landfill sites or ocean. However, the number of available landfill sites is decreasing and ocean disposal of wastes was strictly prohibited. Although physical separation processes have been found to be useful for the remediation of contaminated soil with heavy metals, because the sediments consist of fine particles, it is difficult to remove heavy metal component by physical separation processes. Furthermore, the sediments contain organic matter, so there is a concern about strong stink during the sediment treatment. The present study is aimed to develop the treatment process for removing heavy metals and organic matter from the contaminated sediments. The heavy metals and organic matter were oxidized and dissolved by electro-generated chlorine, and the chlorine used is regenerated and circulated in the leaching process. This system has economic and environmental advantages to reduce the wastewater and the oxidant consumption. The leaching results show that the metal removal efficiency increased to more than 99%, and more than 50% of organic matter was decomposed.

Jan 1, 2014

By Tomás Vargas

Treatment of copper sulphides minerals via leaching-solvent extraction-electrowining (lix-sx-ew) have been for years the process scheme proposed from hydrometallurgy as an alternative to the conventional flotation-smelting route. Many different hydrometallurgical technologies of this type have been proposed, investigated and tested over the years, but until today none of them has come close to endanger the predominance of the flotation-smelting route in the copper industry. However, bioleaching introduced an important breakthrough in this context as this technology enabled the alternative hydrometallurgical treatment of secondary copper sulfide ores and low grade ores by applying heap bioleaching processes. Extension of its application to primary copper ores have been limited until now by the nature of chalcopyrite which is very refractory to leaching. However, important improvements are expected by using of improved designs that enable to operate at high temperature ranges where copper is leached faster under the catalytic action of thermophile microorganisms. The present work analyzes future scenarios for the application of this technology either in heaps or in-situ, which can become very attractive as copper can be produced reducing drastically the energy consumption associated to the mining, crushing and grinding stages.

Jan 1, 2014

By Yimin Zhu, Dezhou Wei, Haijun Liang, Zhifu Zhou

Pyrite is one of sulfide minerals most widely distributed in the earth. During the flotating separation of metal sulfide and the floating desulfation of coal, it is often requaired to restrain pyrite. However, the result is not very satisfactory because of the little difference of floatability of minerals to be separated. The problem is extremely outstanding for the floating desulfation of coal. Researchers carried out a number of experiments with T. ferrooxidans (Yang et al, 2001; Zhou et al, 2000; Zhou et al, 2001; Wei et al, 2001) for an enhanced restraint to pyrite. The results showed that the T. ferrooxidans after being cultured works the best in restraining pyrite. After 10 min pretreatment, the removal rate of pyrite from coal is up to 80%~90%. The research results by Yuan Xin et al (Yuan et al, 2000) show that T. ferrooxidans can restrain effectively the floatability of pyrite and can reduce the recovery of pyrite from 95% to 8%. Though the restraining properties of T. ferrooxidans to pyrite have been proved by many tests and experiments, there are very limited reports on its restraining mechanism. Thus, a systematic experimental research is carried out in this paper.

Jan 1, 2003

By G. S. Hansford, G. E. Searby

"Ferrous iron oxidation kinetics was measured in continuous culture at temperatures between 65 and 75ºC, using thermophilic microbes suitable for high temperature bioleaching. The resulting specific oxidation rates were modelled as a function of the ferric/ferrous iron ratio. The results were compared to those obtained by conventional mesophilic microbes.Microbially mediated ferrous iron oxidation is a vital sub-process in the mesophilic bioleaching of sulfide minerals and is expected to be of similar importance at elevated temperatures using thermophilicmicrobes.A thermophilic archaeal culture used in chalcopyrite bioleaching was successfully grown in continuous culture on ferrous sulfate, indicating the capability of the microbes to sustain growth for long periods on the energy provided by the conversion of ferrous iron to ferric iron, the primary leaching agent in sulfide mineral bioleaching.Steady state data was obtained at residence times of between 70 and 10 hours, with washout occurring at residence times between 11 and 9 hours, corresponding to a maximum growth rate (µmax) of between 0.09 and 0.11 hr-1. The reactions were followed by measurements of the solution redox potential, soluble iron concentrations, and offgas analysis of the oxygen and carbon dioxide utilisation rates."

Jan 1, 2003

By Hongchun Lou, Yan Zhang, Xianjun Lu

"The potential of Fe(NO3)3 modified attapulgite was assessed for the adsorption of phosphorous removal from aqueous solutions. Natural attapulgite was modified by Fe(NO3)3, and thus the composite adsorption materials could be obtained. Also phosphorous adsorption characterization on the iron-coated attapulgite was investigated in this research. Taken overall the results suggested that phosphorous adsorption capacity could be improved markedly by adopting ferric modified-attapulgite. The optimum phosphorous adsorption were developed in the experimental conditions of the solid-to-liquid Ratio 1g/L, the initial phosphorous concentration 5 mol•L-1, reaction time 120 min, temperature 25 ? and pH=4.0, the removal efficiency is 92.28% for the modified attapulgite. In comparison of batch experimental results of phosphorous adsorption onto natural and iron-coated attapulgite, it has provided much more likely the same change trend. On the other hand, the outcomes of this analysis indicated that ferric modified adsorbent was notably influenced by adsorption element over natural attapulgite. The Freundlich isotherm adequately described the adsorption course of phosphorous onto ferric modified-attapulgite with the correlation coefficient R2 goodness of fit of 0.9690."

Jan 1, 2014

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 Y. H. Hu, X. X. Zhang, Z. F. Dai, H. J. Huang, Q. S. Huang

"In order to get a more comprehensive understanding of the interaction of fine fluorite particles and bubbles, the electro-flotation was conducted, a high speed CCD was used to determine the bubble size distribution, and the results were analyzed using the models of collision and attachment. The electro-flotation was performed in a modified Hallimond tube. The results show that for different fluorite with size of 0 -10µm, 10 -38µm, 38-74um, there is a matching range between the different particle sizes and cathode aperture size (38um, 38um, 74um) to reach highest recovery. In addition, improving the concentration of collector is benefit for the flotation recovery due to the ability of collector to decrease the bubble size and prevent the merger of bubbles. Furthermore, theoretical calculation of collision, adhesion probability and collection probability is in agreement with the experimental data for electro-flotation recovery of fluorite.INTRODUCTIONFluorite is widely distributed, and has been found in more than 40 countries around the world. As fluorite minerals are easy to float, flotation is used as the best method of the current recovery of fluorite (Yang and Ai, 2016; Chen et al., 2017). In the field of mineral processing, the flotation of fine particles is particularly difficult, caused by the small quality and large surface (Sun et al., 2006). Flotation is now considered to contain three subprocesses, including the collision, attachment and detachment between the particles and bubbles (Sun et al., 2011). The collision of particles with bubbles is related to factors such as particle size and density, bubble size, bubble rise rate (Qin et al., 2012). Zongfu Dai et al. (2000) summarized the particle-bubble collisions models in detail and compared the experiment results with the model predictions. However, particle-bubble interactions occur in complex and turbulent flotation environments (Xing et al., 2017). Both thermodynamics and dynamic factors affect the adhesion of particles on bubbles surface (Schulze, 1989; Shahbazi et al., 2010). Scheludko et al. (1976) believes that the necessary condition for successful adhesion of particles on bubbles is that the contact time is greater than the induction time. The probability of adhesion between particles and bubbles increases with the increase of bubble size, but if the particle size is too small, the probability of adhesion decreases instead (Ren, et al. 2014; Norori- McCormac et al., 2017). After the particles adhere to the bubble surface, the particles will slide down to the bottom of the bubble due to gravity. At this point, the gravity, inertial centrifugal force and fluid force promote the desorption of particles from the bubble surface, while the surface tension, surface hydrophobic force maintain the stability of particles-bubble combination (Ralston et al.,2017). In addition, in the process of rising to the foam layer, desorption can also occurs. The purpose of this study is to observe the flotation behavior of fine fluorite with different size and to get a more comprehensive understanding of the interaction of fine fluorite particles and bubbles."

Jan 1, 2018

By Fedotov Pavel

"Today the main objective of improving the process of ore dressing is to reduce energy consumption and performance of the basic rule of auxiliary enrichment processes. One of the modern means of ore destruction with a minimum of energy consumption is the interparticle fracturing, for example in the roller press. There still remain the unsolved problems of the energy consumption determination and the particle size distribution of the crushing product .Existing methods for determining these parameters are empirical and require considerable expenses and time. The improved finite element was developed to solve these problems using finite element contact interaction of the rock. Also it was developed to clarify the stress-strain and the rock fracture energy in a layer. These elements are used to study ore destruction in the layer and allows to create a model taking into account the interaction of crushing particles surfaces. The same algorithm was developed determining crushing product particle size distribution in the layer on the base of the Bond’s theory of crushing. The presented finite element model allows to determine the quantity of strain energy and particle size distribution of crushed product during the destruction processes of the material in the roller press with fair accuracy."

Jan 1, 2014

By S. K. Pan, M. P. Srivastava, N. Prasad

"Dalli mines of Bhilai Steel Plant, India is fully mechanized iron ore mine having wet processing system. It has a capacity to treat 5 Mt of ROM. In processing circuit of Dalli mines, scrubbing, wet screening and classification operation is being carried out. Average slime generation for the last five years is about 14% of ROM. As the mining has reached to the lower benches, fines generation is more. As a result, very fine and heavy particles, which settle fast, are being discarded with the slime to tailing pond. These heavy and faster settling particles were causing jamming in the thickeners.Laboratory investigations were carried out at Research and Development Centre for Iron and Steel (RDCIS) to suggest a suitable scheme to recover high quality fines so that the availability of thickener can improve. The slime samples were first treated in a floatex density separator. The concentrate obtained from the floatex density separator was treated in hydrocyclone. Laboratory investigations indicated that the slime with 49% Fe, 12.7% silica and 10.7% alumina can be upgraded to 62.1% Fe, 3.3% silica and 2% alumina with 36% material recovery. Based on the laboratory studies ,a system was installed at Dalli minesThe general arrangement of the system is shown in Fig. below."

Jan 1, 2003

By B. McFadzean, K. Moller

"The mechanism whereby thiol collectors adsorb onto a mineral surface has been widely studied. Most of these studies examine the mineral surface at equilibrium state and, often, well above monolayer coverage. However, the formation of species at the mineral surface is a complex process that is both kinetically and thermodynamically controlled. In a typical industrial application, the dosage of collector will be around monolayer level. This study uses isothermal titration microcalorimetry to measure the heat evolved during the reaction of sodium ethyl xanthate (SEX) and sodium isobutyl dithiocarbamate (SIBDTC) and their mixtures with galena and chalcopyrite. The sensitivity of the instrument allows heats of reaction of sub-monolayer quantities of thiol collector to be measured.The study found that both the magnitude and kinetics of the reaction between a collector and a mineral surface was dependent on the type of mineral and the type of collector. Proposals are made for the reactions based on literature and the evidence from this investigation. It was found that when collectors were reacted with mixtures of galena and chalcopyrite, that the magnitude of the reaction was lower compared to that measured for the pure minerals alone. This may be due to galvanic interactions between the minerals, resulting in a passivating effect on one or both of the minerals."

Jan 1, 2014

By Hamid-Reza Manouchehri, Hassna Aitahmed-Ali, Stellan Sundberg

"In flotation practices as the particle size becomes fine, e.g., p<20µm, the process efficiency declines markedly. When flotation of fine particles is necessary, one method to improve the recovery is selective agglomeration of fines. Different agglomerating methods have been introduced and tested; however, one method which has recently received great attention is to aggregate fine particles by magnetic force. A selective aggregation of the fine paramagnetic particles, e.g., sulfide minerals, can be achieved by applying a high intensity magnetic field during flotation conditioning. The method is simple, selective and effective for fines, and has a low operation cost. However, a minimum magnetic susceptibility of fine grains is required for this process. This method has been demonstrated in different sulfide minerals flotation plants. The implementation of the magnetic conditioning device, ProFlote, in both the Cu-Pb and the Zn circuit of Boliden’s Garpenberg plant in Sweden, where a complex massive sulfide ore is treated, is discussed herein. The preliminary results have shown an increase in the recovery of value minerals, in particular Zn and Ag and considerable reduction in Zn deportment to the final tail. Increases in Cu grade and decreases in Pb grade in Cu-con have been achieved as well."

Jan 1, 2014

By Matt Kilburn, Liming Lu, Steve Suthers, Alexander Edenton, Rodney Smyth, Matt Adam

"Magnetite concentrate is naturally good for pelletising. However, due to the depletion of reserves of traditional high grade hematite iron ore worldwide, the usage of magnetite concentrate in sintering is expected to increase substantially. Magnetite concentrate is typically high in Fe grade and releases extra heat when oxidised to hematite. This will not only improve the quality of agglomerates produced, but also have a potential beneficial impact on the fuel consumption of the agglomeration process. There are an increasing number of magnetite projects currently being considered and developed. However the concentrates from some of these projects are too coarse as pellet feeds and much finer and denser than normal sinter feeds. Accordingly, special strategies are required in order to utilise the coarse magnetite concentrates at high proportions in pelletising and sintering. This paper discusses various raw material preparation technologies and blending optimization methods demonstrated to be effective for improving the sintering and pelletising performance of coarse magnetite concentrates. Pre-granulation and addition of binders, such as hydrated lime, were found to be particularly efficient in recovering the sintering productivity of sinter blends containing magnetite concentrates. In pelletising, HPGR (High Pressure Grinding Rolls) grinding enabled production of good quality pellets from 100% magnetite concentrates and optimized blends containing up to 70% magnetite concentrates also formed good pellets.INTRODUCTION Magnetite concentrates are often used as feed materials for pellet plants or sinter plants depending on their grades and grain sizes. To achieve the required quality of green and fired pellets, the concentrates often need to be ground very fine (80% passing 38- 45 m), which inevitably consumes a great amount of energy. Therefore, efficient grinding processes are required to reduce the amount of energy consumption needed to achieve the sizing and chemical specification of pellet feeds. The use of magnetite concentrates in sintering has the potential to improve sinter chemistry, which is deteriorating over time due to the slowly decreasing grade of direct shipping iron ores around the world (Yang and Mathews, 1997; Yang and Witchard, 1998, Yang and Davis, 1999; Yang 2005). However, methods are required to allow the utilisation of high levels of concentrates in sintering while maintaining productivity and sinter quality (Han and Lu, 2017)."

Jan 1, 2018

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 K. Bru, D. B. Parvaz

"Electric pulse fragmentation (EPF) is an innovative technology that uses high-voltage pulsed power (HVPP) for the selective comminution of the treated material. Highly energetic electrical pulses are used to fracture rocks at mineral grain boundaries, resulting in extensive micro-cracks and/or complete fragmentation of the processed particles depending on the applied energy. In this study, the EPF technology was applied to a low-grade schist ore containing cassiterite, in order to investigate its influence on the selectivity of the fragmentation. The impact of EPF on subsequent comminution steps was studied by combining either the EPF technology or jaw crushing (used here as conventional crushing) with subsequent impact crushing. The results confirmed the potential of the EPF treatment for a selective fragmentation of the schist ore. This gave the potential advantages of reducing the total energy consumption in the comminution circuit and of improving the recovery of cassiterite during concentration processes. INTRODUCTION The rise in the world consumption of metals combined with the depletion of most of the high-grade ore reserves leads to an increased need to process ores of lower grade, with a more complex mineralogy. To recover the valuable minerals contained in such ores, fine grinding is required, but this process has high energy requirements and leads to the production of a large amounts of fines, which can impede the performance of later concentration processes such as gravimetric separation and flotation. Addressing this issue is a major challenge for the mining industry. Selective comminution methods have the potential to significantly contribute to the profitable exploitation of these low-grade ores by generating cracks specifically at mineral grain boundaries rather than randomly as in a conventional mechanical comminution system. This allows the liberation of valuable components from the feed material without overgrinding, i.e. this prevents unnecessary size reduction of already liberated grains, which leads to a reduction in production of fines and in energy consumption. In addition, rock not fragmented by the process are usually weakened by the dense fracture network generated, which increases grindability of the rock (Wang et al., 2011)."

Jan 1, 2018

By Andrew T. McBride, Malcolm S. Powell, Indresan Govender

"Current empirical Semi Autogneous Grinding (SAG) mill models contain assumptions concerning the amount of energy available for particle breakage, how this energy is distributed in the mill media and the rate at which the energy is transferred. Despite these assumptions the JK SAG model is highly functional with a proven track record. DEM provides a numerical framework to greatly improve on these assumptions that were necessary due to a lack of available input data.This paper identifies key outputs of the Discrete Element Method (DEM), such as energy spectra and impact probability distribution. It describes how they can be incorporated into current mill models utilising practical rock breakage data. The rate at which energy is transferred is shown to be dependent on particle position and mill rotational speed and not once per cycle as previously assumed. Based upon this an improved energy transfer methodology is presented. The fundamental outputs of the DEM are validated against experimental three-dimensional particle tracking data, so as to establish whether the DEM produces correct energy transfer predictions on the inter-particle level, which is crucial to this work. • Effective Experimental ValidationAn automated 3D tracking technique is employed to study the motion of plastic particles deep within the charge of a 142mm diameter experimental mill. A Biplanar Angioscope is used to film the mill with X-rays in two planes simultaneously at 50 frames/s, with a shutter speed of 1/3000s. A marked ball within the charge is tracked for over 2 minutes and the resulting digital images processed by a fully automated imaging technique, which locates the 3D coordinates of the particles to within 0.2mm. With over 250 circulations, this provides statistically rigorous and accurate data for verification in the numerical modelling of mills.A series of validation techniques have been developed to compare the experimental trajectory data with the outputs of both 2D and 3D DEM simulations. The techniques describe the shape of the charge and the particle motion within mill, allowing objective comparisons to be made of key features of the charge, such as the equilibrium surface and the centre-of-circulation."

Jan 1, 2003

By Vladimir Grujic, Milorad Grujic, Ivan Djurkovic, William Riggs

"In this paper, we have studied efficiency of different chemical structures of frothers on copper ore Majdanpek. The goal was to find out the best suitable chemical structure, which allows good bubble formation, size and interaction, forces on mineral surface. The particle size distribution was between 10 and 100 µm. Our research was oriented on three groups of frothers: alcohols, poly glycol ethers and poly glycols.We also used different collectors based on similar or less similar alcohol group in chemical structure as in frothers, expecting higher stability of air bubble in contact with copper minerals. Minimal potential energy was determinate in the hydrodynamic conditions for different particle sizes. It was founded that it is proportional to air bubble volume on 0.88° and contra proportional to the kinematics viscosity of pulp.Analyzing the results, we find out that poly glycols decreased surface tension the most. Also, it increased amount of coarser particles to the top of pulp. This group is less sensitive to pH value compared to the others too.From the class of frothers with higher efficiency in flotation process, we pointed OrePrep X-75 and X-133 as frothers with expressive frothing characteristic in the alkaline and low alkaline circuits. Special suitability of those frothers for floating copper minerals makes their hetero polar molecules with more than one hydrocarbon radicals connected to one polar group. However, those froters have no nether high or low solubility and are non-ionization. Those froters indicate some small collecting characteristics too. In dispersion systems, those two frothers make two types of bubbles: smaller di < 1.2 mm; V = 0.0008 cm3 and middle di < 16 mm; V = 2.2 cm3 (di – Bubble diameter, V – Bubble volume). Transportation speed of air bubbles it is possible to calculate based on equation: ?s = 27.5 * di 1.4 cm/s. Distribution of free surfaces for chalcopyrite is average 80% in the class –74µm +10µm and 20% in the class -10µm. If we get all class sizes, we find that recovery (R) depends from the time by equation:"

Jan 1, 2003

By Glasswell K. Nkonde, Fredrick Chileshe, Steven Simukanga

Improving food security by increasing crop output is a primary objective for most countries in the sub- Saharan region. Most soils in the region are deficient in phosphorus an important nutrient after nitrogen in plant growth. Phosphates are therefore used together with nitrogen and potassium based fertilizers in improving crop production. Zambia like many other countries in the sub-Saharan region imports phosphate fertilizers for use by its farmers. However, indigenous phosphate deposits of igneous origin exist in Zambia of which the main ones are Chilembwe in Petauke district, Mumbwa sugarloaf in Mumbwa district, and Nkombwa Hill in Isoka district. If commercially exploited, these deposits could become a ready source of inexpensive phosphate for the local farming communities in the country. Previous research has shown that partially acidulated phosphate rock (PAPR) produced using Chilembwe phosphate rock is just as agronomically effective as the imported superphosphate fertilizers on food crops in most of the Zambian soils. PAPR is produced by reacting finely ground phosphate rock with only a fraction (usually 50%) of the stoichiometric quantity of the sulphuric acid required to fully convert the apatite in the rock to monocalcium phosphate thus making the process cheaper in terms of acid consumption than the poduction of single superphosphate where the apatite in the rock has to be fully converted to monocalcium phophate. The factors that affect the percentage of available P2O5 in the PAPR produced include particle size of phosphate rock, degree of acidulation, concentration of acid, process temperature profile, residence time and drying temperature. These factors were studied in a 25kg/hr PAPR pilot-plant in the School of Mines at the University of Zambia using Chilembwe phosphate rock. The PAPR pilot-plant consisted of a rotary drum granulator as the main reactor, a rotary drum dryer and a triple deck vibrating screen. Ground phosphate rock was reacted with sulphuric acid and water in the granulator to form a granular product which entered the drier. The dry product was screened into three size fractions. The 1 – 4 mm size fraction was the final product whilst the –1mm size fraction was recycled. The +4mm size fraction was sent to a crusher while the crushed PAPR was sent back to the screens.

Jan 1, 2003