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By A. V. Myaskov, V. S. Valavin, K. D’Elia, I. O. Krylov, Vu Van Toan, T. I. Yushina

"The purpose of this research is the development of a technology for processing the concentration tailings of limonite ores at Kamysh-Burun Mining and Processing Complex into an iron- containing product for ROMELT process. In the concentration and metallurgical processing of iron ores, substantial amounts of waste are generated (tailings, slimes, slags, metallurgical dust, etc.) that can be converted into a commercial product if is properly processed. This article presents the results of experimental studies of the feasibility of processing iron- containing tailings of Kamysh-Burun deposit (slimes from Verkhne-Churbash storage facility (VCh sample) and dumped waste of Kamysh-Burun (KB sample)) to obtaining valuable components for further uses. Applying Mossbauer spectroscopy has been shown that the iron minerals in the studied samples are technogenic varieties of iron hydroxides. The most suitable technology is the magnetic roasting process, since iron hydroxides are easily transformed into a magnetic, easily recoverable form. Have been shown the advantages of a two-stage magnetic separation, in a strong field and with polygradient separators. Concentrating this material through magnetic methods makes it possible to increase the iron content to 20-21 %. Magnetic concentration of the deeper layers materials results in iron- containing concentrates with a total iron content of up to 58% whit a phosphorus content of about 2%. Reducing the phosphorus content is possible by blending upper and lower layers of the material at a ratio of 1:1. Thus, it is possible to process the material accumulated in the slime storage facility using the ROMELT process combined with preliminary three-stage magnetic concentration and disposal of tailings with the permissible iron content of less than 5%. Wastes from smelting will be slags suitable for use in construction industry. The Kamysh-Burun material preparation for ROMELT process should be performed applying gravity-washing with separation of the -0.4 mm fraction and additional blending for dephosphorizing the concentrate. The smelting will produce pig iron and slag."

Jan 1, 2018

By S. Kelebek, A. Khan, J. Caceres, Beckett. D.

"Pyrrhotite in nickel-copper sulphide ores is nickeliferrous and intimately associated with pentlandite. Due to its low nickel content, pyrrhotite rejection is an important step in processing of these ores. In our laboratory, various reagents were investigated for their effect on pyrrhotite depression. Among these, a combined use of triethylenetetramine and sodium metabisulphite was found highly promising.The work described in this paper has two parts. The firs part is a summary of results obtained on process samples from the pyrrhotite rejection circuit at Falconbridge’s Strathcona Mill. The second part involves simulation of the circuit performance based on the reagent conditions used in the experimental work and the concept of split factors.Two sets of reagent conditions were evaluated. The first one involved baseline conditions, simulating the current reagent performance. The second one simulated the impact these new reagents can have on the circuit performance.The simulation results obtained under baseline conditions indicated a good agreement with the current circuit performance. The simulation results suggest that at practically the same pentlandite recovery and tailing grades, the circuit concentrate nickel grade can be doubled through effective pyrrhotite rejection by these new reagents."

Jan 1, 2003

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

Sep 1, 2012

By S. M. Narbekova, L. V. Semushkina, Zh. A. Aldybayeva, A. A. Mukhanova, D. K. Turysbekov, N. K. Tussupbayev

More complex and heavy-concentrating oxidized and mixed sulphidic oxidized ores are involved in production because of increasing needs for non-ferrous metals. Extraction of the oxidized minerals of non-ferrous metals general in current technologies. The problem is solved to oxidized polymetallic ores the preliminary selective sulphidization is the oxidized surface of the components by means of hydrogen sulphide, sulphide and calcium polysulphide, and also element sulphur, etc. The task was to study details of flotation of the oxidized lead-bearing ore of the Zarechnoe deposit (Kazakhstan) with calcium polysulphide as sulphidation. According to chemical analysis initial samples of the ore contain, %: lead – 4.8; zinc – 0.1; general iron – 0.81. Phase structure lead is characterized by the oxidized form of 57.76-64.5 %; by sulphidic status of 25.5-32.64 %, and by plyumboyarosite of 19.0-23.02 %. Researches on flotation beneficiation of oxidized lead-bearing ore are conducted. Sulfidizator gave to grinding, on flotations there were used the collector (butyl xanthogenate of sodium) and frother (T-80). Experiments were conducted with a new sulphidizator – calcium polysulphide. The latter is obtained by interaction of CaO calcium oxide with sulfur and water at high temperature with component mass the following ratio S:CaO:H2O =10:5:85. Extent of mill initial ore was optimised (88.42 % of - 0.071 mm grade). For the optimum mill mode the studies on matching of an optimized consumption of sulphidizators are performed. On basic technology with application of sulphide of sodium as a sulphidizator at optimum mill, the lead concentrate with the content of lead 64.4 % at extraction of 86.6 % was received. The content of cerussite in a concentrate was about 14.9 % at extraction of 82.1 %. With application of calcium polysulphide as a sulphidizator the lead concentrate with the content of lead 68.0 % at extraction of 88.0 % was received. The content of cerussite has made 16.0 % at extraction of 85.0 %. Application of a new sulphidizator allows to increase extraction of lead in a concentrate from 64.4% to 68% and cerussite extraction – from 82.1 % to 85.0 %. The possibility of application of calcium polysulphide as a sulphidizator instead of conventional sulphurous sodium was established.

Jan 1, 2018

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

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

Jan 1, 2014

By M. Barsky, E. Barsky

It is generally accepted that the optimal flow velocity for gravity classification is the final particle settling velocity or, which is the same, the particle spin velocity. However, we have found differences between the settling velocities and the spin velocities of the very same particles. Here we present the results of experiments in a cascading gravitational classifier comprising eleven stages. The feed material was introduced at each stage successively, from the first to the eleventh. The optimal flow velocity was determined for each case. From these experiments it follows that the value of the optimal flow velocity depends to a greater extent on the point of introduction of the feed (i.e., on the design of the device) than it does on the interaction between the solid phase and the flow.

Jan 1, 2003

By A. Vidyadhar

Printed circuit boards (PCBs) constitute a majorpart of electrical and electronic equipment containing valuable metals such as Cu, Ni, Au, Ag, Pd, Fe, Sn, Pb, etc. The concentration of base and precious metals in PCBs is observed to be several times more than those in their respective ores. Therefore, the recycling of PCBs is necessary to recover the valuable materials. However, the ultrafine particles pose a big challenge. Removal of them is suggested prior to processing by flotation. Froth flotation is observed to be a promising technique for rejecting the plastics from the comminution product. In the present work, enrichment of ground 1.0 mm PCB powder was investigated through flotation route by varying the operating variables such as frother dosage, pulp density, air flow rate and rotational speed of the impeller. The liberation studies indicate that excellent liberation of metal values is achieved from the non-metallic constituents at -1.0 mm size and below. The particulate system is quite rich in metal value with about 23% total metal content. The non-metallic constituents such as plastics are observed to possess strong hydrophobicity while the metal particles in the pulverized mass are hydrophilic in nature. Froth flotation kinetics is studied in depth with a view to facilitate high rejection of plastics and identify optimum operating conditions for the same. A number of experiments are performed to establish the influence of the operating variables on flotation performance. The conditions for achieving higher yield of the metal-rich fraction with respect to a specified grade are discussed in the light of the experimental results. It is found that the requirement of reagents is negligible which could be a very important factor from the commercial stand point. Single-stage flotation increases the metal content from 23% to over 32% with a mass yield of around 75% and over 90% recovery of metal values. 32% of the materials in the feed could be effectively rejected in the float fraction losing less than 4% metal values. The dependence of kinetics on the process variables is also discussed. It was concluded that a high rotor speed helps in efficiently rejecting the plastics. Frother should be added to help stabilize the froth and enhance the kinetics. A moderate air flow is required while pulp density must be kept low for efficient pre-concentration. It is established that the entire -1.0 mm comminution product could be treated by flotation for generating a pre-concentrate. Keywords: printed circuit boards, recycling, froth flotation, kinetics, recovery

Sep 1, 2012

By E. Kusaka, M. EL Salmawy

"The electrokinetic behavior of EL Maghara coal has been investigated in the presence of nonyl and octyl nonionic surfactants containing polyethoxy of different sizes with trade names of Triton-X100 (TX-100), Triton-X 705 (TX-705), Triton WR-1339 (TWR-1339) and Brij 58. The electrokinetic behavior was determined by zeta potential measurements. Zeta potential measurements were conducted by Laser-Doppler electrophoresis technique using zetamaster. The results show that the zeta potential of Egyptian coal varies with pH value, the hydrocarbon chain and number of ethoxy groups in the hydrophilic group of the surfactant. It was found that the polymeric surfactants reduce the negative values of zeta potential of coal. The shift of PZC of coal upon deashing appears to be closely relevant to silicate mineral removal."

Jan 1, 2014

By P. Matias

The study involved two projects that will produce copper concentrate and may have their operations integrated. Transport by trucks, conveyor belt, and the equipment called RopeCon were evaluated as options for the ore transport between the sites. A rainforest stretch, that Vale S.A. demands to preserve, is situated between the projects. It was suggested to further review RopeCon because it presents minimal interactions with the environment. To design the preliminary system layout, a research was done on the rainforest cover height data, but the information found was scarce and imprecise. For preserving the local forest, the use of this data would conduct to the oversizing of RopeCon towers and, consequently, overestimation of its costs. For the acquisition of reliable information, a method was developed using local data acquired through airborne laser survey. The volume between the digital surface and terrain models was calculated to generate a surface representing the forest cover height. It was revealed, by plotting the points obtained from this surface, that 85% of the vegetation is below 25 m. This height is suggested to review the preliminary project with a more accurate design of the equipment towers, thus reducing its costs estimates. The method permits the precise location of the large trees, making easier to find a pathway, for linear-shaped facilities, that would minimize the creation of vegetation-free zones as a measure to protect equipments against falling trees and forest fires. Keywords: forest cover, mining projects facilities, digital models, airborne laser survey

Sep 1, 2012

By F. K. Ikumapayi

An approach to environmental sustainability and improved process economy, in sulphide minerals production is recycling of process water in flotation of complex sulphide ores, although the chemistry of process water may be a critical issue to flotation efficiency. The influence of major components of calcium and sulphate ions in process water on xanthate collector adsorption and flotation response using pure chalcopyrite, galena and sphalerite minerals were investigated by Hallimond flotation, zeta-potential measurement, FTIR and XPS spectroscopy studies, while bench scale flotation tests were also carried out using complex sulphide ores. The impact of the species in flotation was comprehended using deionised water, tap water, process water and simulated water containing equivalent amount of calcium and sulphate species in process water. Hallimond flotation results showed a decrease of chalcopyrite and galena recovery in process water and also in the presence of calcium and sulphate ions in both deionised and process waters, whereas sphalerite does not respond to flotation. The adsorption of calcium and metal ions but not sulphate ions on the minerals is evidenced by zeta-potential data. FTIR and XPS studies revealed the presence of surface oxidized sulfoxy species and surface calcium carbonates on chalcopyrite in the presence of process water and water containing calcium ions, surface oxidized sulfoxy and carbonate species on galena in the presence of deionised water, process water and water containing calcium and sulphate ions, all at flotation pH 10.5, and these surface species influenced xanthate adsorption. The presence of surface oxidized sulfoxy and carbonate species at the sphalerite flotation pH 11.5 were seen in the presence of deionised water, process water and water containing calcium and sulphate ions, but the surface species does not influence xanthate adsorption. Bench scale flotation using two different complex sulphide ores showed that chalcopyrite, galena and sphalerite recoveries are better in process water than tap water. The studies showed that the process water can be recycled in flotation with no detrimental effect on grade and recovery of sulphide minerals. Keywords: sulphide minerals, flotation, zeta-potential, FTIR, XPS

Sep 1, 2012

By J. Yianatos

The regrinding stage is necessary to achieve the particle size (liberation) for the final upgrade in the cleaning flotation process. At present, vertical mills (Vertimill) are employed at industrial scale, using small steel balls as grinding media. The regrinding process characterization requires mixing regime knowledge along with the mean residence time of the liquid and solid phases. The residence time distribution (RTD) and the actual mean residence time estimations from theoretical relationships are complex due to the presence of three components (solid mineral, balls and water), mass transport characteristics and segregation affecting solid holdup. Therefore, the experimental RTD determination is suitable for industrial mills. In this work, RTD measurements were performed in two different industrial flotation plants. Both plants operate with Vertimills of 1000-1500 HP in the regrinding process with different circuit arrangements according to the plant treatment capacity. The RTD was measured using the radioactive tracer technique. This technique allows for non-invasive tracer detection where the procedure consists of selecting a liquid (Br82 in aqueous solution) or solid (irradiated solid) tracer for RTD data acquisition in real time. The RTD was modelled by the Large and Small Tanks in Series (LSTS) model, which allowed the characterization of the different mixing regimes for both liquid and solids in the regrinding mills. Effective mean residence times in the range 2-11 min were obtained with RTD shapes similar to those obtained in tumbling ball mills. It was found that residence times of liquid and fine particles were similar, which is consistent with results observed in other mixing systems, e.g. mechanical flotation cells. Also, the mean residence times estimated from the RTD allowed the mass balance adjustment to be validated from metallurgical sampling data. RTD estimations are useful for better understanding the Vertimill behaviour and to identify the actual breakage rates. Keywords: vertical mill, residence time distribution, particle size, modeling

Sep 1, 2012

By S. J. Neethling

"This paper examines the impact of several key variables on the performance of flotation cells. The study makes use of a simulator for the froth phase of the flotation vessel, as the main purpose of this study is to separate the impact of different variables from one another. In experiments it is very difficult to separate the effect of variables from one another as a change in a singe input parameter, such as air rate, invariably also affects many of the other key variables, such as bubble size and bubble bursting rates.In order to achieve this a base case simulation that is typical of many of the rougher cells found in sulphide flotation circuits was carried out. Four key variables, air rate, incoming bubble size, surface bubble size and the fraction of the bubbles flowing out of the cell unburst, were varied in order to obtain the system.s response.The most significant results were obtained when the bubble size was varied. An increase in the incoming bubble size has the effect of reducing the pulp zone recovery, but increases the froth zone recovery. If it were possible to vary the incoming bubble size without influencing the overflowing bubble size, the impact of the change would be unclear because of these two competing effects. The effect of a change in the overflowing bubble size has an opposite effect on the froth recovery, with larger bubbles resulting in lower froth recoveries.A final interesting result is that all the effects examined affect the process to a similar degree, which implies that none of these variables can be disregarded, either in industry or in further research."

Jan 1, 2003

By Liu Xiaodong, Jiang Meifen, Kou Xiaoyan, Lu Xianjun

"Two low grade calcium bentonite (Ca-bentonite) from Hejia (HJ) and Boteng (BT), Shandong, China, mainly contain montmorillonite and quartz, and slightly feldspar, calcite, and montmorillonite content are 32.37% and 55.90%, respectively. In order to examine the feasibility to produce activated clay to meet the standard of Chinese chemical industry (HG/T2569-2007), three types of industrial production process are proposed according to the screen test. The first process is grinding the raw bentonite to -200 mesh and then activating. The second is to activate the raw bentonite directly, and then classify the production to -200 mesh. The last one is to activate the raw bentonite after purification. Three parameters, which are sulphuric acid strength, ratio of liquid to solid and treatment time, are optimized via orthogonal test for every process. The preliminary test indicates that it is feasible to produce acid bentonite of difference quality. For the sample of HJ, milling the raw to -200 mesh first and activating afterwards, the parameter is at 9% acid strength, the ratio of liquid to solid 2, the time 4 hours, temperature 90°C, the products can meet the qualified grade for animal and vegetable oils refining. For the sample of BT, milling the raw to -200 mesh first and activating afterwards, the parameter is at 15% acid strength, the ratio of liquid to solid 2.5, the time 2 hours, temperature 90°C, the products can meet the qualified grade of high decolorization for petroleum products refining. For the sample of BT, adopting purification to -325 mesh by stirring and rubbing at first, and then activating at 9% acidity, under the conditions that the ratio of liquid to solid is 2, the time is 2 hours and the temperature is 90°C, the products can meet the first grade of high activation degree for petroleum products refining."

Jan 1, 2014

By Reza M. Rahman

The froth phase plays an important role in the flotation and upgrading of the final concentrate. At present, the techniques that have been used in the mineral industry to measure the froth recovery have limitations. The aim of this paper is to investigate the performance of the froth in an industrial flotation cell. A unique device has been developed which is capable of decoupling the froth zone from the pulp zone. It consists of two concentric tubes. The inner tube acts as a dropback collection chamber or catcher. The particles that detach from the froth phase fall directly into the catcher and are collected as froth dropback. The froth recovery measurements were carried out in a rougher bank of a copper concentrator treating sulphide minerals. The dropback device is designed so that it can be immersed into a plant size flotation cell and plant froth recovery measurements can be taken at any given time. During the experiments, the bubbles laden with valuable mineral particles entered the device from the plant flotation cell, subsequently rising to form a froth layer at the top of the device. The particles that detached from the froth zone were collected in the dropback collection chamber whereas the concentrate sample was collected by a launder. By drying, sizing and chemical analysis of the concentrate and dropback samples, the froth recovery, was estimated on the basis of the valuable component. The effect of air rate on froth recovery was also investigated. Keywords: flotation, froth, dropback, froth recovery

Sep 1, 2012

By S. Leroy

Grinding is the most energy consuming step in mineral processing, however, it is still poorly understood and its efficiency is difficult to monitor. Mill performances are traditionally measured at the mill discharge, on the basis of a mass percentage passing a certain sieve. This approach does not reflect what happens to the fines through milling. On another hand, if many macroscopic parameters, such as the percent solids of the pulp, its viscosity, or its mean residence time in the mill are known to be related to grinding efficiency, the link between these parameters and particle size distribution (PSD) is still unclear. Here, a methodology is established to understand the effect of ball mill operating conditions on particle size and shape distribution in the grinding process. The study focuses on the particular case of an Upper Group 2 (UG-2) platinum ore concentrator. Due to the very low grade of sulphide and PGM (Platinum group metals) in UG-2, grinding mechanisms are entirely control by gangue minerals. These two minerals have contrasted properties and PGM content and these properties immediately affect the entire recovery process. So, the grinding of both gangue minerals is studied on a size by size basis. At particle scale, three parameters are addressed: particle size, particle composition and particle shape. The specific behavior of chromite and silicate grains is demonstrated. As a conclusion, this study gives some clues to operate ball mills in a more efficient manner regarding chromite issue in the processing of UG-2. Keywords: ball milling control, UG-2, chromite, fine particle, particle shape, segregation

Sep 1, 2012

By A. V. Faiberg

Currently, the interest to conditioning of cyanide-bearing recycled solutions with high-grade of base metals, in particular, copper has significantly grown. This can be explained by two major factors. Firstly, there are strict regulations on cyanides amount in tails and higher cost of detoxification processes. This problem can be easily solved during the ore treatment using the conditioning of solutions and water recycling. Secondly, the presence of other metals minerals which react with cyanide during gold leaching resulting in high cyanide consumption. The presence of copper is a major problem. Many copper minerals react with cyanide with the formation of cyanide complexes-Cu(CN)2-, Cu(CN)32- and Cu(CN)43-. One mole of copper consumes 2-4 moles of free cyanide. Availability of cyanide-soluble copper minerals has deleterious effect on gold leaching (gold recovery, cyanide consumption, tails detoxification, efficiency of gold elution and electrowinning). The precipitation of copper from the solutions improves gold recovery by CIL. The removal and recycling of cyanide allowed the reduction of cyanide addition to the process. The recycling of spent cyanide-bearing solutions based on the removal of copper and the return of cyanide bound in complex compounds for leaching process is an optimal way of solving the problem of gold-bearing copper ores treatment. By using the technology, an adverse impact onto gold recovery process is eliminated and at the same time reducing the cyanide consumption. Today, processes and methods of high-grade copper leaching solutions conditioning before and after gold recovery have been developed, which combine two ways of copper recovery: as a saleable product and cyanide to recover gold and reduce total operating costs. They are Cutech, MNR, SART, Cyanomet, AuGment, Vitrokele, Hannah and Cyanisorb. At the same time, these technologies have not been studied enough and described in literature. Data on parameters and operating conditions as well as introduction is not available. The conditioning technology of recycled solutions developed by IRGIREDMET JSC allows copper to be recovered as high quality concentrate and at the same time free cyanide regeneration in the recycled solution which is amenable to gold leaching. This technology was commercialized at a mine in Kazakhstan in 2008. The removal and recycling of cyanide does not require additional cyanide and is cost-effective. More cost-efficiency is achievable through selling copper concentrates, closed water cycle without the detoxification of effluents and enhanced gold recovery by CIL and cathode quality.

Sep 1, 2012

By John O. Marsden

"Technology development and innovation has played a crucial role in the copper industry throughout history. The effective development and implementation of new technologies allows mankind to produce copper at progressively lower cost of production at a given feed grade and mineral type. This progress serves to counteract cost inflation/escalation over time and, depending on the magnitude of the cost benefit, allows ore cut off grades to be reduced thereby improving the availability, accessibility and utilization of resources. Two broad categories of technology development and innovation play a key role in this process: (i) incremental improvements and optimization of existing technology provide benefits in the near- to medium-term, and (ii) step change technologies provide value over the longer term. A key issue for potential developers and implementers of new step change technology is how to justify resource allocation and expenditures for such initiatives. This paper examines this issue with reference to specific case studies in the copper industry. INTRODUCTION AND BACKGROUND Technology development and innovation has played a crucial role in the commodity minerals industry throughout history, and none more important than the copper industry. As new, cost-efficient technologies are commercialized, the cost of production in real terms at a given ore feed grade and mineral type decreases. This progress enables ore cut off grades to be reduced allowing lower grade material to be processed economically, which in turn increases the availability and supply of the metal or mineral of interest. Cost inflation/escalation works in the opposite direction, squeezing margins for operations utilizing incumbent technology to process material from a given ore deposit. A prior review by the author considered the issue of technology development and sustainable competitive advantage. This review concluded that first movers and fast followers can achieve competitive advantage for periods of 10-15 years through effective deployment of new technologies (Marsden, 2004). A new, more comprehensive study and evaluation was completed in 2014. This paper presents a portion of the results from the 2014 study."

Jan 1, 2018

By Kwan Ho Kim

In this study, free fall tests are conducted with the objective of understanding the breakage properties of waste concrete for the guide of the direction of the better comminution process. Waste concrete of 75mm×53mm is tested at various drop heights with or without heat pretreatment. Test results indicate that a block of waste concrete does not break by free fall, but eventually breaks by repeated free falls. The number of free falls required for complete separation of large aggregates from the cement mortar considerably reduced when the samples were heated at 400?. Therefore, a great energy saving can be obtained when heating the waste concrete prior to crushing. The energy saving was found to be dependent on the height of free fall but in all cases, was still large enough to offset the additional energy required for heating. Also, it was found that the heat pre-treatment gives an additional advantage in that it produced cleaner recycled aggregates. This may result from easier detachment of cement mortar from the aggregates due to weakening of bonding forces by heating. By analyzing the size distribution of the crushed product, a free-fall breakage model is developed using the particle breakage ratio and the breakage pattern. The developed model can be used to predict and to analyze waste concrete processing circuit.

Jan 1, 2014

By Mohammad Noaparast, Hamed Haghi

"In this research work, it was attempted to reduce the iron contents (Fe2O3) in relatively low grade silica sample which was obtained from Shenin silica mine, using magnetic separation. The prepared sample contained 93.75%SiO2, 0.44%Fe2O3, 2.78%Al2O3, and 0.08%TiO2 and the Fe2O3 had to be reduced to less than 0.08%. Mineralogical studies and XRD analysis indicated the presence of Hematite, Pyrite, and Goethite as the major iron impurity phases. Mineralogical studies of polished and thin sections demonstrated that the degree of freedom of silica is more than 80% in -300+212 ?m size fraction. After primary stages including crushing, milling, scrubbing and de-sliming, various methods such as low intensity magnetic separation (LIMS), high gradient magnetic separation (WHGMS) and dry high intensity magnetic separation (DHIMS) were applied to find out that which process would be appropriate for the elimination of iron impurities from this sample. Different experiments were carried out on the product of wet rod mill and scrubber. However the magnetic field intensity of WHGMS was increased down to 2Tesla, and there was not any acceptable result in reduction of Fe2O3 content lower than 0.08%. Thereafter the product of wet rod mill and scrubber were dried and reported to DHIMS. In these experiments, maximum magnetic field intensity was 1.4Tesla. In this case very significant results were achieved, and Fe2O3 content of final product were reduced to less than 0.05% which meets glass manufacturing standards."

Jan 1, 2014

By Francisco Vicuña, Miguel Becerra, Gonzalo Montes-Atenas

"The optimization of the copper recovery from smelting slag by a process of comminution / flotation has been a widely discussed topic during the last few decades. Several studies covering the characterization of different types of slag, the impact of the slag cooling rate and of different flotation operating conditions on the flotation performance have been widely published. Currently, at El Soldado mine site (Angloamerican), there are large copper bearing slag deposits that can be a profitable alternative for the concentrator with capacity to treat this material. Under this process schema, the optimization of the flotation process to obtain better copper recoveries is a relevant topic of study. The aim of this study is to assess the efficiency of a sulfidation stage of a slag and to correlate that process with copper recoveries results in a flotation circuit. This article presents a detailed examination at both laboratory and plant scales of the slag sulfidation on the recovery of copper in the flotation process. The experimental methodology involves studying the effect of the sulfidation agent dosage, the pH, the particle size distribution of the flotation feed and the conditioning time. Furthermore, the performance of the sulfidation stage was analysed through the XPS analytical technique. It was observed that the sulfidation of particles in floatable size fractions correlated well with the increase in copper recovery. The results indicated that it is possible to increase between 2 to 4 points the total copper recovery with a sulfidation process implemented appropriately."

Jan 1, 2014