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By Chuanyao Sun, Longli Liu, Fayu He, Guiye Wu

Utilizing thioglycolic acid (TGA) as a template and using structure-guided design, a molecular modelling approach has been used to design and screen appropriate dispersant for chalcopyrite. A series of organic dispersants based on a TGA substitution with hydroxyl and amino and sulfonic functional groups have been examined on chalcopyrite surface. 20 organic depressors of chalcopyrite have been investigated to study the quantitative structure-activity relationship (QSAR) of chalcopyrite depressors. For the chalcopyrite surface, depressor containing carboxyl and sulfonic group is generally stronger than the depressor containing hydroxyl group. Increasing the number of hydrophilic group can effectively enhance the inhibition. Increasing the number of Solidophilic group does not necessarily lead to enhanced inhibition ability; steric factor is one of the important factors to affect the interaction. Extension alkyl carbon chain length cannot enhance the interaction of depressors. Isomers principle as well as applies to the chalcopyrite depressor molecular design. Flotation verification tests further confirmes the accuracy of the molecular simulation. Through the Computer Aided Molecular Design (CAMD) technology, we can save enormous time and resources to design novel surfactants. This research and application will play a great significance role in sulfide flotation.

Jan 1, 2014

By E. Guerrini, M. Maccagni, J. Nielsen

"Lead is the most important material able to meet the current demand for energy storage. The lead acid battery is reliable, efficient and inexpensive. Continuous developments in battery construction are providing improvements in all aspects of the batteries’ performance. The lead acid battery is both the main lead product and the main raw material for lead production. More than half of the lead produced is from secondary smelters where batteries are the most important feed material. The environmental pressure on lead production has been increasing for many years, and continues to become stronger and stronger. The lead production processes and technologies themselves will require updating to meet the new regulations. Although they are very strict now, they will become even more so and current technologies may not be sufficient. Because recyclability is a strong point for the lead acid battery, the secondary smelters need to reduce the environmental impact as much as possible. In recent years, the secondary smelters underwent impressive changes in this direction, but some more improvements are possible. Electrochemistry can offer a step change improvement in making secondary lead production more environmentally friendly. Engitec has developed an electrochemical process base on a very high current density (high productivity) cell using an inexpensive electrolyte and at a relatively low cost. In this paper, the latest update on the technology is provided. INTRODUCTION The lead acid battery is considered the “walking dead” by some new technologies trying to replace it. It has played and will continue to play a very important role in our lives because it is inexpensive, fully recyclable and the existing lead and the mine reserves guaranty enough material for the extended future. According to the life cycle, the lead acid battery is the cheapest and the most effective system to store energy. Lead is practically a close loop where lead battery production consumes more than 80 % of the produced lead but contemporarily it is the most important raw material for the production of metallic lead. In fact, nowadays, the secondary production has overcome primary production and this trend is expected to continue. Primary lead is the make-up material for the existing battery production and the source of lead for the construction of the new ones."

Jan 1, 2018

By Çisem Tuba, Ismail Girgin

"Comparative atmospheric H2SO4 and H2SO4 + H2O2 leaching experiments were performed using a Turkish lateritic nickel ore containing 1.52% Ni. In the leaching tests; acid concentration, temperature and leaching time were considered as main variables and the dissolution behaviour of the ore were studied by Ni, Fe, Co and Mg analyses. The dissolution values in the H2SO4 leaching experiments were observed to increase up to 60% (w/w) acid concentration and a temperature of 95°C. Increase of H2SO4 concentration above 60% at 95°C caused decreases in all the dissolution values and this was attributed to the formation of Fe2(SO4)3 and Fe2(SO4)3•HSO4•H2O type compounds. At the considered optimum H2SO4 leaching conditions ([H2SO4]= 45%, T=95°C, t=45 min); dissolution values reached for Mg, Ni, Fe and Co were determined as 99.12%, 95.58%, 92.63% and 78.37%, respectively. An activation energy of 30.57 kJ/mole was determined for nickel dissolution from the evaluation of the kinetic data obtained for H2SO4 leaching showing that the dissolution was controlled by external diffusion and chemical reactions. Caro’s acid solutions were prepared at H2SO4/H2O2 mole ratio of 3 for varying initial H2SO4 concentrations and leaching tests were performed at 25°C. Co behaved very differently and a dissolution value of 83.94% was obtained even at 5% acid concentration in 60 minutes reaction time. Complete dissolution of Co was obtained after 30 minutes of reaction time at 60% initial H2SO4 concentration and the corresponding dissolution values for Mg, Ni, and Fe were reached to 96.16%, 94.27% and 89.46%, respectively. Although the H2SO4 + H2O2 leaching experiments were conducted at 25°C, it was realized that the temperature of the leaching medium was increased to 122°C after four minutes of reaction time due to probable exothermic reaction(s) at 60% initial acid concentration."

Jan 1, 2014

By M. Minagawa, G. Granata, C. Tokoro, T. Kato

"In this work, we studied the mechanochemical activation of chalcopyrite by vertically stirred ball milling (VSBM) to enhance the dissolution of copper. By using x-ray diffraction (XRD) and laser diffraction particle size analysis, we investigated the effect of milling and milling conditions on particle size and phase composition of the chalcopyrite concentrate. To elucidate the efficiency of VSBM towards the mechanochemical activation of chalcopyrite, we also conducted leaching experiments in sulfuric acid-ferric sulfate media. Under all investigated milling conditions, the VSBM determined a dramatic decrease of particle size from over 100 µm of P80% up to about 13 µm. Furthermore, the XRD of the ground chalcopyrite specimens highlighted a partial oxidation of the mineral. In leaching, upon grinding by VSBM, the increase of copper extraction ranged from a minimum of +10% to a maximum of +40%, depending on the grinding conditions. In order to elucidate whether the dissolution enhancement was a consequence of the decreased particle size or of an induced mechanochemical reaction, we performed leaching experiments on activated and non-activated chalcopyrite sample having the same particle size. In addition, we determined the activation energy of leaching before and after vertically stirred ball milling. The leaching results on samples with the same particle size highlighted a greater copper extraction and higher dissolution rates upon mechanochemical activation. Whether performing grinding by VSBM or not, the leaching kinetics was found to be controlled by the chemical reaction on the surface of the chalcopyrite particle. By grinding the chalcopyrite concentrate by VSBM, the activation energy of leaching slightly decreased from 66 to 58 kJ/mol. INTRODUCTION Due to the increasing demand of copper, low grade ores and refractory minerals like chalcopyrite are expected to become the main copper source in the next years (Klauber, 2008). As a low-grade sulfide, chalcopyrite shall be processed hydrometallurgically. However, due to the formation of a protective layer of elemental sulfur during the oxidizing leaching, the dissolution of copper is slow and/or incomplete (De Bernardi and Carlesi, 2013). Therefore, the development of new technologies to enhance the extraction of copper from chalcopyrite is highly required."

Jan 1, 2018

By K. C. Corin

It is well known that the grinding circuit plays a crucial role in flotation. The effect of pulp chemistry during grinding is often overlooked. It is believed that the surface properties of the minerals can be adjusted during flotation rather than during grinding. However, recent development in surface analysis techniques (TOF-SIMS, XPS, and AES etc.) have proven that the hydrophilic iron hydroxides formed at the mineral surface during grinding make it difficult for collector adsorption, which may result in poor grades and recoveries of the valuable minerals. This study investigated the effect of the pulp chemistry conditions during milling and the subsequent effects on flotation performance of an Nkomati ore. The results have shown that there was approximately 10 % increase in Ni recovery, when the ore was ground at a dissolved oxygen (DO) concentration of 0 ppm (DO-0) in comparison to no gas control during the grinding stage. However, when the pH was changed from 9 to 11, no gas control resulted in a 20 % decrease in the Ni recovery compared to only 5 % decrease in the Ni recovery at DO-0. A noteworthy observation is that controlling a combination of DO and pH during grinding is of critical importance for better floatability and selectivity for such a system. It also became clear from this study that the presence of oxygen is critical to the consumption of xanthate, which is consistent with the general literature. It has been established that grinding at lower DO levels/redox potentials (Eh) in the presence of xanthate significantly improves the flotation performance of the valuable minerals.

Sep 1, 2012

By K. H. Luo, J. Y. Liu

An experimental study on interparticle breakage of three iron ores was carried out by compressive stressing of particle-bed in a piston-die device. The feeds to the particle-bed stressing were five narrow size fractions in the size range of 0.2~3.2 mm respectively and a wide size fraction of 0/4.0 mm. The pressures applied were in the range between 30 and 360 MPa. The results show that the energy absorption of the particle bed increases linearly with the pressure; the increase of the product fineness with the pressure declines as the pressure goes up; and the product size distributions resulting from interparticle breakage of different narrow size fractions have no self-similarity. Using data obtained from the piston-die compressive stressing test, the suitability of a previously-developed material-characterization model for interparticle breakage was verified for these ores. It is shown that the particle-bed compression test conducted with piston-die device in lab-scale may provide a cost-effective way of studying and modelling size-reduction of ore materials by interparticle breakage.

Jan 1, 2014

By Zhaohui Qiu, Yaguan Ji, Hong Zhong, Qingxia Liu, Yada Liu

"Water is a necessary process medium in mineral processing. For the operations where fresh water is not available, sea water could be used as an alternative processing medium to replace the fresh water if we would have built a solid understanding of the interfacial science and flotation chemistry of sulfide minerals in sea water. In this research, micro flotation of chalcopyrite and pyrite in deionized water and sea water were investigated. For chalcopyrite, the recovery starts to drop from pH 10 in deionized water and from pH 9 in sea water, and the flotation recovery close to 100% before it drops. We can found from the results that chalcopyrite flotation is depressed much at pH 11 in sea water when much white precipitate is found. For pyrite, the flotaion recovery changs a lot in deionized water with 1*10-5 M potasium amyl xanthate (PAX) as collector. It’s relatively high at pH 8~10 compared with that at pH<8 and pH>10. While in sea water, the trend is almost the same as that of chalcopyrite. Zeta potential measurements and chemical equilibrium calculation were used to find out that magnesium hydroxide species coats on the surfaces at high pH in sea water are the main reason for depressing the chalcopyrite and pyrite flotation. Through understanding the interfacial science and the solution chemistry of sea water and its impact on processing efficiency, this research will not only helps reduce the use of fresh water to minimize the disposal of spent water for environment protection, but also enhance the recovery of valuable minerals in froth flotation."

Jan 1, 2014

By M. Amani, A. Akbarinasab, M. Bazmandeh

The fourth iron ore production plant at Golgohar mining and industrial company consists a ball mill and wet medium intensity magnetic separators (3500 Guss). The ball mill product applies directly into MIMS and the MIMS concentrate directly transfers to the low intensity magnet separators (LIMS). The thickener is used for dewatering of MIMS and LIMS tailings. By sampling from streams of ball mill and separators, it was defined that the grade and iron recovery were 60% and 80% respectively while they would be 66% and 90% respectively based on the designing documentation. The average of P80 in the ball mill was 342 micron in a month that is coarser than optimized value. This problem was related to unsuitable operation of grinding. For optimization of grinding efficiency, the amount of charge volume was investigated as a variable. The measurements illustrated 32% charge volume. For increasing grinding efficiency, the charge volume was increased 1% and 1.6% in two stages. The results showed ball mill power draw average reached from 3650 kw to 4120 kw and the average P80 reached to 329 micron. All of parameters calculated about 35 day before and after change. Average of Ball mill feed iron ore grade before and after change was 49.24% and 50.32% respectively. This change caused an increase in efficiency of MIMS, where the enrichment ratio and concentrate grade increased 2% and 1.9% respectively. For increasing the iron grade of the MIMS concentrate; the orifice of the tail flow exit pipe changed from 4 inch to 6 inch and the concentrate grade increased 2%.

Jan 1, 2014

By J. Yianatos

The cleaning circuit operation and its effect on the plant metallurgical performance, was characterized at Los Pelambres concentrator, Chile. Experimental data from columns, scavenger and regrinding stages was evaluated. Each stage was characterized by size fractions, and the interaction between stages (columns, scavenger, regrinding) as well as the effect that circulating loads had. The approach for describing the flotation equipment operation (cells and columns) considered the collection and froth zones separately. This methodology allows for the identification of the impact that particle size distribution has on the actual collection and froth transport processes, independently. In this way the effect of regrinding on flotation recovery and grade can be related to the corresponding operating variables, e.g. pulp flowrate, air flowrate, pulp level and pulp density, among others. The effect of gangue entrainment was considered using an empirical dimensionless correlation. The interaction between the circulating load and the overall performance of cleaning circuit was evaluated. The circulating load was defined as the ratio between the solid flowrate of scavenger concentrate over rougher concentrate. It was found that the overall cleaning flotation circuit (column-scavenger) can operate at a relatively high and constant recovery for a wide range of mineral treatment. This is possible when the scavenger has the capacity for compensating the decrease in column recovery at larger circulating loads. Under the condition of high overall cleaning recovery, the rougher stage closely represents the plant recovery. However, at large mineral treatment conditions, the design characteristics (e.g. concentrate transport channel capacity and transfer boxes) as well as instrumentation, control facilities and operation, became bottle-necked. A simulator was developed using empirical models for each stage (column, scavenger and regrinding) in order to evaluate proper operational strategies for cleaning flotation circuits Keywords: flotation, cleaning, regrinding, Vertimill, simulation.

Sep 1, 2012

By M. Maldonado

A flotation bank is a serial arrangement of cells. How to optimally operate a bank remains a challenge. This article reviews three reported strategies: air profiling, mass-pull (froth velocity) profiling and Peak Air Recovery (PAR) profiling. These are all ways of manipulating the recovery profile down a bank, which may be the property being exploited. Mathematical analysis has shown that a flat cell-by-cell recovery profile maximizes the separation of two floatable minerals for a given target bank recovery when the relative floatability is constant down the bank. Available bank survey data are analyzed with respect to recovery profiling. Possible variations on recovery profile to minimize entrainment are discussed.

Sep 1, 2012

By Takaaki Wajima

The slag was attempted to be converted into the adsorbent for phosphate removal from aqueous solution using the alkali fusion method. The slag was mixed with NaOH powder (NaOH / Slag = 1.6), and then the mixture was heated at 600 ? for 6 h to prepare the precursor. The precursor was stirred at room temperature in distilled water to synthesize the adsorbent, and the removal ability of the obtained adsorbent for phosphate from aqueous solution was investigated. The products, including hydrocalumite and calcite, were synthesized, and the removal efficiencies of the adsorbent for F- and PO43- showed higher than those of raw BF slag, especially removal of PO43- are excellent (99 %), while the removal of Cl- , SO42-, NO3- and Br- was minor. The adsorbent could remove PO43- selectively in the co-existing solution of Cl-, NO3-, F-, SO42- and Br-.The equilibrium adsorption capacity of the adsorbent for PO43- was extrapolated using Langmuir and Freundlich models, and experimental data are found to fit Langmuir than Freundlich. The calculated maximum adsorption capacity of the adsorbent for PO43- is higher than other adsorbent reported previously. The high affinity of the adsorbent for PO43- in aqueous solution was caused by the formation of hydroxyapatite [Ca5(PO4)3(OH)] and brushite [CaHPO4?2H2O]. The prepared adsorbent is expected to be anew inorganic adsorbent for the removal of phosphate ion from wastewater. Keywords: Blast furnace slag, Alkali fusion, Phosphate removal, High selectivity

Sep 1, 2012

By F. Bello, R. Yanez, M. Pinto, T. Errazuriz

"Throughout its operation, Minera Esperanza (ME) will deposit one of the largest tonnages rates of high density tailings in the world with a solids concentration of close to 67% by weight. The tailings presently being deposited have a solids concentration of approximately 62% - 64%, and so cannot be used to directly characterize the flow and deposition characteristics of the higher concentration tailings. To increase the knowledge of the behaviour of their tailings, ME, with the support of Golder Associates, developed a large scale tailings deposition test with a capacity of 30,000 tonnes per day (1,200 m3/h). The test was carried out by depositing tailings into two large cells or paddocks in five stages. Multiple spigots were used to allow different flows to be deposited. At each stage, topographical surveys of the natural terrain and the geometry of the deposited tailings, as well as continuous flow measurement and periodical rheological characterization of the discharged tailings were conducted to create a data set against which beach slope models could be calibrated and then used to predict performance of the higher concentration tailings under different deposition scenarios. This study adopted the beach slope prediction model proposed by A.L. Li (Li, 2011). The model was successfully adjusted to the observations made during this large scale field test by applying the values measured in-situ. This paper presents the application of the model, the results obtained, the findings, and the analysis of possible future scenarios. The model was applied, using the discharge flow rate, the geometrical characteristics of the discharge area, and the rheology with its variation along the beach to estimate the development of the beach formation and the steady state of the beach slopes. Through the use of dimensionless variables, the model was applied to higher solids concentration scenarios, obtaining results that would allow beach slope estimation at different deposition stages in multiple scenarios of varying parameters such as flow, rheology and geometry of discharge area."

Jan 1, 2014

By P. Huang

Chitosan was tested as a depressant in the differential flotation separation of chalcopyrite and galena. The micro-flotation results showed that chitosan selectively depressed chalcopyrite while galena was floated by potassium ethyl xanthate in acidic solutions. With the addition of 0.67 mg/L chitosan (CT) in the slurry of chalcopyrite-galena mixtures at pH 4, the recovery of galena was 95% while that of chalcopyrite was 30%, indicating that chitosan could potentially be used in differential Cu-Pb sulfide separation. Furthermore, the study of chitosan with either low (LCT) or high molecular weight (HCT) in the differential flotation indicated that better separation was achieved by the chitosan with lower molecular weight (LCT). ToF-SIMS was employed to study the distribution of CT on the surface of chalcopyrite-galena mixtures by ion mass spectra analysis and surface imaging, which showed that CT adsorbed heavily on chalcopyrite but barely adsorbed on galena. The positive ion ToF-SIMS spectra provided information on the surface composition of chalcopyrite-galena mixtures after CT treatment. It appears that the selective adsorption of CT on chalcopyrite was originated from the strong interaction between NH3+ and Cu. Keywords: differential flotation, micro-flotation

Sep 1, 2012

By Eberhard Gock

The worldwide implementation of corrosion protection of low alloyed automotive metal sheets by zinc coating has led to supply shortages of zinc free non-alloyed steel scrap in the foundry industry. In a collaboration project, partly financed by the German government, with partners from the automotive industry, zinc metallurgy, scrap trading, foundry industry, and plant construction, a new acid-based pre-dezincification process has been developed, and a pilot plant with throughput of about 100 t/d has been put into operation at the Clausthal Umwelt-Institut GmbH (CUTEC) together with the Clausthal University of Technology in Germany. During the treatment process, acid from a zinc metallurgical plant will be supplied, which can be recycled to the primary zinc extraction process after raising the zinc concentration to more than 130 g/L zinc. This way, the separation of small amounts of iron by precipitation and the electrolytic zinc winning are outsourced and, as a result, the pre-dezincification process runs waste-free. In Germany, the amount of zinc coated ?new scrap? is approx. 3 million t per year with a zinc content of approximately 60,000 t per year. The possible additional incomes for the dezincified steel scrap are in the order of ?40 to ?50 per ton of scrap, the gained zinc content during the acid dezincification defrays the costs of the process, if the throughput is more than 700 t/d. In comparison to conventional dezincification during the electro arc furnace process followed by upgrading in a rotary kiln (Waelz process), the acid dezincification process running at a maximum of 50°C will cut the energy costs by up to 74% and the CO2 emission by up to 88%. In fact, a demonstration plant is being designed right now and will be installed for processing a part of the zinc coated metal scrap from the Volkswagen AG in Wolfsburg, Germany. Keywords: steel scrap, zinc-coated steel, dezincification methods, acid

Sep 1, 2012

By V. A. Vorob'ev, A. N. Larionov, E. N. Ivanov, M. A. Ryazanov

The aim of the research is to increase the recoverable value of poor, difficult-to-digest complex mineral raw materials. In the Republic Belarus, universal centrifugal-shock devices have been created for crushing, grinding and classification of ore and non-metallic mineral raw materials in dry mode, which differ from foreign analogues by a higher level of resource saving (including electricity and materials). Dry-ore preparation technologies have been tested on a pilot scale: fine crushing (replacing grinding in rod mills), grinding (replacing wet grinding in ball mills) and pneumoclassification on different types of ores, - which made it possible to identify the criteria for the suitability and unsuitability of mineral raw materials for their use. The possibility of regulating the upper and lower limit of the size of the finished material is established, the possibility of reducing the yield of slimes (-0,020 mm) to 5 times is determined. A comparative analysis of the contrast degree of separation characteristics of ore a similar granulometric composition, crushed in wet and dry conditions, is performed. The achieved contrast degree of gravity properties after dry grinding - 104,3; after the wet - 60.2. When enriching on a centrifugal concentrator, gold recovery to the concentrate after dry grinding increased by 13.12% while improving the quality of the concentrate from 78.3 g/t gold to 90.75 g/t. When testing the gravity-flotation scheme for the enrichment of poor gold-quartz ore, the increment of gold extraction after dry grinding was 7.6% for gravity and 9.5% for flotation; the total increase in extraction is 17.1%. The possibility of flotation of a larger material after dry grinding has been proved, with an extraction gain of 1.5% compared to the flotation of thinner material after ball milling. The influence of dry technology on the efficiency of the leaching of gold and silver (an extraction gain of 5.5 to 10%) and heap leaching (extraction gain - up to 18.3%) was studied. The positive influence of dry ore preparation technologies on the efficiency of gold and silver heap leaching and heap leaching, thickening and filtration is proved. Lines of dry ore preparation technologies can be used in modular complexes for processing technogenic raw materials.

Jan 1, 2018

By Filippov. L. O.

Metallurgical results for the copper-nickel disseminated ores were improved by using ultrasonic treatment. This has been demonstrated by processing of bubble-pulp flow in the reactor zone of a reactor-separator flotation machine. It can be explained by changes in mineral and size composition of the concentrates. Thus, the selectivity of Cu/Ni flotation for different values of the hydrodynamic parameter increases considerably. Additional effect of the ultrasound treatment has been noticed under conditions when the detachment of the gangue minerals from bubbles can be controlled. Reactor-separator has advantages over other types of flotation cells for this purpose. In fact, the weak diameter of reactors make more efficient the ultrasound effect than in conventional impeller or column flotation machines. A combination of ultrasonic treatment with original organization of gas pulp flux in the reactor-separator allows to obtain significant effects in the flotation attachment/detachment subprocesses, enhancing the process selectivity and/or recovery of substances. Increase of total recovery of copper and nickel into bulk concentrate at constant grade is 8.5% for coarse disseminated ores. For copper ore, copper recovery increased by 10% at unchanged grade of 11.7% and nickel recovery gained 17% at the constant grade of 2%.

Sep 1, 2012

In this study, the effect of different regrinding conditions on the depression of fine pyrite at the cleaner stage was investigated. The regrinding condition was manipulated by different grinding media (stainless steel and mild steel). It was found that, after regrinding with mild steel grinding media, the flotation recovery of pyrite is much lower than with stainless steel grinding media, and it is more difficult to restore by adding more collector. This means, regrinding with mild steel grinding media may depress the pyrite flotation more strongly. The copper activation of pyrite during regrinding was also studied. It was found, that the mild steel grinding media is more beneficial to copper activation than stainless steel media. Additionally, the analysis of pyrite recovery based on the particle size indicated, that the fine particles are more susceptible to different grinding environments than intermediate and coarse particles and therefore are more easily depressed with a much lower flotation recovery.

Sep 1, 2012

By E. V. Manlapig

"Some researchers postulate that the floatability of a particle in the streams of an industrial flotation circuit is a function of only its size and liberation properties. They do not discount the importance of reagent or surface contamination on particle floatability but assume that the reagent coverage or surface contamination of the particle surfaces is essentially uniform for particles of similar size and mineralogical properties when subjected to a particular chemical regime.In this paper, samples collected from batch tests performed on different streams of the Red Dog lead cleaning circuit have been characterised with respect to their size, mineral association and mineral surface exposure. An analysis of this data set indicates that particle recovery rate is significantly affected by mineralogical properties - in particular the degree of valuable mineral surface exposure.However, through the use of a nodal analysis technique developed by the authors, it can be concluded that particles with the same size and mineral composition do not all float with the same rate, but with a distribution of rates. The characteristics of this floatability distribution do not change during the flotation process. It is suspected that the particles within a size mineralogical class in this study have varying degrees of surface oxidation and that this results in varying particle flotation rates. Thus basing particle floatability on size and mineralogical effects, alone, in a flotation model is not always valid."

Jan 1, 2003

By Johannes Käppeler, Alexander Weh, Harald Giese, Klaas van der Wielen

"High voltage breakage is a novel comminution technology that shows considerable potential for various applications in mineral processing and recycling. It relies on highly energetic electrical pulses to induce electrical explosions in rocks, causing the formation of plasma channels and powerful shockwaves that result in fracturing and fragmentation of particles. There has been considerable interest in this technology for its ability to weaken rocks and preferentially liberate target minerals, as previously demonstrated by other authors. Weakening has the potential to significantly reduce energy consumption during comminution, or alternatively, increase throughput of a circuit. Better liberation may yield an improved concentrate grade and/or recovery, reduce fines losses to tailings or increase the grind size required for liberation, thereby also reducing energy requirements. This review aims to increase general awareness of this technology, and promote further research into it by providing a general summary of the theory underlying high voltage breakage, as well as the envisaged applications of this novel technology. A brief history of the technology is presented before progressing onto the physics underlying the electrical and geomechanical aspects of high voltage breakage. Key results regarding weakening and liberation are discussed, and some mention is made of current challenges and planned developments to increase the throughput of high voltage breakage units to the level where it can be integrated into new and existing processing circuits."

Jan 1, 2014

By T. Kalliomäki, B. P. Wilson, M. Lundström, J. Aromaa

"The diffusion rate of cupric ions and limiting current density are important factors that determine the rate of copper mass transport in electrolytes proximal to the electrodes used in electrorefining. As a result, diffusion can critically affect both the rate and morphology of cathodic copper deposition thus it is crucial that it be optimized to improve production. The diffusion coefficient of cupric ion (DCu(II)) in copper electrolytes has been previously researched but the values detailed in the literature show significant variations and moreover arsenic - a common impurity - is generally excluded. Furthermore, in previous studies the models have only been validated with model solutions rather than real industrial electrolytes. In this study, the limiting current density and diffusion coefficient of cupric ion were defined for synthetic electrolytes containing copper (40–60 g/dm3), nickel (0–20 g/dm3), sulfuric acid (130–220 g/dm3) and arsenic (0–45 g/dm3). The cathodic limiting current densities were measured at three temperatures (50, 60, 70 °C) with linear sweep voltammetry using the rotating disc electrode (RDE) as the cathode. By utilizing the limiting current density, models for the DCu(II) were constructed using Levich equation. The validity of the DCu(II) model for technical use was confirmed through their application to industrial electrorefining electrolytes with known element concentrations. Consequently, this study provides improved models for limiting current density and diffusion coefficient of cupric ion that can be utilized for industrial process optimization.INTRODUCTION For an efficient copper electrolysis process, the mass transport of copper ions has to be effective. The main mass transport types of copper ions in copper electrorefining are diffusion and convection while the effect of migration is negligible. In order to maintain good mass transport, it is beneficial to keep both diffusion and convection high to maximize the limiting current density (jlim) (Bard and Faulkner, 2001). jlim is the highest current density that can flow in the cell, thus by maximizing jlim the deposition time can be minimized and the good quality of cathodic copper can be ensured. Thus, it is important to research possibilities that increase and optimize diffusion since slow diffusion in the electrolyte limits the rate of the process (Levich, 1947; Moats et al., 2000; Subbaiah and Das, 1989)."

Jan 1, 2018