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.     

Column Bioleaching of a Low-Grade Silicate Ore of Uranium

The bioleaching of a low-grade Indian uraninite ore (triuranium octoxide, U₃O₈: 0.024%), containing ferro-silicate and magnetite as the major phases, and hematite and pyrite in minor amounts, has been reported. Experiments were carried out in laboratory scale column reactors inoculated with enriched culture of Acidithiobacillus ferrooxidans isolated from the source mine water. The pH effect on uranium recovery was examined with the same amounts of ores in different columns. With the presence of 10.64% Fe in the ore as ferro-silicate, the higher uranium biorecovery of 58.9% was observed with increase in cell count from 6.4 × 10⁷ to 9.7 × 10⁸ cells/mL at pH 1.7 in 40 days as compared to the uranium recovery of 56.8% at pH 1.9 with a corresponding value of 9.4 × 10⁸ cells/mL for 2.5-kg ore in the column. The dissolution of uranium under chemical leaching conditions, however, recorded a lower value of 47.9% in 40 days at room temperature. Recoveries were similar with 6-kg ore when column leaching was carried out at pH 1.7. The bioleaching of uranium from the low-grade ore of Turamdih may be correlated with the iron(II) and iron(III) concentrations, and redox potential values.     

Kinetics of Ammonia Leaching of Multimetal Sulphides

This paper briefly describes the studies carried out on oxidative ammonia leaching of Cu-Zn-Pb multimetal sulphides. Kinetics of zinc and copper dissolution were studied with ? 200 + 300 mesh BSS fraction and 1% solids in the slurry. It is observed that the dissolution of sphalerite proceeds by a phase boundary reaction model and that of copper via diffusion through product layer in the temperature range of 70-100°C. The rate equations for zinc and copper dissolution are given by:

1 ? (1 ? α)^1/3 = k _Zn[NH₃][pO₂]^1/2

1 ? 2/3α ? (1 2/3α )^2/3 = k^Cu[NH₃]²[pO₂]^1/2

where the symbols have the usual meanings.

Activation energies for zinc and copper dissolution reactions are estimated to be 66.5 and 55.4 kJ/mole, respectively. Activation energy values thus obtained are also comparable to those obtained using a differential approach.

The leaching results obtained with 10% solids using a wide range of particle size (? 140 + 500 mesh) indicate that copper dissolution is chemically controlled in ammonia as well as ammonia-ammonium sulphate medium in the temperature range of 115-135°C. However, at lower temperature (?55°C). the leaching reaction follows a diffusion model. Zinc dissolution data show deviations from the shrinking core model due to high extractions in the initial stages.     

Leaching Kinetics of Bauxite in Hydrochloric Acid

The Bayer process is currently used to produce cell-grade alumina from bauxite. However, if the reactive silica content in the bauxite exceeds 7%, losses of caustic and aluminium as sodalite (2 Na₂0.2 Al₂O₃·3 SiO₂·2 H₂O) become economically unacceptable. As Australia possesses large tonnages of bauxite containing more than 7% reactive silica, a process for the production of cell-grade alumina horn these bauxites using hydrochloric acid leaching was developed by the authors.

The process consists of the following steps: calcination of bauxite, leaching of calcined bauxite in hydrochloric acid, filtration of residue, crystallization of A1C1₃.6 H₂O, decomposition of AlCl₃.6 H₂O crystals to produce A1₂O₃, regeneration of hydrochloric acid for recycle into the leaching step.

The kinetic studies of the dissolution showed that the extraction of aluminium is independent of the solid to leachant ratios studied. The acid concentration has a marked effect on leaching kinetics of bauxite and the leaching time decreases substantially as the concentration of acid increases. The order of the reaction is greater than one. The rate equation can be described by the following least squares line of best fit.

r = 4.1 × 10^?5 HCl^0.5 + 1.27 × 10^?5 HCl² where

r = the initial reaction rate, g Al extracted/sec.

Aluminium extraction greater than 90% can be achieved from bauxite in 4 h using 25% HCl while 30% acid is considered the maximum acid strength due to the crystallization of A1C1₃.6 H₂O from leach solutions.

The rate of leaching increases as the temperature of leaching is increased. An Arrhenius plot of the initial reaction rate (g Al extracted/sec) against the inverse absolute leaching temperature produced a straight line. The activation energy for the reaction was determined to be 83.3 kJ/mole with a standard deviation of 2.9 kJ/mole. The correlation between the initial reaction rate and the inverse absolute temperature can be described by the following regression equation:

r = 4.71×l0⁸e^83820/RT

The large activation energy value suggests that the reaction is chemically controlled rather than diffusion controlled. The additives such as NaCl and FeCI₃ do not effect the leaching rate.     

Kinetics of dissolution of synthetic heazlewoodite (Ni3S2) in nitric acid solutions

The dissolution rate of heazlewoodite in nitric acid solution has been determined. The effects of nitric acid concentration, temperature, particle size, stirring intensity and addition of Cu²⁺ ions have been investigated. Solid residues after leaching were examined by SEM, X-ray diffraction and chemical analysis. In the solutions containing less than 2.0 M HNO₃, dissolution was observed to be completely inhibited after 30 min leaching time, and the rate of hydrogen sulphide production was faster than its oxidation to S⁰ and HSO₄^?. In 3 M HNO₃, an abrupt increase in dissolution rate of Ni₃S₂ was found. Two different regions of the dissolution of heazlewoodite were observed below and above 50°C. At temperatures below 50°C, the dissolution rate was very slow, even in 3.0 M HNO₃ solution, and H₂S gas was evolved. Above 50°C, the dissolution rate rapidly increased. Over the temperature interval 60–90°C in 3.0 M HNO₃ dissolution followed a linear rate law, and the activation energy was found to be 42.1 kJ mol^?1. Most of the oxidized sulphide ion was found in the solution as sulphate. The leaching rate was independent of stirring speed. The rate-controlling step of the Ni₃S₂ dissolution is the oxidation of hydrogen sulphide to elemental sulphur or sulphate ions on the Ni₃S₂ surface. Addition of small amounts of Cu²⁺ ions to the nitric acid acted as catalyst for the dissolution of Ni₃S₂. Bubbling air through the leach suspension increased the dissolution rate of Ni₃S₂ in solutions containing less than 2.0 M HNO₃.     

Pyrite oxidation with ozone: stoichiometry and kinetics

One of the most frequent causes of refractoriness in precious metals leaching is their occlusion or fine dissemination into a pyritic matrix. This study experimentally explores the acid leaching of pyrite with ozone, suggests the stoichiometry of the reaction, estimates its activation energy and defines the effect of the main variables on the leaching kinetics. The results of stoichiometry tests allow establishing that one mole of pyrite requires 7.7 moles of ozone to produce one mole of ferric ion and 2 moles of HSO₄? ions. A decrease in the particle size, solution pH and solids’ concentration of the leaching system increases pyrite dissolution. The type of acid (nitric, sulphuric and hydrochloric) does not affect pyrite dissolution rate. Up to 60% of pyrite is dissolved when the optimal experimental conditions are employed (1?g pyrite (?25?µm), 800?mL of 0.18?M of H₂SO₄, 800?rev?min^?1, 1.2?L?min^?1 gas stream O₂/O₃ with 0.079?g O₃?L^?1 and 25°C). The apparent activation energy of the pyrite-ozone reaction is 14.92?kJ?mol^?1, and the absence of a passive layer on the pyrite surface and the linearity of the dissolution profiles suggest that the dissolution kinetics is controlled by the chemical reaction.     

Dissolution behaviour of a beryl ore for optimal industrial beryllium compound production

This study involves the leaching of the beryl ore with sulphuric acid (H₂SO₄) solution for predicting optimal beryllium extraction conditions with the aim of assessing the importance of leachant concentration, reaction temperature and particle size on the extent of dissolution. A kinetic model to represent the effects of these variables on the leaching rate was developed. It was observed that the dissolution of beryl ore increases with increasing H₂SO₄ concentration, temperature, decreasing particle size and solid to liquid ratio. At optimal leaching conditions, 89.3% of the ore was reacted by 1.25?mol/L at 75°C temperature and 120 minutes with moderate stirring, where 1612.0?mg/L Be²⁺, 786.7?mg/L Al³⁺, 98.1?mg/L Fe³⁺ and 63.4?mg/L Ag⁺ were found as major species in the leach liquor. The unleached products constituting about 10.7% were examined by X-ray diffraction (XRD) and found to contain primarily, siliceous compounds such as Xonotlite, Antigorite, Chrysolite and Kaolinite.     

Dissolution kinetics of malachite in ammonia/ammonium carbonate leaching

The leaching of oxide copper ore containing malachite, which is the unique copper mineral in the ore, by aqueous ammonia solution has been studied. The effect of leaching time, ammonium hydroxide, and ammonium carbonate concentration, pH, [NH₃]/[NH₄⁺] ratio, stirring speed, solid/liquid ratio, particle size, and temperature were investigated. The main important parameters in ammonia leaching of malachite ore are determined as leaching time, ammonia/ammonium concentration ratio, pH, solid/liquid ratio, leaching temperature, and particle size. Optimum leaching conditions from malachite ore by ammonia/ammonium carbonate solution are found as ammonia/ammonium carbonate concentrations: 5 M NH₄OH+0.3 M (NH₄)₂CO₃; solid/liquid ratio: 1:10 g/mL; leaching times: 120 min; stirring speed: 300 rpm; leaching temperature: 25 °C; particle size finer than 450 μm. More than 98% of copper was effectively recovered. During the leaching, copper dissolves as in the form of Cu(NH₃)₄⁺² complex ion, whereas gangue minerals do not react with ammonia. It was determined that interface transfer and diffusion across the product layer control the leaching process. The activation energy for dissolution was found to be 15 kJ mol⁻¹.     

Preparation of Industrial Manganese Compound from a Low-Grade Spessartine Ore by Hydrometallurgical Process

The increasing global demands for pure manganese in steel production and manganese compound as dietary additives, fertilizer, pigment, cells and fine chemicals production cannot be over-emphasized. Thus, continuous efforts in developing low cost and eco-friendly route for purifying the manganese ore to meet some defined industrial demands become paramount. Therefore, this study focused on reductive leaching and solvent extraction techniques for the purification of a Nigerian manganese ore containing admixture of spessartine (O_96.00Mn_24.00Al_16.00Si₂₄.₀₀) and quartz (Si_3.00O_6.00). During leaching, parameters such as leachant concentration and reaction temperature on the extent of ore dissolution were examined accordingly for the establishment of extraction conditions. At optimal leaching conditions (1.5 mol/L H₂SO₄?+?0.2 g spent tea, 75 °C), 80.2% of the initial 10 g/L ore reacted within 120 min. The derived dissolution activation energy (Ea) of 35.5 kJ/mol supported the diffusion reaction mechanism. Thus, the leachate at optimal leaching was appropriately treated by alkaline precipitation and solvent extraction techniques using sodium hydroxide and (di-2-ethylhexyl) phosphoric acid (D2EHPA) respectively, to obtain pure manganese solution. The purified solution was further beneficiated to obtain manganese sulphate monohydrate (MnSO₄.H₂O, melting point?=?692.4 °C: 47-304-7403) of high industrial value. The unleached residue (~?19.8%) analyzed by XRD consisted of silicileous impurities (SiO₂) which could serve as an important by-product for some defined industries.     

Removal of Iron From Quartz Ore Using Different Acids: A Laboratory-Scale Reactor Study

The purification of quartz using chemical processes is extremely important for many industries, including the glass, electronic, detergent, ceramics, paint, refractory, and metallurgy industries, as well as for advanced technology products. The purpose of this work was to investigate the removal of iron as an impurity from quartz ores using a chemical leaching method with different reagents.

The iron content of the quartz ore sample was 310 ppm, and the iron was in the form of Fe₂O_3. In the first step, pre-enrichment studies were conducted based on the particle size, and the Fe₂O₃ content of the quartz ore was decreased to 88 ppm. A statistical design of the experiments and an ANOVA (analysis of variance) were performed in the second step to determine the main effects and interactions of the researched factors, which were the concentration of the leaching reagent (H₂SO₄, HCl, H₃PO₄, HClO₄, and NTA [Nitrilotriacetic acid]), solid/liquid ratio, leaching temperature, and leaching time. The highest Fe₂O₃ removal was 86.6%, and a 11.8 ppm Fe₂O₃ content quartz product with a whiteness index (WI) value of 90.6 was obtained after 120 min of treatment at 90°C with a 10% S/L ratio and 1 M H₂SO₄.     

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.     

Extraction of nickel from ferromolybdenum leach residues

Abstract

The extraction of nickel from ferromolybdenum leach residues by sulphation roasting, water leaching and iron removal from subsequent nickel leach solutions was studied. Sulphation roasting and water leaching promote the reaction between sulphuric acid and the residue and decrease the silicon dissolution. Over 90% of Ni was leached. Ferric ions in the solution could be effectively removed as jarosite and ferric hydroxide. The recovery of nickel reached 88·3% under sulphation roasting with the sulphuric acid quality of 1472 kg t^?1 leach residue at 280°C for 4 h followed by iron removal with addition of 0·5 g NaClO₃, 6 g Na₂SO₄ and 10 g CaO/100 mL solution at 95°C for 2·5 h, while the concentration of iron in solution reduced to 0·38 from 56·6 g/L^?1.

On a étudié l’extraction du nickel à partir de résidu de lessivage de ferromolybdène par grillage sulfateur, lessivage à l’eau et enlèvement du fer des solutions obtenues de lessivage de nickel. Le grillage sulfateur et le lessivage à l’eau favorisent la réaction entre l’acide s ulfurique et le résidu et diminuent la dissolution de la silice. On a lessivé plus de 90% du Ni. On pouvait enlever efficacement de la solution les ions ferriques sous forme de jarosite et d’hydroxyde ferrique. La récupération du nickel atteignait 88·3% au moyen du grillage sulfateur, avec 1472 kg d’acide sulfurique par tonne de résidu de lessivage à 280°C pendant 4 h, suivi par l’enlèvement du fer avec l’addition de 0·5 g de NaClO₃, 6 g de Na₂SO₄ et 10 g de CaO par 100 mL de solution à 95°C pendant 2·5 h, la concentration du fer dans la solution étant réduite de 56·6 à 0·38 g L^?1.     

A New Approach to Processing Rutile from Ilmenite Ore Utilizing the Instability of Pseudobrookite

Rutile (TiO₂) is a vital industrial material used in pigments and in many other valuable chemicals. A new production process to synthesize rutile from natural ilmenite ore and therefore overcome the environmental problems associated with conventional rutile extraction processes was developed. Because the simple phase separation of ilmenite (FeTiO₃) into Fe₂O₃ and TiO₂ occurs due to air oxidation, extracting TiO₂ by removing Fe₂O₃ may be possible if pseudobrookite (Fe₂TiO₅), known as a stable compound in the Fe₂O₃-TiO₂ system at higher temperatures, is of unstable phase in the lower-temperature range. In order to clarify the potential of this new approach, the phase stability of pseudobrookite in the lower-temperature range is discussed. The free energy of formation of pseudobrookite from the respective pure oxides was measured at temperatures ranging from 1073 K to 1473 K by the chemical equilibrium technique using Al₂O₃ as the reference oxide. The observed free energy is given as a function of temperature: ?G⁰ = 7715 ? 7.7T (J/mol). The results indicate that pseudobrookite has an unstable phase below 929 K. This has important industrial implications as a new approach to producing synthetic rutile from ilmenite ore by oxidation at low temperatures and acid leaching.     

<Emphasis Type="Italic">In Situ</Emphasis> Observation of Calcium Aluminate Inclusions Dissolution into Steelmaking Slag

The dissolution rate of calcium aluminate inclusions in CaO-SiO₂-Al₂O₃ slags has been studied using confocal scanning laser microscopy (CSLM) at elevated temperatures: 1773 K, 1823 K, and 1873 K (1500 °C, 1550 °C, and 1600 °C). The inclusion particles used in this experimental work were produced in our laboratory and their production technique is explained in detail. Even though the particles had irregular shapes, there was no rotation observed. Further, the total dissolution time decreased with increasing temperature and decreasing SiO₂ content in the slag. The rate limiting steps are discussed in terms of shrinking core models and diffusion into a stagnant fluid model. It is shown that the rate limiting step for dissolution is mass transfer in the slag at 1823 K and 1873 K (1550 °C and 1600 °C). Further investigations are required to determine the dissolution mechanism at 1773 K (1500 °C). The calculated diffusion coefficients were inversely proportional to the slag viscosity and the obtained values for the systems studied ranged between 5.64 × 10^?12 and 5.8 × 10^?10 m²/s.     

The dissolution of chalcopyrite in ferric sulfate and ferric chloride media

The literature on the ferric ion leaching of chalcopyrite has been surveyed to identify those leaching parameters which are well established and to outline areas requiring additional study. New experimental work was undertaken to resolve points still in dispute. It seems well established that chalcopyrite dissolution in either ferric chloride or ferric sulfate media is independent of stirring speeds above those necessary to suspend the particles and of acid concentrations above those required to keep iron in solution. The rates are faster in the chloride system and the activation energy in that medium is about 42 kJ/mol; the activation energy is about 75 kJ/mol in ferric sulfate solutions. It has been confirmed that the rate is directly proportional to the surface area of the chalcopyrite in both chloride and sulfate media. Sulfate concentrations, especially FeSO₄ concentrations, decrease the leaching rate substantially; furthermore, CuSO₄ does not promote leaching in the sulfate system. Chloride additions to sulfate solutions accelerate slightly the dissolution rates at elevated temperatures. It has been confirmed that leaching in the ferric sulfate system is nearly independent of the concentration of Fe³⁺, ka[Fe³⁺]^0.12. In ferric chloride solutions, the ferric concentration dependence is greater and appears to be independent of temperature over the interval 45 to 100 °C.     

Recovery of yttrium from deep-sea mud

Deep-sea mud rich in rare earth yttrium has received lots of attention from the international community as a new resource for Y. A novel process, which mainly includes acid leaching, solvent extraction, and oxalic acid precipitation-roasting, is proposed for recovery of Y from deep-sea mud. A series of experiments were conducted to inspect the impacts of various factors during the process and the optimum conditions were determined. The results show that the Y of deep-sea mud totally exists in apatite minerals which can be decomposed by hydrochloric acid and sulfuric acid solution. The highest leaching efficiency of Y is 94.53% using hydrochloric acid and 84.38% using sulfuric acid under the conditions of H~+concentration 2.0 mol/L, leaching time 60 min, liquid-solid ratio 4:1 and room temperature 25 ℃(only in case of sulfuric acid, when using hydrochloric acid, the leaching temperature should be 60 ℃). Because of the much lower leaching temperature, sulfuric acid leaching is preferred. The counter current extraction and stripping tests were simulated by a cascade centrifugal extraction tank device. Using 10 vol% P204,15 vol% TBP and 75 vol% sulfonated kerosene as extractant, 98.79% Y~(3+) and 42.60% Fe~(3+) are extracted from sulfuric acid leaching liquor(adjusted to pH = 2.0) after seven-stage counter current extraction with O/A ratio of 1:1 at room temperature, while other metals ions such as Al~(3+), Ca~(2+), Mg~(2+)and Mn~(2+) are almost not extracted. The Y~(3+) in loaded organic can be selectively stripped using 50 g/L sulfuric acid solution and the stripping efficiency reaches 99.86% after seven-stage counter current stripping with O/A ratio of 10:1 at room temperature, while only 2.26% co-extracted Fe~(3+) is stripped. The Y~(3+) of loaded strip liquor can be precipitated by oxalic acid to further separate Y~(3+) and Fe~(3+). The precipitation efficiency of Y~(3+) in loaded strip liquor can be 98.56% while Fe~(3+) is not precipitated under the conditions of oxalic acid solution concentration 200 g/L, quality ratio of oxalic acid to Y of 2, and precipitation time 0.5 h. And the precipitate was roasted at 850 ℃ for 3 h to obtain the oxide of Y in which the purity of Y_2 O_3/REO is 79.02% and the contents of major non-rare earth impurities are less than 0.21%.Over the whole process included acid leaching, solvent extraction, and oxalic acid precipitation-roasting,the yttrium yield is 82.04%.     

ASSESSMENT OF REDUCTION BEHAVIOR OF HEMATITE IRON ORE PELLETS IN COAL FINES FOR APPLICATION IN SPONGE IRONMAKING

Studies on isothermal reduction kinetics (with F grade coal) in fired pellets of hematite iron ores, procured from four different mines of Orissa, were carried out in the temperature range of 850–1000°C to provide information for the Indian sponge iron plants. The rate of reduction in all the fired iron ore pellets increased markedly with a rise of temperature up to 950°C, and thereafter it decreased at 1000°C. The rate was more intense in the first 30 minutes. All iron ores exhibited almost complete reduction in their pellets at temperatures of 900 and 950°C in < 2 hours' heating time duration, and the final product morphologies consisted of prominent cracks. The kinetic model equation 1 ? (1 ? α)^1/3 = kt was found to fit best to the experimental data, and the values of apparent activation energy were evaluated. Reductions of D. R. Pattnaik and M. G. Mohanty iron ore pellets were characterized by higher activation energies (183 and 150 kJ mol^?1), indicating carbon gasification reaction to be the rate-controlling step. The results established lower values of activation energy (83 and 84 kJ mol^?1) for the reduction of G. M. OMC Ltd. and Sakaruddin iron ore pellets, proposing their overall rates to be controlled by indirect reduction reactions.     

Removal of iron from quartz sands: A study by column leaching using a complete factorial design

The aim of the work was to decrease the iron content of ferrous quartz sands by fixed-bed column leaching with recycling of the leaching solutions in order to attain a product suitable for industrial use. Dissolution of iron was achieved by treating the sands in an acid medium with a reducing agent (oxalic acid) to convert Fe^III into Fe^II.The factors assumed to affect dissolution of iron, such as temperature, oxalic acid concentration, pH and flow-rate, were studied with a 2⁴ full factorial design in order to assess the main effects and the interactions among the factors.Removal of 46.1% iron gives a product containing 0.0163% Fe₂O₃ which is fit for industrial applications.     

Beneficiation of High-Aluminium-Content Hematite Ore by Soda Ash Roasting

High-aluminium-content iron ore is one of typical intractable iron ores, and magnetic separation and floatation processes are found impracticable to remove alumina from the ore effectively. In this article, a new process, roasting with addition of soda followed by leaching, is developed to remove aluminium from the ore. Results show that Al₂O₃ content decreases from 8.16% in raw ore to 2.13% in iron concentrate, and total iron grade increases from 48.92 to 63.21% when the ore is roasted at 1000°C for 15 min with the addition of 14.0% (wt.) sodium carbonate. Mechanisms of aluminium–iron separation were studied by using XRD, SEM, and thermodynamic methods, and it is shown that aluminium is transformed into sodium aluminosilicate, sodium aluminate, and corundum during roasting; sodium aluminate is able to be leached by water, so is sodium aluminosilicate by dilute acid solution, while corundum remains in the iron concentrate.