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By S. J. Liu

The conventional collector for bauxite mineral is oleic acid, which is highly insoluble and must be used in high dosage in bauxite flotation operations, that is the great drawback of oleic acid in use. A series of hydroxamic acid were synthesized, and their collecting ability to certain minerals in bauxite: diaspore, kaolinite were studied. The collecting power of various reagents to diaspore in an incremental sequence were shown as follows: benzyl hydroxamic acid<amylic hydroxamic ac-id<heptylic hydroxamic acid<oleic acid<octylic hydroxamic acid; while the incremental sequence of the collecting power to kaolinite was like this: benzyl hydroxamic acid<amylic hydroxamic ac-id<heptylic hydroxamic acid<octylic hydroxamic acid<oleic acid. It could be seen that octylic hydroxamic acid had the capability of high collectivity and good selectivity, and octylic hydroxamic acid was chosen to use in the following experiments. Reduced zeta potential showed the adsorption of the anionic collector on the mineral. By the adsorptive capability and IR tests, the adsorption of octylic hydroxamic acid on diaspore was determined to be chemical absorption, which was greater than that of oleic acid, as it was physical absorption. Results of floatation experiments on manually mixed minerals indicated that the collectivity and selectivity of octylic hydroxamic acid were better than oleic acid.

Jan 1, 2014

By Peng Qian, Xiao Li, Guoyan Fu, Peiwei Han, Yongliang Wang, Shufeng Ye

"Arsenic removal from high arsenic gold concentrate is conducive to increase the mineral value and realize the mineral comprehensive utilization. This study investigated the alkaline pressure oxidation leaching (APOL) of an arsenical gold concentrate in lower temperature to separate the arsenic into filtrate. The effects of temperature, NaOH concentration, solid-liquid ratio, reaction pressure and stirring speed on the leaching efficiency were investigated, and the characteristics of the gold concentrate and leaching residues were evaluated using XRD and SEM analysis methods. The obtained results showed that by using lower temperature (<60?) it is possible to avoid the production of elemental sulfur during the leaching process, which is benefit for the oxidation of arsenopyrite; When the reaction was carried out under appropriate conditions, more than 98% of the arsenic will be extracted into the filtrate as the form of arsenate or thioarsenate. INTRODUCTION In recent years, there has been considerable interest in the recovery of gold from refractory arsenical gold ore. The arsenic in this kind of ore usually exists as the form of arsenopyrite, and the gold surrounded by arsenopyrite exists as the form of fine particles, which will give many difficulties to metallurgical work. When the content of arsenic in the gold ore is more than 0.5%, it will seriously penalize the smelting efficiency (Ruiz et al., 2014). So pretreatment of the refractory gold ore is necessary to obtain the largest economic value and environmental benefits (Balaz et al., 2000). The oxidation pretreatment processes mainly include roasting oxidation, pressure oxidation, chemical treatment and biological treatment (Ann et al., 1992; Kou and Chen, 2012; Tongamp et al., 2010; Fernandez et al., 1995; Lin and Zheng, 1996). Among them, the roasting oxidation and pressure oxidation are the more ordinary processes in industry. The roasting oxidation process is the mature and adaptable technology, but the production of waste gas containing sulfur and arsenic will pollute the environment. The pressure oxidation method possesses certain metallurgical and environmental advantages, and can efficiently decompose the mineral. According to the different solution media, the pretreatment can be divided into alkaline and acidic pressure leaching pretreatments (Papangelakis and Demopoulos, 1990; Hiskey and Sanchez, 1995). The acidic pressure oxidation method has been used in most smelting plants for high arsenic gold ore pretreatment, while the alkaline pressure oxidation is still not widely applied. Compared with the acidic pressure oxidation, the alkaline pressure oxidation can be realized under lower temperatures."

Jan 1, 2018

By D. I. Verrelli

Within the flotation community a belief has developed that some particle shapes are more ?floatable? than others. This is usually attributed to an influence of particle shape or roughness on the induction period required to achieve attachment between the particles and the air bubbles in the pulp. Up to now, such measurements have not been able to readily isolate the effect on individual flotation subprocesses. In contrast, our experimental apparatus, the CSIRO Milli-Timer, enables us to directly observe the process of particle?bubble interaction and attachment by means of a high-speed video recording, thus providing a direct measure of the induction period for attachment. To assess the influence of particle shape on induction time we used two varieties of methylated borosilicate glass particles ? spheres and angular ?frit? ? in a range of tightly-sized fractions. In doing this, we take care to account for the influence of other factors that could affect the induction time, such as the polar angle of sliding commencement, and approach velocity. These parameters are recorded as variables for each interaction, and corrected for using multiple nonlinear regression. Our results illustrate the importance of particle shape on induction period, with angular particles exhibiting induction periods that were an order of magnitude lower than those of spheres. Furthermore, the induction period was seen to decrease with increasing particle velocity, or kinetic energy on approach, but increased as the trajectory approached the limit of just grazing the bubble. Particle shape in mineral processing is a consequence chiefly of the mineral type and the type of grinding employed. The results presented herein indicate that attention should be paid to the shape of particles obtained from the grinding operation, besides particle size. Keywords: attachment, bubbles, flotation, induction period, particles, shape effects, surface roughness

Sep 1, 2012

By Z. G. Song, C. Y. Sun, X. R. Zhang, Y. Yu

In skarn type scheelite deposit, there are calcium minerals with floatability similar to scheelite, such as calcite, fluorite, etc., which are difficult to separate. Heating cleaning method developed by Petrov has been widely applied in processing rough concentrate of scheelite for nearly half a century, but there exists a general lack of full understanding of the interaction between floatation reagent and mineral particles surfaces. The mechanism on separation of scheelite and calcium minerals in heating cleaning process was investigated by mineral flotation experiments, XPS analysis, as well as solution chemistry and crystal chemistry calculations. The results showed that the sequence of minerals floatability using sodium oleate as collector and a large amount of sodium silicate as modifier in heating flotation test was scheelite > calcite > fluorite, which was the same as solution chemistry calculations based on Gibbs free energy vs. pH and also the same as Ca-X intensity order by crystal chemistry analysis. There are two sides about the mechanism of heating cleaning method: one is that the adsorbed collector on calcium minerals surface is desorbed more easily than scheelite, and the other is selective dissolution of calcium ions from the mineral surface because of Ca-X intensity difference in heating and stirring process.

Jan 1, 2014

By D. Johann, E. Zanetti, I. A. S. Brum, P. M. Zancan, A. F. Castro

"This study was done from a sample of coal mined at the Vale-Mozambique mine, located in Moatize district, Tete Province. The aim of this work is to analyze the reagent system in the flotation of coal fines belonging to the Upper Chipanga layer (UC). Among coal processing methods, flotation stands out as one of the most important for the concentration of this material, in particular in the treatment of fine particles. The total feed of the Vale-Mozambique processing plant is 8,000 tph of coal, where 10% of this feed corresponds to the fine fraction that feeds the flotation circuit. The material used in this study had a particle size of 96% smaller than 0.25 mm. The reagents used in the flotation tests were betacol and diesel oil as hydrophobizing agents and MIBC as frother. The range of betacol concentrations in the first test phase was 200 g/t to 500 g/t, and in the second phase 200g /t to 500g /t while diesel oil and MIBC were kept constant at 300 g/t. The immediate analysis followed the Brazilian standards NBR 8289, NBR 8293, NBR 8290; and NBR 8299. The results showed that it is possible, from a feed with ash content around 22.84%, to obtain products with levels below of 10% ash, with a mass recovery around 50%. The recovery of carbonaceous matter was also evaluated and presented positive results. Complementing this study, the effect of H2O recovery was evaluated and it was observed that for the concentrations of betacol the recoveries ranged from 6 to 9%, and for diesel oil plus MIBC were 4 to 7%.INTRODUCTION Considering the increase in population and energy consumption as humanity has been living along the last decades and the consequent demand for non-renewable natural resources, it is clear the challenge of the mineral industry in the search for new technologies and improvement of the existing ones. This challenge is justified by the need for better industrial results and to meet the market specifications, taking into account the reduction of deposits with higher contents and the need for beneficiation of more complex ores, lower levels and lower release granulometry. Usually in these situations, processes already well-established as the gravimetric (Sampaio and Tavares, 2005) have not good results."

Jan 1, 2018

By Polezhaev S. Yu, I. V. Ukraintsev, Boduen A. Ya, Saburbayeva L. Yu

"Comparison studies of different technologies are performed when selecting optimum option for gold ore and concentrate processing. Methods of ultrafine grinding, oxidizing roasting, bacterial leaching and pressure oxidation are used to extract gold from refractory ores and concentrates. Refractory gold flotation concentrate produced from pyrite-arsenopyrite ores of Bestobe deposit is a case study in this paper. The results of the experimental tests of the direct leaching of flotation concentrate and pre-oxidation products using different methods (oxidizing roasting cinder, pressure oxidation and bacterial leaching cakes) are the following: - cyanidation of original flotation concentrate provides recovery of Au at 46.6 %; - oxidizing roasting increases gold recovery from 46.6 % to 85 %; - BIOX increases gold recovery up to 86 %; - preliminary pressure oxidation of flotation concentrate ensures steadily high Au extraction in the follow-up cyanidation at 94.9-96.5 %. Pressure oxidation and bio-oxidation has changed dramatically the strategy of refractory gold ore processing. These solutions provided for development of ore deposits that considered earlier non-commercial. However, refractory ore processing has a complex character in terms of developing process flowsheet and requires large amount of studies to be performed for each new project. Gold extraction is the basic but not the only criteria to technology selection for gold ore and concentrate processing. Opex and capex, ore reserves, regional geography, developed infrastructure and skill personnel should be considered. Process studies and feasibility study of the developed technology shall be provided for each specific gold ore deposit. INTRODUCTION Due to changes in available feedstock and depletion of easy-to-process ores, refractory ores containing finely disseminated gold and silver in sulphides (primarily, in pyrites and arsenopyrites) are more and more involved into processing. Experts say that such ores amount above 40 % of the world’s gold reserves. Conventional technologies of gold extraction from refractory ores include flotation, roasting and cinder cyanidation (Fraser etal., 1991). However, today’s best available practice for refractory gold ore processing are bacterial (BIOX) and pressure (POX) oxidation."

Jan 1, 2018

By Indrajit Paul, Atanu Pal, Kapil Modi, Sanjay Chandra, Chandra Sekhar, Srinivas Dwarapudi

"High alumina content in the Indian iron ore fines (2.5% Al2O3) resulted in poor quality pellets at the recently commissioned 6 MTPA pelletizing plant and a study was under taken to establish suitable pellet chemistry through proper fluxing agents. Limestone fluxed pellets were prepared at varying basicity (0 to 0.8) and pyroxenite fluxed pellets at varying MgO contents (0 to 3%) to study their effect on the metallurgical properties. Fired pellets were tested for their strength, RDI, reducibility and swelling. Microstructural studies, optical and SEM-EDS, were carried out to quantify different phases and their chemistry. Acid pellets exhibited poor quality whereas limestone pellets at 0.8 basicity and pyroxenite pellets at 1.5% MgO exhibited optimum metallurgical quality. Formation of widely distributed silicate melt in limestone pellets and high melting point slag & magnesioferrite in pyroxenite pellets found to improve their quality. However during commercial production of MgO bearing pyroxenite pellets, their room temperature and low temperature properties found to be inferior, but exhibited better high temperature properties, viz. softening-melting behavior. Contrary to this, commercially produced limestone pellets, comprising CaO, exhibited better strength and RDI but showed poor high temperature properties. To yield the benefits of both CaO (for good strength and RDI) and MgO (for low swelling and high softening), a dual flux combination was established by using both limestone and pyroxenite/olivine together that resulted in optimum metallurgical properties. Coke rate at the blast furnaces found to be reduced by 1kg for every1% increase of dual flux pellets in the burden.INTRODUCTION Iron ore agglomerate quality plays a vital role in decreasing the reducing agent consumption and increasing the productivity of blast furnace. In most of the integrated steel works, the burden mix for blast furnace is decided as per the availability of the iron ore agglomerates like sinter and pellets. More attention has been given in recent years to the use of fluxed pellets in blast furnace due to their good strength and improved reducibility, swelling and softening melting characteristics. Generally, quality of pellets is influenced by the nature of ore or concentrate, associated gangue, type and amount of fluxes added and their subsequent treatment to produce pellets. These factors in turn result in the variation of physicochemical properties of the coexisting phases and their distribution during pellet induration. Hence properties of the pellets are largely governed by the form and degree of bonding achieved between ore particles and the stability of these bonding phases during reduction of iron oxides (Panigraphy et al., 1990). It is important to note that conditions and parameters of pelletizing are specific to given ore or concentrate; the present study is undertaken for typical Indian hematite ore fines high in alumina. Except authors earlier work on quality of pellets from Indian iron ore (Dwarapudi et al., 2008, 2010), there is no considerable literature reported on the effect of pellet basicity and MgO content in the basicity range of 0 to 0.8 for Indian hematite ore."

Jan 1, 2018

By A. Gül

This investigation was carried out with low grade Murgul-Cakmakkaya copper ore once used to be dumped away as waste material. Sulphide minerals in the ore sample observed are chalcopyrite, pyrite, sphalerite and to some extent molibdenite; quartz, calcite and dolomite are the gangue minerals. The ore sample assays 0.28% Cu, 2.94% Fe and 3.36% S. Mineralogical study of the ore sample showed that adequate liberation of chalcopyrite for flota-tion was achieved below 100 µm. In flotation tests Aerophine 3418-A was the collector employed, NaOH or Ca(OH)2were the pH modifiers. Deactivators used were sodium meta bi-sulphide, starch, dextrin and lime. MIBC was the frother used throughout the tests. Test results showed that starch alone was the most suitable depressant to give the optimum metal content and recovery val-ues. Activated charcoal was used to adsorb excess collector species in the flotation pulp during cleaning stages After two stages cleaning of the rougher concentrate using starch as the depressant a copper concentrate assaying 21.2% Cu was obtained with 79.5% recovery while the tailings contained 0.01% Cu and metal losses were about 3%

Jan 1, 2014

By Jorge Misle, Daniel Silva

"This paper addresses the need for advanced control technologies and strategies in the mining industry, its potential and actual impact on mining processes and the best practices for their implementation, operation, and long term sustainability. It is well known that advanced process control (APC) can be a very good source for incremental process improvement for the mining industry, yielding to higher production rates or helping maintain consistent production rates despite ore changes. The application and integration of new technology, process knowledge and innovation are key elements in building such a supporting structure. More specifically, and this cannot be emphasized enough, the correct application and assembly of control tools, the in-time presentation of key data to decision makers and the development of relevant statistics and KPI’s (key performance indicators) can and will make a difference not only in terms of production, but also in plant stability, availability and production costs. In deploying such ideas and aiming at the long-term goal of "Operational Excellence", Minera Los Pelambres (MLP) has developed a number of successful control strategies, best practices and technologies, which are not only the application of state-of-the-art automation and computing tools, but rather the assembly of the above presented structuring factors. This work discusses some of the key tools and ideas implemented for SAG & Ball milling and Flotation areas, their performance results and the roadmap taken for the integration of process knowledge and control tools. Also, cases studies are presented and discussed further to not only show the main areas of improvement, but also to highlight the benefits of such strategy that for the most part, deviate from the traditional mining industry control approach. Also, a word is given on the need for innovating in the design of control strategies and leveraging best practices and process knowledge that ultimately, will contribute to long-term reliable operations, improved mineral processing and, although unmeasured, lead to a superior overall performance and productivity of all mining personnel involved."

Jan 1, 2014

By T. X. Hu

Pyrite ore is one of the main raw materials in the production of sulfuric acid. For a long time, the raw material containing the pyrite, which is used to produce sulfuric acid, generally contains S only 32%~45%. A plenty of residue from the sulfuric acid production contains 40%~50% Fe and 1%~2% S. Some of the residue also contains Cu, Pb, Zn and other harmful elements like As, thus it pollutes environment seriously and is solid waste. The research on comprehensive utilization of the residue has been made by many scholars. Through the study, the residue can only be used to pro-duce iron concentrate and the iron oxide red pigment. But because of the complex techniques, the low recovery, the low grade of iron concentrate , high sulfur in iron concentrate, causing second waste residue and the environmental pollution, the residue is mainly used as an accessory material in the production of cement, and is of a low value. Therefore, the iron resource in pyrite has not been used efficiently until now. In the paper, the flotation condition test, the flotation open circuit test and the flotation closed circuit test to produce high-purity pyrite concentrate with the low-grade pyrite ore in oceanic ore deposit have been carried on. Flotation reagent scheme of sulfuric acid used as activator and butyl xanthate as collector is determined. After four times depth cleaning, high-purity pyrite concentrate containing 52.82% S, 44.29% Fe, and only 2.89% impurities has been obtained. The S recovery is 99.13%, and the Fe recovery is 96.00% in the high-purity pyrite concentrate. The oxidizing roasting depth desulfurization experiment of the high-purity pyrite concentrate has been conducted. The results show that the Fe grade in the residue obtained is 68.58%, the content of S is 0.25%, the content of P is 0.0084%, the content of As is 0.0065%. The residue is a kind of high-quality iron-smelting raw material. Based on the above experiments, the industrial experiments on producing high-purity pyrite concentrate with the low-grade pyrite ore, and producing sulfuric acid and iron concentrate with high-purity pyrite concentrate have been done. The results indicate that the high-purity pyrite concentrate with the S grade of 51.24%, the Fe grade of 44.15% has been obtained. The iron concentrate obtained with 65.73% Fe and 0.23% S is a sort of high-quality iron-smelting raw material applied to the production of steel and iron smelting. Through the experimental study, the sulfur and iron resources in the pyrite ore have been recovered and utilized efficiently.

Jan 1, 2014

"The global demand for the copper metal is continuously increasing however, the high-grade copper ore deposits are constantly depleting. As a result, ore bodies with complex mineralogy characteristics and rich in arsenic-copper bearing minerals are increasingly processed. Arsenic-copper sulphides such as enargite can represent a significant amount of arsenic in concentrate, which causes high financial penalties by smelters. On the other hand, arsenic (As) is a highly toxic inorganic pollutant, which causes human health and environmental problems. Recent research studies have demonstrated promising results showing the possibility of separating arsenic-copper sulphides from copper sulphides using controlled pulp potential (Eh) flotation at laboratory scale. However, the selective removal of arsenic-copper minerals in real ore systems is not well understood. Two ore samples were selected from the Tampakan copper-gold deposit located in the Philippines named the HAS (high arsenic sample) and the LAS (low arsenic sample) samples and the mineralogical analyses were performed on both samples. The floatability of enargite in a batch flotation system under controlled pulp potential conditions was investigated for the HAS and the LAS samples. It was observed that it is possible to selectively separate enargite from other copper minerals after reducing the pulp potential to about -200 mV SHE at pH 11 in the LAS sample. However, no separation between enargite and other copper minerals was observed at a reducing potential for the HAS sample. Mineralogical analyses and surface chemistry studies were undertaken for both samples to find out the reasons for different observations. This study has an important implication in sulphide flotation to achieve sustainable development in the utilisation of arsenic-copper mineral resources. INTRODUCTIONAs the demand for copper is increasing, the extraction and development of resources including complex deposits is becoming necessary. These complex deposits often have issues such as: lower grade ores, floatable non-sulphide gangue, very fine-grained ore and therefore poor liberation and significant regrinding, mixed copper mineralogy, complicated galvanic interactions, economic levels of gold and silver, and high levels of impurities such as arsenic. There is currently no established methodology for separating arsenic-bearing copper minerals from other copper sulphide minerals by flotation. The possibility of separating arsenic-copper sulphides from other copper minerals by controlling the potential of the flotation pulp has been shown in previous studies. However most of these studies were conducted on single mineral systems and the selective removal of arsenic-copper minerals in complex ore systems is not well understood. The flotation behaviour of arsenic-bearing minerals within complex ores differs significantly from that of pure mineral systems due to the effects of mineralogical composition, the degree of liberation, and the surface oxidation. The understanding and manipulation of these effects can lead to improvements in the selective flotation recovery of arsenic-bearing species from complex ore systems."

Jan 1, 2018

By D. R. Nagaraj, Raymond Farinato

Water constitutes one of the most valuable resources employed in mineral recovery and processing. Identifying potential problems encountered in mineral processing in waters unusable for drinking or agriculture (including seawater and even hypersaline waters), understanding chemical factor interactions, and optimization of flotation reagents will greatly expand our ability to create more sustainable beneficiation technologies. The current paucity of our knowledgebase regarding high salinity effects on flotation clearly signals a need for a significant research effort. We report here results of a study designed to probe effects of pulp chemistry on metallurgical performance in the flotation of Cu and Cu-Mo ores in saline waters. The results clearly demonstrate that pH (more correctly dosage of pH modifier to target a pH set point) has the most significant effect on Mo recovery when using seawater. This is linked to the formation/precipitation of Mg(OH)2 when pH = 10.3 (in seawater). The type of pH modifier used, NaOH or Ca(OH)2, had no impact on Cu or Mo recovery, indicating that Ca2+ ions were not involved in any reduction in Mo recovery in the pH range of 9-11. Our results in tap water spiked with either Ca2+ or Mg2+ ions in the same pH range (controlled by NaOH), confirmed that Ca2+ ions had no effect on Cu or Mo recovery, but that addition of Mg2+ ions had a severe adverse effect on Mo recovery at pH = 9.5, and to a lesser extent on Cu recovery. The onset of Mg(OH)2 precipitation occurred at a higher pH, and the adverse impact of Mg(OH)2 on Cu recovery (and perhaps Mo recovery as well) was less, in sea water than in tap water spiked with Mg2+. These effects are attributed to the mitigating role of the high electrolyte concentration, notably NaCl.

Jan 1, 2014

By Silvia França, Hudson Couto

"In this paper, an alternative beneficiation process was developed to recover nickel minerals from low-grade weathered ores. The studied sample, 0.3% Ni grade, is essentially composed of nickel sulfides (pyrrhotite and pentlandite); magnetite and silicates (mostly talc and amphibolites) are the primary gangue minerals. The particle size distribution was controlled using a grinding operation to avoid producing an excess of fine particles by overgrinding. The sample underwent a grinding step for the fine degree of liberation. The slimes (-20 ?m) were removed via two desliming stages in hydrocyclones (diameters of 1.5" and 2"); the ground material, which was classified in a range of (-150+20) ?m, was processed by column flotation in three different stages: rougher flotation for magnesium silicates, rougher flotation for sulfides, and cleaner flotation for sulfides, using pilot flotation columns (2" and 6" in diameter and 20 ft high). The goal of using the preliminary magnesium silicate flotation step was to remove hydrophobic minerals (mostly talc) that would float naturally with the sulfides, which would decrease the Ni grade in the concentrate, as well as to reduce the MgO content because this compound is considered harmful in metallurgical processes when a grade greater than 9% is used. The final concentrate obtained in the cleaner flotation step reached up to 2% Ni and 7.0% MgO. The global and flotation recoveries in the process were approximately 30% and 60%, respectively, due to the losses of Ni during the desliming process. Despite this low global recovery, the adopted beneficiation process can be considered an interesting alternative to produce nickel concentrate from tailings or waste materials."

Jan 1, 2014

By S. M. Bradshaw, R. Smith, J. H. van Vreden, C. A. Snyders

"The re-treatment of gold tailing heaps has shown to be a potentially significant revenue stream for both large and small scale gold companies. Limited access to capital and lack of skilled labour remain restrictive especially when dealing with a refractory ore and/or where limited amount of material are available. Alternative techniques such as low temperature pre-leaching options to unlock value are therefore in demand. Several low temperature pre-leaching experiments in acidic and alkaline solutions, followed by cyanidation were performed which ultimately led to the further investigation of a caustic pre-leaching step. It was found that the Au recovery from a refractory ore tailings heap could be significantly improved from ˜ 15 to 20% to greater than 60% with a simple atmospheric NaOH pre-leaching step. It was further considered, that since the NaOH pre-leach liberate the Au associated with refractory sulphide minerals, the overall process may benefit by the inclusion of a froth flotation step to separate refractory sulphides from the free milling ore. Froth flotation on weathered tailing heaps are, however, known to be difficult, ultimately resulting in poor recoveries. A process flow sheet, incorporating a caustic pre-leaching step and a process which includes flotation and caustic pre-leaching of the concentrate, are technically evaluated and ultimately compared to simple cyanide re-processing. INTRODUCTION South Africa consists of a large number of gold tailings dumps. Janse van Rensburg (2016) estimated that there was about 517 dumps in the Gauteng area, covering approximately 320 000 ha of surface area. Recovering value from the dumps has received a great deal of interest but can be problematic due to the dumps often being refractory, preg-robbing and weathered. High recovery of valuable metals (Au and Ag) are possible (Fleming and Brown, 2010) but would most likely require pressure oxidation and subsequent neutralisation which can be very capital intensive demanding a high skill set to operate and maintain. In addition, these flowsheets target high processing tonnages due to the economy of scales at the lower grade feed materials, further increasing the capital requirement."

Jan 1, 2018

By E. Gock

Based on the scientific investigations on mechanochemical synthesis of metal sulphides from metals or metal compounds and sulphur or sulphur compounds, the article describes the industrial production of nanoscale metal sulphides. Special operational conditions such as inert atmosphere, size and type of the grinding balls, grinding time and specific energy input are taken into consideration. A system configuration for the manufacturing of sulphides at a capacity of 100 kg/h is presented.

Jan 1, 2014

By K. Lillkung

The use of secondary sources has become a very attractive option in the production of platinum group metals as the demand for these metals has increased. One of the promising sources are the spent catalysts, which provide an attractive source as they have relatively high metal content compared with the ores where the metal is originally extracted from. The concentration of platinum group metals in catalytic converters is in the range of 0.1-1 mg/g as the concentration in ores is usually less than 10 ppm. Aqueous chloride solutions are claimed to be a cost effective way in which platinum group metals can be brought into solution and concentrated. In this work the effect of different parameters on the dissolution rates of Pt, Pd and Rh were investigated. The parameters investigated were the Cl- concentration, the temperature and applied potential. Chloride concentrations used were 62.5, 90, 125 and 250 g/l. Temperatures were 25, 40, 60, 80 and 94°C. Potentials were 950, 1050, 1150 and 1250 mV vs. SCE for Pt and Rh and 380, 410, 440 and 470 mV vs. SCE for Pd. Dissolution rates were calculated from the current densities determined by potentiostatic measurements. Thin wire electrodes were used in the measurements. As expected, higher temperature, higher potential and higher chloride content generally increasedthe dissolution rate. However, the effect was nonlinear and slightly different for different metals. The results show that strongly oxidizing conditions are needed in order to achieve reasonable dissolution rates. With the highest chloride content and at potential of 1250 mV vs. SCE and 94°C the dissolution rate of Pt was nearly 1 mm/h, whereas in similar conditions Rh achieved a rate of only nearly 100 µm/h. Similarly, with the highest chloride content and at potential of 470 mV vs. SCE at 94 °C the dissolution rate of Pd was nearly 1 mm/h. Keywords: Pt, Pd, Rh, redox potential

Sep 1, 2012

By V. F. Petrov

Cyanide is widely used gold lixiviant for hydrometallurgical processes. The resulting tails contain complex toxic compounds. With reference to this, the protection of the environment should be the major issue. A complete detoxification of cyanides before disposal, a partial detoxification to the allowable concentrations and a disposal of tails at a polygon followed by the detoxification of excess water into the environment are offered as an environmentally friendly way of disposal. However, a complete detoxification of the cyanidation tails does not give the required results due to the presence of toxic compounds, cyanides and thiocyanates in tailings ponds. This is due to adsorption-desorption process occurring between solid and liquid phases. A partial detoxification is used to down cyanide to the allowable concentrations, which are safe to the environment. The storage of tails without detoxification is mainly used at low concentrations of toxic compounds in the disposed tails. With reference to this, water recycling is practiced by Russian plants in order to reduce as much as possible or almost eliminate toxic effluents into the environment. Fresh water is applied if necessary and to make up for water losses, with 8-12% of total water supply. This allows for the stability of water balance and the decrease of consumption of the reagents applied. The conditioning technologies of water recycling followed by regeneration of free cyanide and metals recovery as saleable concentrates are specially developed due to adverse impact of water recycling on gold recovery processes due to high concentration of base metals (copper). These technologies are also profitable. In order to detoxify cyanide in the tails for tailings dam, effective and inexpensive technologies involving a seasonal use of detoxified reagents were developed. For example, a technology with aldehydes use which is applied by a series of operations. Effluents disposed to the environment are detoxified using chlorination followed by almost a complete destruction of toxic cyanides and thiocyanates. Before disposal, chlorinated water is settled to precipitate the resulting hydroxides of metals and complete chemical reactions such as a full decomposition of hypochlorite ions. In order to protect ground water, watertight screens from local water-resistant materials are used. The samples of these materials were previously tested and chosen according to their maximum ability to adsorb toxic compounds in the tailings pond. If local water-resistant materials are not available, whole welded impermeable polymer screens are used. The options mentioned allow the storage of cyanidation tails in accordance with national ecological standards. Keywords: cyaninides, thiocyanates, wastes, environmental safety, recycled water-supply, detoxification, storage

Sep 1, 2012

By A. N. Zagorodnyaya, T. N. Bukurov, Z. S. Abisheva, A. S. Sharipova

"The paper describes the results of stripping of copper production dusts by leaching with aqueous soda, nitric acid solutions and recovery from solutions resulting in the production of lead sulphate, tribasic lead sulfate (TBLS), ammonium perrhenate, zinc, copper-cadmium cake.The combination of two stripping operations under chosen optimal conditions allows the recovery of 98% rhenium, 82% lead, 92 % copper, and 96% zinc and cadmium each. The major amount of rhenium (~ 80%) is recovered at the carbonization stage, and other metals - at the nitric acid leaching stage.Lead sulfate was precipitated by mixing solutions resulting from carbonization and leaching. Tribasic lead sulfate was produced by dissolving lead sulfate in sodium hydroxide solutions. Rhenium was recovered from mother liquor resulting from lead sulfate precipitation by liquid-liquid extraction and zinc was recovered from raffinates. Trialkylamine (TAA) was used as an extractant for rhenium recovery and di-2- ethylhexylphosphoric acid was used for zinc recovery. Optimal liquid-liquid extraction parameters have been defined allowing the recovery of 96% rhenium in crude ammonium perrhenate and 94% zinc in sulfuric acid solution. Based on the experimental data, the flowchart for dust processing to produce salable products - lead, rhenium, zinc salts has been suggested."

Jan 1, 2003

By Susanna G. de Jager, Francis W. Petersen

"The aim of the project was to develop a methodology applicable to different South African coal types to be able to predict with a high degree of certainty what kind of reagent will be suitable for flotation of a certain coal type.The study included 5 coal samples from different collieries and included Highveld seam 3 and 4 coal (as one sample), Waterberg coal (zone 10 -11), Highveld seam 5 upper coal and Witbank seam 2 and seam 4 coal. Seam 3 and 4 (mixture) coal, Waterberg and the seam 5 upper coal were all found to be vitrinite rich coals. The coal characterisation study included petrographic analysis, proximate analysis, float-sink analysis, oxygen-containing functional groups, surface extraction, flotation release analysis and contact angle measurements.The flotation tests showed that the best reagents to float all of the coal samples studied are iso-decanol and dodecene. These reagents produced flotation products that are near to the release curve. Statistical relationship for predicting ash content and yield that can be expected for a certain coal type and selected collector was obtained. From the relationship, the accuracy of the yield is above 95% (p-value below 0.05) with a linear regression of 0.9289. The accuracy of the ash is slightly lower with a linear regression of 0.7841 but also above 95%."

Jan 1, 2003

By E. L. Chanturiya

During the mineral processing of pyrites sulphide ores of some deposits of Russia loss of gold with tailings and pyrite concentrates amounts up to 40%~90%. The problem of selective allocation of gold-bearing pyrite is crucial. There are 4 mineralogy-technological groups of pyrite determined in copper-zinc sulphide ores, based on the increase of contents of gold in pyrite. The gold associated with group 4 pyrite contents of 85% the total mass of gold found in the pyrite. The flotation behaviour of the enriched fractions of pyrite (98%), of the allocated from the ore of Gay deposit, were studied essentially in various conditions ?? and the charge of Butyl xanthogenate of potassium (ButX). Some products of selective pyrite flotation were investigated by methods of scanning electronic microscopy, energy-dispersive and X-ray spectral microanalyses, definition of specific resistance and thermoelectric moving force.

Jan 1, 2014