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By Xiang Lan

A strain of Thiobacillus ferruuxidans No. 8 was isolated from the acidic mine water of an arsenopyrite deposit, South China. It has been found from the experimental results that the growth of bacteria was inhibited by flotation agents. The order of toxic effect of different flotation agents on the growth of Thiobacillus ferrooxidans is ethyl xanthate > butyat ammonium aerofloat > butyl xanthate ? isobutyt xauthate >? flother No. 2 oil. The bioteaching treatment of arsenic and antimony-bearing gold concentrate in Longshan mine, Hunan province, China, was influenced considerably by the presence of flotation agents. The possible causes of the influence of flotation agents were discussed and an improved method was suggested.

Jan 1, 1991

By Anjan Kumari, B. K. Soni, Manis Kumar Jha, A. Vidyadhar, Vinod Kumar

"In the present paper, recovery of valuable metals from the poly-cracked ash obtained by pyrolysis of discarded PCBs has been studied following physical beneficiation and a hydrometallurgical process alternate to smelting. The ash is subjected to comminution for the separation of metallic components from non-metallic fractions by scrubbing, tabling and froth flotation which enriched the metal in the concentrate -97% which was further screened to separate high and low metal concentrate. The finer fraction containing -12% metallic value was leached in acidic solution. Almost all (Cu, Pb, Fe, etc) were leached in 4M HN03 at 90°C in 75 min at a pulp density 100 g/L. The characterization of both ash and residues were made by XRD, EPMA, EDS, etc. to validate the experimental results. The metals from the leach liquor could be recovered/separated using solvent extraction techniques which can further be processed to get pure metal salts/ sheets.IntroductionPrinted circuit boards (PCBs) are an essential component of electrical and electronic equipments (EEEs) as it plays an important role during operation of these devices. The world market of PCBs has reached nearly $60 billion in 2012, with 1.7 percent real growth over 2011 [1]. The diverse technology-driven global market is expected to grow at an average annual rate of 5% by 2014 [2]. The waste EEEs are generated due to replacement of obsolete equipments with modem ones and thus their demand is also increasing rapidly [3]. The discarded e-wastes are being piled up as landfills which affect the environment and cause health threats to living organisms, if not disposed of properly. These e-wastes contain metals some of which are precious, few are hazardous and others as valuable."

Jan 1, 2014

By Krzyszof Borowiec

With the aim of determining the thermodynamic conditions for sulfation of impurity elements, in particular magnesium, in titaniferous slags the equilibrium gas composition for the sulfation reaction: MO + so2 + 1/2 02 = MS04 has been studied by an EMF technique with a solid zro2 + Cao electrolyte. Here M denotes the impurity elements: Mg, Ca and Mn, and the following _oxide systems have been studied: Ti02-MgO-Fe2 o 3 , Ti02-MnO-Mn2 o 3 , and Ti02 -CaO-Si02 , all in the . temperature range 700-1000ºc. Reproducible EMF values were obtained by adding small amounts of -Na2so4 to the oxide-sulfate mixture. From the observed results the Gibbs energy of formation of the phases Mg 2Ti04 , MgTi03 , MgTi 2o 5 , MnTi03 , CaTi03 , and CaTiSio5 , as well as the activity of MgTi05 in the solid solution with Fe2Ti05 have been calculated. The results show that, from a thermodynamic viewpoint, sulfation of these impurities in titaniferous stags i:s favorable- at around and below 800ºC.

Jan 1, 1985

By Kafui Nyavor

The indirect, and the direct contact mechanisms of bacteria oxidation of sulfides were studied using Thiobacillus ferrooxidans and pyrite. The aim was to determine which of the two generally aeeepted mechanisms played a dominant role, if any, during acid mine drainage (AMD) formation from the oxidation of sulfide minerals. The results show that at a pH of 4.5 where T. ferrooxidans are quite active, but where initial chemical oxidation of pyrite to Fe2+ is limited, the bacteria were unable to directly oxidize pyrite to any significant extent. Whereas, in the presence of Fe2+, T. ferrooxidallS assisted oxidation of pyrite by the indirect mechanism was observed and sustained over the pH and Eb range where the bacteria is active. The results lead to the conclusion that the direct contactmechanism of pyrite oxidition by T. ferrooxidans is insignificant. The predominant, and possibly the only, pathway involves the indirect bacteria oxidation of available Fe2+ to Fe3+, which in turn oxidizes FeS2 to generate additional Fe2+ and acidity in a cyclic process.

Jan 1, 1996

By Zuomei Jin

A laboratory study was conducted on the separation of Zn, Pb, Cu and Ag from a finely dissem-inated interlocked silver-rich multimetal sulphidc ore. These are often difficult to separate from each other by flotation. The thermodynamic calculations of Zn-S-O, Cu-S-O systems at different tem¬peratures were presented on predominance diagrams. Based on these diagrams the optimum temperature and gas composition were selected for roasting. The results of separating Zn, Cu from calcine by neutral leaching and acid leaching were found to be satisfactory. The distribution of silver in the roasting and leaching processes were tested. A flow sheet was proposed.

Jan 1, 1993

By M. Sizgoric

Certain types of mineralogical alteration due either to weathering or serpentinization, and the effects these transformations have on the flotation of nickel sulfide ores are discussed. These effects are generally deleterious due to: (a) the poor flotation properties of the alteration minerals (as in the case of tochilinite), (b) the nature of the alteration texture (as in the case of violarite), or (c) .close association of the alteration products with other undesirable minerals (such as nickeliferous smythite). Several examples from INCO Metals Company's mines and mineral prospects are given as illustration.

Jan 1, 1981

By Allan J. MacRae

A copper alloy to be used for coils in cast copper cooling blocks has been identified and samples prepared for metallurgical examination and destructive testing to compare with cast cooling blocks manufactured from copper and nickel-copper alloy pipes. The copper alloy allows for the formation of a metallurgical bond between the pipe and the cast copper during the casting process, which does not normally occur with pure copper or nickel-copper alloy pipes. The alloy's coefficient of thermal expansion is also much closer to that of pure copper compared with nickel-copper alloys, so there is less strain induced at the pipe-cast copper interface due to differential thermal expansion. It is expected that this new pipe coil alloy and associated casting techniques will allow for the production of cast copper cooling blocks that have greatly enhanced service life in high heat load applications compared to existing technologies.

Jan 1, 2002

By Thomas Havlik

It is proposed to realise the direct leaching, during which will be generated elemental sulfur and not sulfate or sulfuric acid as end product, under conditions that sulfur will not cover the leaching surface and Hot hinder the leaching course. The following improvements have been considered: -use of organic sulfur solvents -ultra high-frequency leaching -ultrasonic leaching -use of surface detergents or surfactants in leaching processes -simultaneously crushing and leaching These proposals are under investigation and preliminary results are promising

Jan 1, 1996

By Shuqiang Guo, Dingsheng Tan, Weizhong Ding, Zunyu Hu

"The mixture of hydrochloric acid and magnesium chloride is a good lixiviant for leaching of laterite. The solution can enhance the leaching ability because of the proton activity deviated positively from ideal. Otherwise, the existence of magnesium chloride in the lixiviant affects the dissolution behavior of magnesia, which may be an advantage of the process with less hydrochloric acid consumption. An investigation of magnesia particle dissolved in strong hydrochloric acid solution has been conducted. Influences of concentration of the magnesium chloride and hydrochloric acid and temperature on magnesia dissolution were studied. The dissolution mechanism was also discussed in this paper.IntroductionThere has recently been much interest in the use of hydrochloric acid leaching processes for recovering metal from ores or secondary resources [l-41_ Compared with sulfuric acid, the use of hydrochloric acid as lixiviant offers significant advantages including faster leaching kinetics even at atmospheric conditions, higher leaching rate and greater solubility for most metal chlorides [5-61 .With the gradual decline of the nickel sulfide supply, the extraction of nickel from laterites has become increasingly attractive [7-s1. Nickel may be co=ercially recovered from laterites both by pyrometallurgical and hydrometallurgical processes. The drawbacks of pyrometallurgical processes include the requirement of high grade ores, substantial energy input and loss of cobalt bound within ores. The extraction of Ni from lower grade nickel laterites ores is suited by hydrometallurgical routes, which play an increasingly important role with depletion of higher grade deposits."

Jan 1, 2014

By Ian G. H. du Preez

It is illustrated how differences in the basic chemistry of individual metal ions can be exploited to develop new or improved separating agents. Factors like coordination number preferred by the specific metal ion, the degree of coordinate saturation reached by the metal ion and its relation to the phase transferability of the complex species, kinetic factors as opposed to thermodynamic factors controlling the success of the separation, the nature and type of the donor atom employed in a separating agent in relation to the preferences of the metal ion, the significance of stereochemical factors and the hydrophobicity of the media are some of the important factors which are discussed and applied for the development of pa1ladiwn(TI) and nickel(TI) specific separating agents, respectively.

Jan 1, 1999

By Jose Hector Figueroa P.

Mexicana de Cananea, S.A de C.V. is a major copper producer mining a large deposit of the disseminated porphyry type considered one of the largest copper resources of the world with 2.7 billion tons of 0.60% copper for the mill plus 4.1 billion tons of averaging 0.26% copper for leaching. Mining of low grade sulfide leach ore, has been an important part of the overall production of the open pit mine which started in 1944. Before this time all the mining in Cananea was by underground methods. The leaching process in Cananea began on the 1920's with in situ operations in the underground mine and small dumps up to the 1950's. Large scale leaching operations began in 1960 and the copper solutions produced were treated in precipitation plants until 1,980 when the first SX/EW Plant was put in operation in Cananea. Leaching action is essentially ferric-bacterial, with no fiesh acid added to the leaching solutions. Traditionally, 5540% copper extraction is recovered fiom ROM, 70- 120 meter high leach dumps in 80 months of leaching. Cananea has successfully improved copper recoveries by implementing several actions that include: 1. Stripping of old dumps. 2. Leaching of old dumps through slotted casing. 3. Crushing conveying leach ore to minus 8 inch and stacking it in 30 meter lifts. As result of these actions, copper recovery has been increased' to about 70% of total copper with actions 1 and 2 and up to 85% in half the time by crushing-conveying the leach ore, based on the same solubility index of the run-of-mine ore. This paper describes the above actions that were implemented to enhance the leaching recoveries from. sulfide leach dumps and the economics of such actions.

Jan 1, 1999

By Dong Hoon Han

The simultaneous sulfation by sulhr dioxide and oxygen of CaO and MgO in calcined CMA (calcium magnesium acetate) were investigated. he experiments were carried out in the temperature range from 700 to 830 "C and at the SO2 concentrations of 0.3 to 1 mol %. The oxygen concentration was kept at 5 %. Each sample was calcined completely before the temperature was lowered to the sulfation temperature. According to the ranges of partial pressure of sulhr dioxide and temperature, there exist three regions for the sulfation reaction: region I in which neither oxide is sulfated, region I1 in which only CaO is sulfated, and region 111 in which both oxides are sulfated. It was shown that the MgO in calcined CMA is sulfated at a rate comparable to that of CaO by forming a mixed sulfate phase. Thus, unlike MgO contained in other desulhrization materials, the MgO content in calcined CMA can be fully utilized for flue gas desulhrization.

Jan 1, 1999

By Dai Xi

In the present work, the viscosities of molten FenO-MgO-Si02 and FenO-MgO-CaO-Si02 semi-synthetic slags for nickel flash smelting were measured in the temperature range from 1473 to 1723K by a rotational viscometer. The ratio of Fe to Si02 was fixed at 1.2, calcium and magnesium oxide contents varied in the range from 2 to 8 mass% and 9 to 12 mass%, respectively. The results showed that silicate anions became smaller by increasing CaO content, which results in a decrease of the viscosity of the slag. In the case of MgO addition the viscosity behavior was complicated. At MgO content below 11 mass%, the viscosity increased with the increase of MgO at all temperatures tested. However, when the MgO content was above II mass%, the viscosity decreased slightly for FenO-MgO-CaO-Si02 system. At higher MgO contents, low-viscosity slags can be obtained by adding CaO. At a given composition, the viscosity decreased with increasing temperature. The higher the temperature, the more MgO can be added before saturation. The effect of Fe304 on the viscosity was quite significant. The viscosity of slag increased by 300mPa's at l573K when Fe304 content varied from 3 to 9 mass%. The viscosity values are presented as the functions of temperature and composition. A comparison between FenO-MgO-Si02 and FenO-MgO-CaO-Si02 systems is given in this study.

Jan 1, 2003

By Patrick R. Taylor

Titanium nitride has been synthesized in a thermal plasma reactor from ilmenite ore concentrate using methane and ammonia as the reactive gases. The product has been characterized by x-ray diffraction and SEM. Results show that a high concentration of methane favors the formation of iron, titanium nitride and titanium carbonitride. The formation of titanium carbonitride can be minimized by controlling the concentration of methane. The formation of TiN proceeds by the progressive reduction of the oxides of titanium to titanium, which is converted to titanium nitride.

Jan 1, 1998

By Robin D. Rogers

Room temperature ionic liquids (RTIL) can replace certain traditional organic solvents in liquid/liquid extraction of metal ions from aqueous systems. The very low vapor pressure and non-flammable nature of RTIL warrant consideration of these neoteric solvents for new processing technologies which will reduce pollution, Increase safety, and lower costs. Metal ion separations in RTIL, however, will typically require the use of an extractant and research is needed to identify suitable selective extractants which have appreciable solubility in the RTIL and low solubility In the aqueous phase. Choice of extractant should also take into account the toxicity, expense, and disposal of the ligand.

Jan 1, 1999

By W. Petruk

The behaviour of galena, chalcopyrite and tetrahedrite during floration of a fine-grained base metal ore from New Brunswick was investigated by image analysis study of the mill concentrates and tailings. The minerals were analyzed to determine the grades and recoveries as related to free and unliberated mineral grains and their size distributions. The results show that only 37 wt % of the galena was liberated but 88 wt % of the free galena was recovered, indicating good mill performance for the lead. The lead recovery was 65 wt %, with the main loss due to unliberated galena. Most of the unliberated galena lost to mill tailings is so fine-grained that galena liberation and Pb recovery would not improve significantly by re-grinding. Chalcopyrite liberation was 56 wt % but only 75 wt % of the free chalcopyrite was recovered. Copper recovery was 45 wt %, with the chalco-pyrite loss in the mill tailings being both as large free grains, and as small and large unliberated grains. It was calculated that copper recoveries could be improved about 8 to 10 wt % by changing the concentrator operations and by regrinding. Concentrator operations hive already been modified and copper recoveries have improved about 5 wt %. About 29 wt % of the silver was lost to tailings mainly as both free and unliberated, large and very small, Silver-bearing tetrahedrite grains. It was calculated that about 25 wt % of the silver in the tailings might be recovered by improved milling conditions and by regrinding.

Jan 1, 1981

By Saskia Duyvesteyn

Ion flotation is a technique applicable to the removal of metal ions from dilute solutions, using ionic surfactants.- While there are a number of physical and chemical parameters that influence the ion flotation process, this paper focuses on two of the important chemical ones, namely frother concentration and ionic strength. The effect of several frothers, including ethanol, iso-propanol and MIBC at concentrations ranging from 0% to 0.4%, was studied for the model Cd-dodecylsulfate and Cu-dodecylsulfate systems. The ionic strength, as well as the effect of monovalent and multivalent counter ions, was also studied in the range of 1@3 - 1 M, at constant pH. The results are discussed in terms of applicability to a range of effluents from minerals and materials processing operations.

Jan 1, 1994

By Yuanyuan Zhang, Patrick J. Masset

"The reduction of FexO in slag is one of the fundamental reactions in iron and steelmaking technology, including the blast furnace process (BF). Experiments on the reduction reaction of iron oxide using CaO-SiO2-Al2O3-FexO-C mixtures were carried out using thermogravimetric analysis (TGA) under reducing conditions at temperatures up to 1100°C. The composition of the primary slag was chosen so that the FexO concentration varied between 0 and 75 wt % and the CaO/SiO2 ratio being kept constant to 0.90. From the TGA curves recorded isothermally at different temperatures the kinetic parameters of the iron oxide reduction were determined.IntroductionIn blast furnace the iron ores and coke are charged in layers from the top. The reduction of iron oxide by coke is one of the fundamental reactions in the blast furnace process. Although a great amount of work has been done on the topic of reduction of composites [1, 2, 3], gasification of carbon [4, 5] and reduction of iron oxides by CO [4, 5, 6], the mechanism of reduction in iron oxide - carbon powder mixtures is not fully understood. Fruehan [1] and Rao [2] studied the reaction rate of the reduction of iron oxide by different types of carbon and concluded that the reaction mechanism is controlled by the oxidation of carbon at temperatures up to 1200°C. Rao [2] studied the reaction of hematite with graphite as carbon source up to 1080°C. An activation energy close to 300 kJ/mol was calculated within the temperature range of 950°C to 1087°C. In his study, Fruehan [1] determined the rate of reduction of Fe2O3 and FeO by coal, coke and coconut charcoal as carbon sources within the 900°C to 1200°C temperature range. He found that the reduction of Fe2O3 by carbon occurs in two stages and that the reduction of Fe2O3 to FeO is faster than that of FeO to Fe. This is due to the fact that the CO2/CO equilibrium ratio is higher, and hence, the rate of oxidation of carbon is faster. Fortini and Fruehan [3] have studied the reduction reaction of wustite by graphite and wood charcoal. They found that the wustite reduction can have a significant effect on the overall rate of reduction in composites at high temperatures or in the presence of a large excess of carbon. The calculated activation energy of the reduction reaction by wood charcoal was close to 368 kJ/mol, while the activation energy of the reduction reaction by graphite was 444 kJ/mol."

Jan 1, 2011

By Nolberto Moreno Borjas

This paper outlines the detrimental impact that unblended ore can have on the Ni and Co recoveries and the ore slurry density with real data from the Caron low-pressure ammonia leaching process for the Punta Gorda ore. The author briefly discusses the reason for the decrease in slurry density that the process experiences. The lower metal recoveries in the process that are caused by the unblended ore chemical characteristics and variations are shown with a data base statistical analysis drawn from the results obtained from a pilot test, designed with different ore bypassed from the normal process flow and a similar ore from a similar metallurgical treatment. Keywords: Ni-Co Laterite, Caron Process, Ore Blending, Slurry Density, Metal Recovery

Jan 1, 2004

By Minjun Tang, Jianfeng Zhou, Linjiang Qing, Yufeng Guo, Tao Jiang

"For the comprehensive recovery ofV, Ti, Fe in vanadium-titanium magnetite in the process of pre-reduction-electric furnace smelting, it is expected that vanadium oxide is reduced into molten iron and titanium oxide is available in the slag. Thus the effect of slag basicity, the content of MgO, MnO and CaF 2 of slag content on the separation of slag and molten iron and the comprehensive recovery of V, Ti, Fe were investigated. The results show that increasing slag basicity, MgO, MnO and CaF2 are all beneficial to the separation of slag and molten iron. Higher slag basicity and the addition of CaF2 strengthen the reduction of vanadium oxide, and inhibite the reduction of titanium oxide, which are in favour of the recovery ofV and Ti. However, the addition ofMgO, MnO was not efficient for the comprehensive recovery ofV, Ti, Fe. For the vanadium-titanium magnetite of Panxi in China, the recovery of Fe was over 98%, the recovery ofV in molten iron was over 85% and the Ti02 content in slag was about 43% under the optimal conditions of slag phase type.IntroductionIt is well known that there are abundant vanadium-titanium magnetite resources which are located in Panzhihua of Sichuan Province[!]. At present the vananium-titanium magnetite concentrates are smelted in the blast furnace, which just make use of iron and vanadium, but lots of titanium is discarded in the slag. At the same time, the process has a large number of disadvantages, such as low recovery of vanadium, long route, pollutions, large investment and expenditure of energy[2-3]. In recent years, in order to efficiently utilize the vanadium-titanium magnetite concentrates, several new non-blast furnace processes have been proposed, such as pre-reduction-electric furnace, pre-reduction-grinding separation and so on[4]. Especially pre-reduction-electric furnace is operating well in South Africa and New Zealand,However,Du to the high slag basicity of the slag, the grade ofTi02 of slag is Corresponding author:Minjun Tang.E-mail:450435123@qq.com. Tel:13627429224."

Jan 1, 2013