Microbial Beneficiation of Salem Iron Ore Using Penicillium purpurogenum

High alumina and silica content in the iron ore affects coke rate, reducibility, and productivity in a blast furnace. Iron ore is being beneficiated all around the world to meet the quality requirement of iron and steel industries. Choosing a beneficiation treatment depends on the nature of the gangue present and its association with the ore structure. The advanced physicochemical methods used for the beneficiation of iron ore are generally unfriendly to the environment. Biobeneficiation is considered to be ecofriendly, promising, and revolutionary solutions to these problems. A characterization study of Salem iron ore indicates that the major iron-bearing minerals are hematite, magnetite, and goethite. Samples on average contains (pct) Fe₂O₃-84.40, Fe (total)-59.02, Al₂O₃-7.18, and SiO₂-7.53. Penicillium purpurogenum (MTCC 7356) was used for the experiment. It removed 35.22 pct alumina and 39.41 pct silica in 30 days in a shake flask at 10 pct pulp density, 308 K (35 °C), and 150 rpm. In a bioreactor experiment at 2 kg scale using the same organism, it removed 23.33 pct alumina and 30.54 pct silica in 30 days at 300 rpm agitation and 2 to 3 l/min aeration. Alumina and silica dissolution follow the shrinking core model for both shake flask and bioreactor experiments.     

Recovery Behavior of Rare Earth from Bayan Obo Complex Iron Ore

A novel process is presented for recovering rare earth from Bayan Obo complex iron ore. The iron ore was reduced and melting separated to produce iron nugget and rare-earth-rich slag. In order to investigate the influence of cooling rate on mineral components, especially the enrichment behavior of RE-containing mineral, the slag was remelted at 1673 K (1400 °C) and the liquid slags were cooled using three types of cooling conditions, water quenching, air cooling, and furnace cooling. Subsequently, the slags were leached by hydrochloric acid to evaluate the relations between leaching efficiency of rare earth and cooling conditions. The results indicated that the slags under different cooling conditions mainly contained fluorite, cefluosil, and cuspidine. The rare-earth mineral is more fully crystallized when the cooling rate of the liquid slag was decreased. The proportion of Ce (III) to Ce (IV) increases with the increase of heating time and decrease of cooling rate. It has been found that the influence of cooling rate on the leaching rate of the rare earth is slight. From water quenching to furnace cooling, the leaching rate of rare earth increases from 97.00 pct to 99.48 pct. After being filtered, filtrate can be used to produce rare-earth chloride. Leached residue, with CaF₂ of 64.45 pct and ThO₂ of 0.05 pct, can be used to recover CaF₂ and extract nuclear source material.     

Microbial Dissolution of Zn-Pb Sulfide Minerals Using Mesophilic Iron and Sulfur-Oxidizing Acidophiles

The purpose of this study is to test the feasibility of using mixed culture of iron and sulfur-oxidizing bacteria for the dissolution of metals from high-grade zinc and lead sulfide ore. Considering that the roll crusher could reduce the ore size to less than 2 mm, this size fraction was selected in order to study the possibility of removing mill circuit. Effects of parameters such as pulp density, initial pH, Fe²⁺, oxidation–reduction potential (ORP), and pH fluctuations were investigated, as well. The maximum Zn dissolution was achieved under the conditions of initial pH 2, initial 75 g/L FeSO₄ · 7H₂O, and pulp density of 50 g/L. The results indicated that under the optimum conditions, about 68.8% of zinc was leached during 24 days of bacterial leaching treatment. The lead recoveries were low (about 1%), because of precipitation of Pb as lead arsenate chloride. Furthermore, the surface studies by using SEM images showed that during chemical leaching the ore dissolution starts from surface discontinuities, but in bacterial leaching all surface becomes involved. In addition, in another process the ore was leached separately with sulfuric acid and sodium hydroxide, and then final results were compared to the bacterial leaching tests in order to find the optimum hydrometallurgical method to extract zinc and lead from these ores.     

Catalytic-Oxidative Leaching of Low-Grade Complex Zinc Ore by Cu (II) Ions Produced from Copper Ore in Ammonia-Ammonium Sulfate Solution

The catalytic-oxidative leaching of a mixed ore, which consists of low-grade oxide copper ore and oxide zinc ore containing ZnS, was investigated in ammonia-ammonium sulfate solution. The effect of the main parameters, such as mass ratio of copper ore to zinc ore, liquid-to-solid ratio, concentration of lixivant, leaching time, and temperature, was studied. The optimal leaching conditions with a maximum extraction of Cu 92.6?pct and Zn 85.5?pct were determined as follows: the mass ratio of copper ore to zinc ore 4/10?g/g, temperature 323.15?K (50?°C), leaching time 6?hours, stirring speed 500?r/min, liquid-to-solid ratio 3.6/1?cm³/g, concentration of lixivant including ammonia 2.0?mol/dm³, ammonium sulfate 1.0?mol/dm³, and ammonium persulfate 0.3?mol/dm³. It was found that ZnS in the oxide zinc ore could be extracted with Cu(II) ion, which was produced from copper ore and was used as the catalyst in the presence of ammonium persulfate.     

Oxidative Pressure Leaching of Silver from Flotation Concentrates with Ammonium Thiocyanate Solution

The thermodynamics and technologies of the selective pressure leaching of silver from flotation concentrates were investigated in an ammonium thiocyanate medium. Thermodynamic analyses, which include silver solubility in NH₄SCN solution and Eh-pH diagrams of the Me-MeS-NH₄SCN-H₂O system at 25 °C, were discussed. The effects of several factors, such as temperature, leaching time, oxidant, pH value, flotation concentrates concentration, surfactant concentration, and so on, on the extraction percentages of silver and zinc were investigated. The following optimal leaching conditions were obtained: NH₄SCN concentration 1.5 M, lignin concentration 0.5 g/L, Fe³⁺ concentration 2 g/L, flotation concentrates addition 200 g/L, and oxygen pressure 1.2 MPa at 130 °C for 3 hours. Under these optimum conditions, the average extraction percentage of silver exceeded 94 pct, whereas the average extraction percentage of zinc was less than 3 pct. Only 7 pct of ammonium thiocyanate was consumed after 4 cycles, which indicated that ammonium thiocyanate hardly was oxidized under these oxidative pressure leaching conditions.     

Leaching of Chalcopyrite Concentrate with Hydrogen Peroxide and Sulfuric Acid in an Autoclave System

In this study, leaching of chalcopyrite concentrate was investigated in an autoclave system using hydrogen peroxide and sulfuric acid. By decomposition of hydrogen peroxide, the active oxygen formed can provide both high oxidation potential and high pressure in a closed vessel for leaching. Preliminary studies showed that hydrogen peroxide can be used as an oxidant instead of oxygen gas in the autoclave. Central composite design (CCD) was used to examine the effects of the experimental parameters on the copper and iron extraction as a response. The proposed model equation using CCD showed good agreement with experimental data, the correlation coefficients R ² for copper and iron being 0.84 and 0.86, respectively. The optimum conditions to obtain the main goal of maximum copper and minimum iron extraction from chalcopyrite were determined as to be sulfuric acid concentration of 2.5 M, hydrogen peroxide concentration of 2.3 M, leaching time of 24 minutes, chalcopyrite amount of 3.17 g (in 50-mL solution), stirring speed of 630 rpm, and leaching temperature of 351 K (78 °C). Under the optimum condition, 76 pct of copper and 9 pct of iron were extracted from chalcopyrite concentrate. Extraction yield results of metals indicate that selective leaching of chalcopyrite can be achieved using hydrogen peroxide and sulfuric acid in an autoclave system.     

Extraction of Ni (II) from Spent Hydrodesulfurization HDS Catalyst Through Leaching and Electroless Precipitation of Ni(OH)2

The extraction of nickel (II) from a spent hydro-desulfurization catalyst containing 11.6 pct Ni was carried out through sulfuric acid leaching. Variations of parameters such as the concentration of acid, temperature, and time, were studied and optimized. Nickel hydroxide was precipitated from the leach liquor via neutralization with 1 M sodium hydroxide up to pH 12 in three different methods: normal neutralization precipitation, and then neutralization precipitation followed by aging at 353 K (80 °C) for 4 hours and neutralization of the leach liquor with 10 pct (v/v) of 0.1 N sodium lauryl sulfate. X-ray diffraction (XRD) and transmission electron microscopy (TEM) microanalysis shows a difference in crystallinity with the method of precipitation. The nickel hydroxide contains Cu(II), Co(II), Zn(II), and Mn(II) as trace impurities. The discharge capacities of the precipitated nickel hydroxides were 120 mAhg^?1, 140.72 mAhg^?1, and 145.2 mAhg^–1 for aged sample, sample without surfactant, and with surfactant respectively.     

Investigation into Oxygen-Enriched Bottom-Blown Stibnite and Direct Reduction

The direct oxidation of stibnite (Sb₂S₃) using a gas mixture of nitrogen-oxygen was investigated in a pilot plant. Steady-state pilot operation of 5 and 10 t/d was normally observed during the pilot test of 100 days, and a cleaning experiment of high-antimony molten slag from oxygen-enriched bottom-blown was tested by direct reduction in a laboratory-scale electric furnace. Autogenous smelting was achieved without adding any other fuel, which guaranteed the feasibility and advantage of oxygen-enriched bottom-blown stibnite. Through analysis and calculation, the sulfur dioxide concentration in offgas was more than 8 pct, which meets the requirement for the preparation of sulfuric acid. In the reduction experiment, the effects of added CaO, the ratio of coal (ω = actual weight of coal/theoretical weight of coal), and the slag type on the reduction procedure were considered. The residual slag obtained after reduction averaged less than 1 g/ton Au and less than 1 wt pct Sb. The metal phase contained iron less than 3 wt pct, and the recoveries of Au in the metal phase were more than 98 pct. This process shows significant environmental and economic benefits compared with previous processes.     

Stress Corrosion Cracking in Al-Zn-Mg-Cu Aluminum Alloys in Saline Environments

Stress corrosion cracking of Al-Zn-Mg-Cu (AA7xxx) aluminum alloys exposed to saline environments at temperatures ranging from 293 K to 353 K (20 °C to 80 °C) has been reviewed with particular attention to the influences of alloy composition and temper, and bulk and local environmental conditions. Stress corrosion crack (SCC) growth rates at room temperature for peak- and over-aged tempers in saline environments are minimized for Al-Zn-Mg-Cu alloys containing less than ~8 wt pct Zn when Zn/Mg ratios are ranging from 2 to 3, excess magnesium levels are less than 1 wt pct, and copper content is either less than ~0.2 wt pct or ranging from 1.3 to 2 wt pct. A minimum chloride ion concentration of ~0.01 M is required for crack growth rates to exceed those in distilled water, which insures that the local solution pH in crack-tip regions can be maintained at less than 4. Crack growth rates in saline solution without other additions gradually increase with bulk chloride ion concentrations up to around 0.6 M NaCl, whereas in solutions with sufficiently low dichromate (or chromate), inhibitor additions are insensitive to the bulk chloride concentration and are typically at least double those observed without the additions. DCB specimens, fatigue pre-cracked in air before immersion in a saline environment, show an initial period with no detectible crack growth, followed by crack growth at the distilled water rate, and then transition to a higher crack growth rate typical of region 2 crack growth in the saline environment. Time spent in each stage depends on the type of pre-crack (“pop-in” vs fatigue), applied stress intensity factor, alloy chemistry, bulk environment, and, if applied, the external polarization. Apparent activation energies (E _a) for SCC growth in Al-Zn-Mg-Cu alloys exposed to 0.6 M NaCl over the temperatures ranging from 293 K to 353 K (20 °C to 80 °C) for under-, peak-, and over-aged low-copper-containing alloys (<0.2 wt pct) are typically ranging from 80 to 85 kJ/mol, whereas for high-copper-containing alloys (>~0.8 wt pct), they are typically ranging from 20 to 40 kJ/mol for under- and peak-aged alloys, and based on limited data, around 85 kJ/mol for over-aged tempers. This means that crack propagation in saline environments is most likely to occur by a hydrogen-related process for low-copper-containing Al-Zn-Mg-Cu alloys in under-, peak- and over-aged tempers, and for high-copper alloys in under- and peak-aged tempers. For over-aged high-copper-containing alloys, cracking is most probably under anodic dissolution control. Future stress corrosion studies should focus on understanding the factors that control crack initiation, and insuring that the next generation of higher performance Al-Zn-Mg-Cu alloys has similar longer crack initiation times and crack propagation rates to those of the incumbent alloys in an over-aged condition where crack rates are less than 1 mm/month at a high stress intensity factor.     

Interfacial Reaction Between CaO-SiO2-MgO-Al2O3 Flux and Fe-xMn-yAl (x = 10 and 20 mass pct, y = 1, 3, and 6 mass pct) Steel at 1873 K (1600 °C)

This study investigated the interfacial reaction kinetics and related phenomena between CaO-SiO₂-MgO-Al₂O₃ flux and Fe-xMn-yAl (x = 10 and 20 mass pct, y = 1, 3, and 6 mass pct) steel, which simulates transformation-induced plasticity (TRIP) and twinning-induced plasticity (TWIP) steels at 1873 K (1600 °C). It also examines the effect of changes in the composition of the steel and slag phases on the interfacial reaction rate and the reaction mechanisms. The content of Al and Si in the 1 mass pct Al-containing steel was found to change rapidly within the first 15 minutes of the reaction, but then it remained relatively constant. The content of Al and Si in the 3 to 6 mass pct Al-containing steels, in contrast, changed continuously throughout the entire reaction time. In addition, the content of Mn in the 1 mass pct Al-containing steels initially decreased with increasing time, but the content did not change in the 3 to 6 mass pct Al-containing steels. Furthermore, the mass transfer coefficient of Al, k _Al, in the 1 mass pct Al-containing systems was significantly higher than that in other systems; i.e., the k _Al can be arranged such that 1 mass pct Al systems >> 3 mass pct Al systems ≥ 6 mass pct Al systems. The compositions of the final slags were close to the saturation lines of the [Mg,Mn]Al₂O₄ and MgAl₂O₄ spinels when the slags reacted with 1 mass pct Al and 3 to 6 mass pct Al-containing steels, respectively. These results, which show the effect of Al content on the reaction phenomena, can be explained by the significant increase in the apparent viscosity of the slags that reacted with the 3 to 6 mass pct Al-containing steels. This reaction was likely caused by the precipitation of solid compounds such as MgAl₂O₄ spinel and CaAl₄O₇ grossite at locally alumina-enriched areas in the slag phase. This analysis is in good accordance with the combination of Higbie’s surface renewal model and the Eyring equation.     

Oxidative leaching of an offgrade/complex copper concentrate in chloride lixiviants

Laboratory studies have been conducted on chloride leaching as a possible route for the simultaneous recovery of copper, zinc, and lead from an off grade and complex chalcopyrite concentrate (from Sikkim, India) associated with appreciable amounts of sphalerite, galena, and pyrite. The effects of temperature, concentration, and quantity of ferric chloride, stirring speed, and leaching time on metal dissolution have been investigated. Leaching tests have also been conducted with in-dividual (HC1, NaCl, CuCl₂, FeCl₃) and mixed chlorides (two-, three-, and four-component mix-tures). Results show the possibility of recovering not only 99 pct Cu and 89 pct Zn but also 82 pct Pb and 58 pct elemental S by treatment of the concentrate with 4 M FeCl₃ at 383 K (110 °C) for 7.2 ks (2 hours) employing 25 pct excess FeCl₃ and a stirring speed of 700 rev min⁻¹. Though 64 pct iron of the concentrate is found to dissolve, the pyrite seems to remain unattacked. Kinetic studies indicate that the chalcopyrite, sphalerite, and galena of the concen-trate dissolve simultaneously in the FeCl₃ lixiviant as if each mineral is separately leached, and the Cu and Zn dissolution reactions are under chemical control (linear kinetics). The addition of NaCl to the chloride lixiviants is found to be beneficial only up to a common salt concen-tration of 100 g/l. Leaching of the copper concentrate with CuCl₂ or mixed FeCl₃-CuCl₂-NaCl has not been as effective as its direct leaching with 4 M FeCl₃. N.V. NGOC, formerly Visiting Scientist with the Department of Metallurgical Engineering, Institute of Technology, Banaras Hindu University.     

Evaluation of Isothermal Separating Perovskite Phase from CaO-TiO2-SiO2-Al2O3-MgO Melt by Super Gravity

Perovskite phase was successfully separated from CaO-TiO₂-SiO₂-Al₂O₃-MgO melt by super gravity. Under the hypothesis that the titanium exists in the slag in terms of TiO_2, with the gravity coefficient G = 600, time t = 5 minutes, and temperature T = 1563 K (1290 °C), the mass fraction of TiO₂ in the concentrate is up to 52.94 pct, while that of the tailing is just 5.88 pct. The recovery ratio of Ti in the concentrate is up to 81.28 pct by centrifugal separation.     

Bacterial leaching of low-grade ZnS concentrate using indigenous mesophilic and thermophilic strains

In this study low-grade sphalerite has been treated by the bioleaching process using native cultures of Acidithiobacillus ferrooxidans and Sulfobacillus thermosulfido-oxidans in order to determine the ability of these bacteria to the leaching of zinc. The effects of bacterial strain, pH, temperature, pulp density, iron precipitation, and initial concentration of ferric iron on the zinc leaching were evaluated. The bioleaching experiments were carried out in shake flasks at pH 1.5, 180 rpm, 33 °C and 60 °C for mesophilic and thermophilic bacteria, respectively. Compared with the use of laboratory reference strains or control conditions, zinc recovery from the respective concentrate was greater when native isolates were employed. The experimental data show that the selection of the suitable pH and temperature during the bioleaching processes would be important. The results indicate that the increase in pulp density generates a decrease in the dissolved zinc concentration. The maximum zinc extraction reached was 87% using native thermophile S. thermosulfido-oxidans culture after 30 days.     

Study on the Formation and Characterization of the Intermetallics in Friction Stir Welding of Aluminum Alloy to Coated Steel Sheet Lap Joint

Multimaterial fabrication such as joining of steel and aluminum is currently prominent in a variety of industries. Friction stir welding is a novel solid-state welding process that causes good joint strength between steel and aluminum. However, the phenomenon contributing significant strength at the interface is not yet clear. In the present study, the interface of the friction stir lap-welded aluminum and coated steel sheet having joint strength maximum (71.4 pct of steel base metal) and minimum, respectively, under two parameter combinations, i.e., 1000 rpm 50 mm min^?1 and 500 rpm 100 mm min^?1, was exclusively characterized by X-ray diffraction, transmission electron microscopy (TEM), concentration profile, and elemental mapping by electron-probe microanalysis. A TEM-assisted EDS study identifies the morphologies of large size Al₁₃Fe₄ and small size Fe₃Al-type intermetallic compounds at the interface. The diffusion-induced intermetallic growth (thickness) measured from a backscattered image and concentration profile agreed well with the numerically calculated one. The growth of these two phases at 1000 rpm 50 mm min^?1 is attributed to the slower cooling rate (~3.5 K/s) with higher diffusion time (44 seconds) along the interface in comparison to the same for 500 rpm 100 mm min^?1 with faster cooling rate (~10 K/s) and less diffusion time (13.6 seconds). The formation of thermodynamically stable and hard intermetallic phase Al₁₃Fe₄ at 1000 rpm and travel speed 50 mm min^?1 in amounts higher than 500 rpm and a travel speed of 100 mm min^?1 results in better joint strength, i.e., 71.4 pct, of the steel base metal.     

Phase Transformation and Microstructure of Zn2Ti3O8 Nanocrystallite Powders Prepared Using the Hydrothermal Process

This paper examines the phase transformation and microstructure of Zn₂Ti₃O₈ nanocrystallite powders prepared using the hydrothermal process that includes TiCl₄ and Zn(NO₃)₂·6H₂O as the initial materials. Differential thermal analysis, X-ray diffraction, transmission electron microscopy (TEM), selected area electron diffraction, nanobeam electron diffraction, and high resolution TEM were utilized to characterize the transition behavior of zinc titanate precursor powders after calcination. Nanocrystalline Zn₂Ti₃O₈ powders with a size range of about 5.0 to 8.0 nm were obtained when the precursor powders were calcined at 773 K (500 °C) for 1 hour. When the zinc titanate precursor powders were calcined at 1073 K (800 °C) for 1 hour, the cubic crystal of Zn₂Ti₃O₈ with a _o = 0.8399 ± 0.0003 nm still remained the predominant crystalline phase and the crystallite size increased to 20.0 nm. In addition, ZnTiO₃ phase first appeared because of the 13.8 pct of Zn₂Ti₃O₈ decomposition when the zinc titanate precursor powders were calcined at 1073 K (800 °C) for 1 hour. When the zinc titanate precursor powders were calcined at 1073 K (800 °C) for 9 hours, the Zn₂Ti₃O₈ crystallites grew continuously to 80.0 nm and enhanced the crystallinity. When the precursor powders were calcined at 1273 K (1000 °C) for 1 hour, Zn₂TiO₄ crystallites with a _o = 0.8461 ± 0.0002 nm were the predominant crystalline phase.     

Preparation of High-Grade Titania Slag from Ilmenite-Bearing High Ca and Mg by Vacuum Smelting Method

Ilmenite produced from the Panxi area in China has high impurities such as Ca and Mg. High-grade titanium (Ti) slag can be obtained by the electric arc furnace process, a traditional method of treating ilmenite. Thus, Ti slag prepared from the Panxi ilmenite contains high CaO and MgO, exceeding 5 pct of the slag content. This high CaO and MgO content confers considerable difficulty in producing titania (TiO₂) white using fluidizing chlorination. In this study, a new process named vacuum separation was found to produce high-grade TiO₂ materials. The effects of separation temperature and time on the TiO₂ grade were studied. The high-grade TiO₂ slag, which has 93 pct TiO₂, <0.1 pct MgO, <1.2 pct SiO₂, and <0.5 pct CaO, can be produced at 1823 K (1550 °C) in 45 minutes through the proposed method.     

Effect of Na2O and B2O3 on the Crystallization Behavior of Low Fluorine Mold Fluxes for Casting Medium Carbon Steels

An investigation has been conducted to study the effect of Na₂O and B₂O₃ on the crystallization behavior of low fluorine (F) mold powders for casting medium carbon (MC) steels in this article. The results of this study indicated that B₂O₃ tends to lower the crystallization temperature and increase crystallization incubation time of the low F powders; however, Na₂O plays an opposite role compared with that of B₂O₃. The crystalline phase of Ca₁₁Si₄B₂O₂₂ was formed in Sample D2 [F = 3 pct, Na₂O = 10 pct, B₂O₃ = 8 pct (in wt pct)], which exhibited the most similar crystallization behavior to that of cuspidine, such that Sample D2 showed closest crystallization kinetics to that of a conventional high-F mold slag for casting MC steels. The precipitated crystalline phases for all the samples have been analyzed and discussed in the article.     

Leaching Process Investigation of Secondary Aluminum Dross: The Effect of CO2 on Leaching Process of Salt Cake from Aluminum Remelting Process

For the recycling/disposal of aluminum dross/salt cake from aluminum remelting, aqueous leaching offers an interesting economic process route. One major obstacle is the reaction between the AlN present in the dross and the aqueous phase, which can lead to the emission of NH₃ gas, posing a serious environmental problem. In the current work, a leaching process using CO₂-saturated water is attempted with a view to absorb the ammonia formed in situ. The current results show that at a solid-to-liquid ratio of 1:20 and 3 hours at 291 K (18 °C), the extraction of Na and K from the dross could be kept as high as 95.6 pct and 95.9 pct respectively. At the same time, with continuous CO₂ bubbling, the mass of escaping NH₃ gas decreased from 0.25 mg in pure water down to <0.006 mg, indicating effective absorption of ammonia by carbonized water. Furthermore, the results in the case of the leaching experiments with synthetic AlN show that the introduction of CO₂ causes hindrance to the hydrolysis of AlN. The plausible mechanisms for the observed phenomena are discussed. The concept of the leaching of the salt cake by carbonated water and the consequent retention of AlN in the leach residue opens up a promising route toward an environment-friendly recycling process for the salt cake viz. recovery of the salts, utilization of CO₂, and further processing of the dross residue, toward the synthesis of AlON from the leach residues.     

Vaporization Studies from Slag Surfaces Using a Thin Film Technique

The investigations of vanadium vaporization from CaO-SiO₂-FeO-V₂O₅ thin film slags were conducted using the single hot thermocouple technique (SHTT) with air as the oxidizing atmosphere. The slag samples were analyzed after the experiments by SEM/EDX. The vanadium content was found to decrease as a function of time. The loss of vanadium from the slag film after 30 minutes of oxidation was approximately 18 pct and after 50 minutes, it was nearly 56 pct. The possible mechanism of vanadium loss would be the surface oxidation of vanadium oxide in the slag, VO _x to V⁵⁺, followed by surface evaporation of V₂O₅, which has a high vapor pressure at the experimental temperature.     

The Effect of CaO on Gas/Slag/Matte/Tridymite Equilibria in Fayalite-Based Copper Smelting Slags at 1473 K (1200 °C) and <Emphasis Type="Italic">P</Emphasis>(SO<Subscript>2</Subscript>) = 0.25 Atm

Fundamental experimental studies have been undertaken to determine the effect of CaO on the equilibria between the gas phase (CO/CO₂/SO₂/Ar) and slag/matte/tridymite phases in the Cu-Fe-O-S-Si-Ca system at 1473 K (1200 °C) and P(SO₂) = 0.25 atm. The experimental methodology developed in the Pyrometallurgy Innovation Centre was used. New experimental data have been obtained for the four-phase equilibria system for fixed concentrations of CaO (up to 4 wt pct) in the slag phase as a function of copper concentration in matte, including the concentrations of dissolved sulfur and copper in slag, and Fe/SiO₂ ratios in slag at tridymite saturation. The new data provided in the present study are of direct relevance to the pyrometallurgical processing of copper and will be used as an input to optimize the thermodynamic database for the copper-containing multi-component multi-phase system.