Release of Uranium by an Ore Treatment Unit at Caldas,MG,Brazil

This study aims to assess the behavior of the release of Unat (i.e. uranium isotopes in natural relative abundance) from the release of water treated by the Ore Treatment Unit (UTM) in Caldas, state of Minas Gerais, Brazil, during the years 1999 to 2009. During this period, the unit showed no industrial activity, except between 2004 and 2005, when UTM operated 400 tons of monazite in a process to obtain rare earths. Unat was analyzed by spectrophotometry once per week in the effluent waters at sampling point 014. Two fractions were considered for analysis: the soluble one (that passes through a 0.45 μm filter) and the particulate one (retained on a 0.45 μm filter). Statistical analyses were performed: the “Z” test and Pearson’s r2 correlation index. The values for the soluble fraction were consistently lower than those of the particulate fraction and no statistically significant correlation was observed between the soluble and particulate fractions. However, the particulate and the total fractions (the sum of soluble and particulate) strongly correlated.

Recovery of valuable metals from a low-grade nickel ore using an ammonium sulfate roasting-leaching process

Metal leaching from a low-grade nickel ore was investigated using an ammonium sulfate roasting-water leaching process. The nickel ore was mixed with ammonium sulfate, followed by roasting and finally leaching with water. During the process the effects of the amotmt of ammonium sulfate, roasting temperature, and roasting time on the leaching recovery of metal elements were analyzed. The optimum technological parameters were determined as follows： ammonium sulfate/ore ratio, 0.8 g/g; roasting temperature, 400℃; and roasting time, 2 h. Under the optimum condition the leaching recoveries ofNi, Cu, Fe, and Mg were 83.48%, 76.24%, 56.43%, and 62.15%, respectively. Furthermore, the dissolution kinetics of Ni and Mg from the nickel ore was studied. The apparent activation energies for the leaching reaction of Ni and Mg were 18.782 and 10.038 kJ.mo1-1, which were consistent with the values of diffusion control reactions. Therefore, the results demonstrated that the leaching recoveries of Ni and Mg were controlled by diffusion.

Beneficiation of the gold bearing ore by gravity and flotation

Gold concentration usually consists of gravity separation, flotation, cyanidation, or the combination of these processes. The choice among these processes depends on the mineralogical characterization and gold content of the ore. Recently, the recovery of gold using gravity methods has gained attention because of low cost and environmentally friendly operations. In this study, gold pre-concentrates were produced by the stepwise gravity separation and flotation techniques. The Knelson concentrator and conventional flotation were employed for the recovery of gold. Gold bearing ore samples were taken from Gümüshane Region, northern east part of Turkey. As a result of stepwise Knelson concentration experiments, a gold concentrate assaying around 620 g/t is produced with 41.4wt% recovery. On the other hand, a gold concentrate about 82 g/t is obtained with 89.9wt% recovery from a gold ore assaying 6 g/t Au by direct flotation.

Optimization of flotation variables for the recovery of hematite particles from BHQ ore

The technology for beneficiation of banded iron ores containing low iron value is a challenging task due to increasing demand of quality iron ore in India. A flotation process has been developed to treat one such ore, namely banded hematite quartzite （BHQ） containing 41.8wt% Fe and 41.5wt% SiO2,by using oleic acid, methyl isobutyl carbinol （MIBC）, and sodium silicate as the collector, frother, and dispersant, respectively. The relative effects of these variables have been evaluated in half-normal plots and Pareto charts using central composite rotatable design. A quadratic response model has been developed for both Fe grade and recovery and optimized within the experimental range. The optimum reagent dosages are found to be as follows： collector concentration of 243.58 g/t, dispersant concentration of 195.67 g/t, pH 8.69, and conditioning time of 4.8 min to achieve the maximum Fe grade of 64.25% with 67.33% recovery. The predictions of the model with regard to iron grade and recovery are in good agreement with the experimental results.

Effect of desliming on the magnetic separation of low-grade ferruginous manganese ore

In the present investigation, magnetic separation studies using an induced roll magnetic separator were conducted to beneficiate low-grade ferruginous manganese ore. The feed ore was assayed to contain 22.4% Mn and 35.9% SiO2, with a manganese-to-iron mass ratio （Mn：Fe ratio） of 1.6. This ore was characterized in detail using different techniques, including quantitative evaluation of minerals by scan- ning electron microscopy, which revealed that the ore is extremely siliceous in nature and that the associated gangue minerals are more or less evenly distributed in almost all of the size fractions in major proportion. Magnetic separation studies were conducted on both the as-received ore fines and the classified fines to enrich their manganese content and Mn：Fe ratio. The results indicated that the efficiency of separation for deslimed fines was better than that for the treated unclassified bulk sample. On the basis of these results, we proposed a proc- ess flow sheet for the beneficiation of low-grade manganese ore fines using a Floatex density separator as a pre-concentrator followed by two-stage magnetic separation. The overall recovery of manganese in the final product from the proposed flow sheet is 44.7% with an assay value of 45.8% and the Mn：Fe ratio of 3.1.

Aerosol flotation of low-grade refractory molybdenum ores

The characteristics of aerosol flotation, which include the effect of the concentration and particle size of kerosene aerosol on the molybdenum （Mo） flotation index and the effect of kerosene aerosol dosing method on the kerosene dosage and flotation time, were studied in the flotation of low-grade refractory molybdenum ores using kerosene aerosol. The results revealed that the particle size and concentration of kerosene aerosol had little effect on the Mo grade but had significant effect on the Mo recovery. A smaller particle size and a lower con-centration of kerosene aerosol were beneficial to the Mo aerosol flotation. For the received Mo ore samples, the optimized particle size of kerosene aerosol was 0.3-2 μm and the optimized aerosol concentration was 14 mg/L. The compressed air atomizer had a more uniform dis-tribution of aerosol particles than the ultrasonic atomizer, and the aerosol concentration was controlled easily, so the compressed air atomizer was more suitable for the research of aerosol flotation. Compared with conventional flotation in which kerosene was directly added into the ore pulp, the flotation time was reduced by ~30%, and the dosage was decreased by ~20% in aerosol flotation, while the Mo flotation index was similar.

Kinetics of leaching refractory gold ores by ultrasonic-assisted electro-chlorination

The resources of refractory gold ores are abundant, and their effective treatment can bring good economic benefits. This paper investigated the kinetics of leaching gold from refractory gold ores by ultrasonic-assisted electro-chlorination. The effects of ultrasound time ratio, initial hydrochloric acid concentration and leaching temperature on the kinetic parameters were discussed. It is found that the leaching ratio goes up with all the factors increasing. The reaction kinetics is controlled by diffusion. When ultrasound improves the diffusion by reducing the diffusion resistance, the activation energy increases to 37.1 kJ/mol.

Reductive leaching of manganese from low-grade pyrolusite ore in sulfuric acid using pyrolysis-pretreated sawdust as a reductant

Manganese （Mn） leaching and recovery from low-grade pyrolusite ore were studied using sulfiaric acid （H2SO4） as a leachant and pyrolysis-pretreated sawdust as a reductant. The effects of the dosage of pyrolysis-pretreated sawdust to pyrolusite ore, the concentration of sulfuric acid, the liquid/solid ratio, the leaching temperature, and the leaching time on manganese and iron leaching efficiencies were inves- tigated. Analysis of manganese and iron leaching efficiencies revealed that a high manganese leaching efficiency was achieved with low iron extraction. The optimal leaching efficiency was determined to be 20wt% pyrolysis-pretreated sawdust and 3.0 mol/L H2SO4 using a liq- uid/solid ratio of 6.0 mL/g for 90min at 90℃. Other low-grade pyrolusite ores were tested, and the results showed that they responded well with manganese leaching efficiencies greater than 98%.

Kinetics of chemical leaching of chalcopyrite from low grade copper ore: behavior of different size fractions

The kinetics of the chemical leaching of copper from low grade ore in ferric sulfate media was investigated using the constrained least square optimization technique. The experiments were carried out for different particle sizes in both the reactor and column at constant oxidation-reduction potential （Eh）, pH values, and temperature. The main copper mineral was chalcopyrite. About 40% of Cu recovery is obtained after 7 d of reactor leaching at 85℃ using -0.5 mm size fraction, while the same recovery is obtained at 75℃ after 24 d. Also, about 23% of Cu recovery is obtained after 60 d of column leaching for ＋4--8 mm size fraction whereas the Cu recovery is as low as about 15% for ＋8--12.7 and ＋12.7--25 mm size fractions. A 4-stage model for chalcopyrite dissolution was used to explain the observed dissolution behaviors. The results show that thick over-layers of sulphur components cause the parabolic behavior of chalcopyrite dissolution and the precipitation of Fe3＋ plays the main role in chalcopyrite passivation. In the case of coarse particles, transformation from one stage to another takes a longer time, thus only two stages including the initial reaction on fresh surfaces and S0 deposition are observed.

Reaction condition optimization and kinetic investigation of roasting zinc oxide ore using (NH_4)_2SO_4

An orthogonal test was used to optimize the reaction conditions of roasting zinc oxide ore using（NH_4）_2SO_4. The optimized reaction conditions are defined as an（NH_4）_2SO_4/zinc molar ratio of 1.4：1, a roasting temperature of 440°C, and a thermostatic time of 60 min. The molar ratio of（NH_4）_2SO_4/zinc is the most predominant factor and the roasting temperature is the second significant factor that governs the zinc extraction. Thermogravimetric-differential thermal analysis was used for（NH_4）_2SO_4 and zinc mixed in a molar ratio of 1.4：1 at the heating rates of 5, 10, 15, and 20 K·min-1. Two strong endothermic peaks indicate that the complex chemical reactions occur at approximately 290°C and 400°C. XRD analysis was employed to examine the transformations of mineral phases during roasting process. Kinetic parameters, including reaction apparent activation energy, reaction order, and frequency factor, were calculated by the Doyle-Ozawa and Kissinger methods. Corresponding to the two endothermic peaks, the kinetic equations were obtained.

Deselenization and detellurization of precious-metal ore concentrates by swelling oxidizing roasting and successive alkaline leaching

A new technique of swelling oxidizing roasting and alkaline leaching was proposed for deselenization and detellurization of pre- cious-metal ore concentrates. Alkali-metal and alkaline-earth-metal chlorides and carbonates were preliminarily selected as swelling agents. The roasting removal rate and alkaline leaching rate of selenium and tellurium were investigated, and NaCl was selected as an appropriate swelling agent. Furthermore, the effects of various factors on the selenium gasification rate and leaching rate of selenium and tellurium were investigated. The results show that the selenium gasification rate reaches 88.41% after swelling oxidizing roasting for 2 h at 510~C using an NaC1 dosage coefficient of 100 and a sulfuric acid dosage coefficient of 1.3; the amorphous elemental tellurium is completely transformed into TeO2. The roasted product is subjected to alkaline leaching using a 100 g/L NaOH solution, which results in a selenium leaching rate of 10.51%, a total selenium removal rate of 98.92%, and a tellurium leaching rate of 97.64%. In the alkaline leaching residue, the contents of selenium, tellurium, gold, platinum, and palladium are 0.7825%, 5.492%, 8.333%, 0.2587%, and 1.113%, respectively; the precious metals are enriched approximately sixfold.

Comprehensive recovery of gold and base-metal sulfide minerals from a low-grade refractory ore

The comprehensive recovery of small amounts of valuable minerals such as gold and base-metal sulfide minerals from a low-grade refractory ore was investigated. The following treatment strategy was applied to a sample of this ore： gold flotation-gold concen- trate leaching-lead and zinc flotation from the gold concentrate leaching residue. Closed-circuit trials of gold flotation yielded a gold concen- trate that assayed at 40.23 g·t-1 Au with a recovery of 86.25%. The gold concentrate leaching rate was 98.76%. Two variants of lead-zinc flotation from the residue--preferential flotation of lead and zinc and bulk flotation of lead and zinc--were tested using the middling processing method. Foam from the reflotation was returned to the lead rougher flotation or lead-zinc bulk flotation, whereas middlings from reflotation were discarded. Sulfur concentrate was a byproduct. The combined strategy of flotation, leaching, and flotation is recommended for the treatment of this kind of ore.

Hydrometallurgical application for treating a Nigerian chalcopyrite ore in chloride medium:Part I.Dissolution kinetics assessment

The dissolution kinetics of a Nigerian chalcopyrite ore in hydrochloric acid was studied in this article. Acid concentration, reaction temperature, and ore particle size were chosen as experimental parameters. The chemical and morphological studies of the ore before and after leaching at optimal conditions were carried out by X-ray diffraction （XRD） and scanning electron microscopy （SEM）. It is revealed that increasing the acid concentration and system temperature and decreasing the ore particle size greatly enhances the dissolution rate. The dissolution kinetics was found to follow the shrinking core model for the diffusion control mechanism where the activation energy （Ea） of 32.92 kJ.mo1-1 was obtained for the process and supported by morphological changes at a higher dissolution of 91.33%.

Reductive leaching of low-grade manganese ore with pre-processed cornstalk

Cornstalk is usually directly used as a reductant in reduetive leaching manganese. However, low utilization of cornstalk makes low manganese dissolution ratio, In the research, pretreatment for cornstalk was proposed to improve manganese dissolution ratio. Cornstalk was preprocessed by a heated sulfuric acid solution （1.2 M of sulfuric acid concentration） for 10 min at 80℃. Thereafter, both the pretreated solution and the residue were used as a reductant for manganese leaching. This method not only exhibited superior activity for hydrolyzing cornstalk but also enhanced manganese dissolution. These effects were attributed to an increase in the amount of reductive sugars resulting from lignin hydrolysis. Through acid pretreatment for cornstalk, the manganese dissolution ratio was improved from 50.14% to 83.46%. The present work demonstrates for the first time the effective acid pretreatment of cornstalk to provide a cost-effective reductant for manganese leaching.

Kinetic approach to the study of froth flotation applied to a lepidolite ore

The number of published studies related to the optimization of lithium extraction from low-grade ores has increased as the demand for lithium has grown. However, no study related to the kinetics of the concentration stage of lithium-containing minerals by froth flotation has yet been reported. To establish a factorial design of batch flotation experiments, we conducted a set of kinetic tests to determine the most selective alternative collector, define a range of pulp p H values, and estimate a near-optimum flotation time. Both collectors（Aeromine 3000 C and Armeen 12D） provided the required flotation selectivity, although this selectivity was lost in the case of pulp p H values outside the range between 2 and 4. Cumulative mineral recovery curves were used to adjust a classical kinetic model that was modified with a non-negative parameter representing a delay time. The computation of the near-optimum flotation time as the maximizer of a separation efficiency（SE） function must be performed with caution. We instead propose to define the near-optimum flotation time as the time interval required to achieve 95%–99% of the maximum value of the SE function.

Infiltration behavior of sintering liquid on nuclei ores during low-titanium ore sintering process

Sinter strength is dependent not only on the self-intensity of the residual rude and bonding phase but also on the bonding degree between them. The infiltration behavior of sintering liquid on nuclei ores influences the bonding degree, which ultimately determines the sinter strength. Infiltration tests were conducted using micro-sinter equipment. The infiltration area index of original liquid（IAO）, infiltration volume index of secondary liquid（IVS）, and sinter body bonding strength（SBS） were proposed to study the melt infiltration behavior. The results show that the IVS first increases and then decreases with increasing TiO2 content in adhering fines, whereas the IAO exhibits the opposite behavior. Compared with the original liquid, the secondary liquid shows lower porosity, smaller pores, and more uniform distribution. The SBS increases first and then decreases with increasing IAO and TiO2 content, and reaches a maximum when the IAO and TiO2 contents are approximately 0.5 and 2.0wt%, respectively. The SBS first increases and then tends to be stable with increasing IVS. The TiO2 content is suggested to be controlled to approximately 2.0wt% in low-titanium ore sintering.

Alkaline pretreatment for chlorine removal from high-chlorine rhodochrosite

Chloride in manganese ore adversely affects mineral extraction. The mechanisms and the factors that influence an alkali pretreatment to removal chlorine from manganese ore were explored to eliminate hazards posed by chlorine during the electrolysis of manganese. The results showed that sodium carbonate, when used as an alkaline additive, promoted the dissolution of insoluble chloride, enhanced the migration of chloride ions, and effectively stabilized Mn^（2＋）. The optimal conditions were a sodium carbonate concentration of 0.45 mol·L^（-1）, a liquid-solid ratio of 5：1 mL ·g^（-1）, a reaction time of 1 h, and a temperature of 25°C. The chlorine removal efficiency was greater than 95%, and the ore grade（Mn） was increased by 2.7%.

Cost-effective integrated strategy for the fabrication of hard-magnet barium hexaferrite powders from low-grade barite ore

Ultrafine barium hexaferrite（BaFe12O19） powders were synthesized from the metallurgical extracts of low-grade Egyptian barite ore via a co-precipitation route. Hydrometallurgical treatment of barite ore was systematically studied to achieve the maximum dissolution efficiency of Fe（～99.7%） under the optimum conditions. The hexaferrite precursors were obtained by the co-precipitation of BaS produced by the reduction of barite ore with carbon at 1273 K and then dissolved in diluted HCl and FeCl3 solution at pH 10 using NaOH as a base; the product was then annealed at 1273 K in an open atmosphere. The effect of Fe^3＋/Ba^2＋ molar ratio and the addition of hydrogen peroxide（H2O2） on the phase structure, crystallite size, morphology, and magnetic properties were investigated by X-ray diffraction, scanning electron microscopy, and vibrating sample magnetometry. Single-phase BaFe（12）O（19） powder was obtained at an Fe^3＋/Ba^2＋ molar ratio of 8.00. The formed powders exhibited a hexagonal platelet-like structure. Good maximum magnetization（48.3 A×m^2×kg^–1） was achieved in the material prepared at an Fe^3＋/Ba^2＋ molar ratio of 8.0 in the presence of 5% H2O2 as an oxidizer and at 1273 K because of the formation of a uniform, hexagonal-shaped structure.

Bioleaching of copper oxide ore by P seudomonas aeruginosa

Bioleaching is an environmentally friendly method for extraction of metal from ores. In this study, bioleaching of copper oxide ore by Pseudomonas aeruginosa was investigated. Pseudomonas aeruginosa is a heterotrophic bactermm that can produce various organic acids in an appropriate culture medium, and these acids can operate as leaching agents. The parameters, such as particle size, glucose percentage in the culture medium, bioleaching time, and solid/liquid ratio were optimized. Optimum bioleaching conditions were found as follows： particle size of 150-177 μm, glucose percentage of 6%, bioleaching time of 8 d, and solid/liquid ratio of 1：80. Under these conditions, 53% of copper was extracted.

Pre-desilication and digestion of gibbsitic bauxite with lime in sodium aluminate liquor

The effect of lime on the pre-desilication and digestion of gibbsitic bauxite in synthetic sodium aluminate liquor at different tem- peratures was investigated. The bauxite is comprised of gibbsite, aluminogoethite, hematite, kaolin, quartz, and minor boehmite. Lime in- creases the desilication efficiency of the bauxite during the pre-desilication process by promoting the conversion of sodalite and cancrinite to hydrogamet. Desilication reactions during the digestion process promoted by lime result in the loss of A1203 entering the red mud, but the amount of aluminogoethite-to-hematite conversion promoted by lime leads to the increase of aluminogoethific A1203 entering the digested liquor. The alumina digestion rate at 245~C is higher than that at 145 C due to the more pronounced conversion of aluminogoethite to hema- tite. The soda consumption during the digestion process decreases due to lime addition, especially at higher temperatures.