m. abdollahy

This paper studies the extraction of alumina from kaolin through an acidic leaching process. The kaolin sample used in this study was taken from the Zonuz kaolin mine (Iran Porcelain Industries Company) at particle sizes smaller than 150 microns. Calcination of the kaolin sample was conducted at 700 ºC for 2 hours. Aluminum was leached from calcined kaolin with an H2SO4 acid solution. Temperature and time of leaching, acid concentration, and liquid-to-solid ratio were investigated as leaching parameters. The optimal leaching conditions were considered at 90 °C, 3 hours duration, 2.5 M acid concentration, 7 ml/g liquid-to-solid ratio, and 600 rpm stirring speed, and under these conditions, aluminum leaching recovery was obtained at 94.87%. To remove the iron impurity that is transferred to the solution during leaching, the pregnant leaching solution was purified. Purification was done by solvent extraction with an organic phase containing D2EHPA as an extractant and oil as a diluent, iron removal was achieved at 88.71%. By precipitating the aluminum hydroxide from a solution of aluminum sulfate with sodium hydroxide followed by calcination of the produced aluminum hydroxide at a temperature of 1200 °C for 2 hours, alumina with a purity of 97% was produced. The total recovery of alumina obtained from the studied kaolin under optimal conditions was 89.46%, which is desirable.

In this study, the feasibility of improving the flotation circuit performance of a phase 1 of Sungun copper plant was investigated. For this purpose, circuit performance was first monitored ., It was found that in the rougher stage the some copper exists in fractions greater than 106 and smaller than 38 microns was lost. Final concentrate analysis taken on different days showed that the copper grade decreased due to the iron content increasing. Mineralogical studies also confirmed the maximum presence of pyrite in the final concentrate. Field surveys of the flotation circuit showed that there are many technical problem in the circuit that affect on the flotation process. These drawbacks including the presence of thick tubes at the surface of the re-cleaner cell, the lack of a standard pattern of wash water distribution and jet spraying. Measurement of pH in different points of the circuit showed a decrease of pH below 11 and so hydrophobicity of pyrite caused. Measurement of operating variables in the circuit showed that the wash water rate and gas rate used in the circuit were higher and lower than the values suggested by the researchers and the designer, respectively. Due to the deficiencies identified in the performance evaluation phase, laboratory tests were designed and performed in two stages using mechanical flotation cell and column flotation cell. Adjusting the gas rate and the wash water rate at the laboratory scale at the intervals suggested by the researchers resulted in the increase in 3% grade and 10% recovery.

One of the challenges in the rare earth elements (REE) production industry is the difficulty of separating them from each other. Therefore, in the present study, the possibility of cerium separation from carbonate concentrate of REEs containing 22.75% Ce, 9.4% La and 8.6% Nd (total REEs of 41.51%) was investigated. Two methods were used to separate Ce from other REEs. In the first method, due to the very low solubility of Ce4+, the calcination of carbonate concentrate and conversion of cerium carbonate to cerium oxide and selective dissolution of other REEs from the calcine were investigated. In the second method, prior to the carbonate precipitation of REEs, the possibility of cerium oxidation in the sulfate solution (by using potassium permanganate) and its subsequent selective separation from other REEs was investigated. Results of the calcination (first method) showed that only 35 and 45% of La and Nd were dissolved at 80⁰C after 2 hours leaching with 1 molar nitric acid, respectively. Results of the calcine leaching with 4 molar nitric acid at 80⁰C for 2 hours showed that more than 90% of the La and Nd were dissolved. In this conditions, 35% of the Ce was also dissolved and the leaching process was not selective. In the second method, using potassium permanganate with a stoichiometric ratio of 1.5 to 1 for oxidant to Ce, pH 3.5 for 1 hour, 98% of the Ce and less than 1% of La and Nd were precipitated. The resulting hydroxide was calcined at 850⁰C for 1 h to produce CeO2 and finally a Ce oxide concentrate was obtained with 99% purity.

Nugget effect is a common feature of many vein-style gold deposits. This style of mineralization results in complication in many exploitation stages including resource and average grade estimation, representative sample selection, ore characterization as well as processing. Accurate and reliable ore characterization is a vital stage in any mineral processing project. Ore characterization in high nugget effect gold ores is so complex due to uncertainties associated with representativity of the selected sample. The main challenge is to provide a representative sample for identification of each process mineralogy parameter such as average grade, liberation, grinding size, mineral association, floatability, etc. Here, different protocols are collected and compared to draw a comprehensive picture of the available methods. The results are examined on a high nugget effect gold ore from Northeast of Iran, Azghad mine. Processing experiments were conducted to develop a process suitable for Azghad gold mine. To investigate the feasibility of pre-concentration, gravity methods using a Nelson separator was applied on the oxide sample as the main gold vein. 43.3% of total gold was concentrated in the gravity concentrate with only 2% weight recovery, suggesting gravity concentration can be considered as an efficient method in this ore. Gold preg-robbing experiments were performed for Shale sample due to the organic matter, clay minerals and mica. Preg-robbing potential of the shale ore was determined to be 1.1 ppm. Optimization of operating parameters was performed by central composite design method for the combined oxide and shale samples.

Copper oxide minerals such as malachite do not respond well to the traditional copper sulfide collectors, and require alternative flotation schemes. In many copper ore mines, significant copper oxide minerals, especially malachite, are associated with sulfide minerals. Considering that xanthates are most widely used in the flotation of sulfide minerals as well as copper sulfide minerals and, hydroxamate has shown a good selectivity for copper oxide minerals. Use of the synergistic effect of xanthate and hydroxamate can be an effective way to increase the flotation efficiency of copper oxide minerals along with sulfide minerals. In this work, we investigate the individual interactions of potassium amyl xanthate (PAX) and potassium alkyl hydroxamate (HXM) with the natural malachite and explore their synergistic effects on the malachite flotation. The results of solubility of malachite in collector solutions, changes in the malachite surface potential, adsorption kinetics, adsorption densities, dynamic contact angles, FT-IR analyses, and small-scale flotations, are discussed. The results obtained demonstrate that PAX and HXM are chemically co-adsorbed on the malachite surface, and the amount of PAX adsorbed on the malachite surface is considerably increased in the mixed PAX/HXM systems because of the co-adsorption mechanism. The flotation results confirm that the mixed PAX/HXM exhibit a superior flotation performance of malachite compared to the individual system of PAX or HXM. Based on these results, the mixed PAX/HXM exhibit a remarkable synergism effect on malachite surface hydrophobicity

The Esfordi phosphate concentrate with total assay of REE equal to 1.2 percent is one of the richest source of rare earth elements in Iran. It contains Ce(5608 ppm), La(1959ppm) and Nd(2227ppm). In this research, the effect of mechanical activation on the leaching of rare earth elements by nitric acid from Esfordi phosphate concentrate was investigated. For the investigation of mechanical activation effect on leaching of REE, a planetary ball mill was used with steel balls of diameter about 20 mm, ball to powder ratio of 15 and 2 and with a retention time of 90 minuts in dry situation and air atmosphere. Because of high and low leachability of fluorapatite and REE minerals by nitric acid, respectively, tow stage leaching methode was used. The first leaching stage residue, that contained about 99 % of REEs in initial phosphate concentrate, was activated mechanically at ball to powder ratio of 12 and then leached. Results showed that the mechanical activation had important influence on leaching of REE by nitric acid from phosphate concentrate and improved the leaching of Ce, La and Nd from about 1% (without mechanical activation) to 35.79, 42.61 and 30.34% , respectively. The amorphization degree increased to 1.2 and 57% after 90 min intensive milling with ball to powder ratio of 2 and 15 to 1, respectively. Also, the crystallite size and microstrain of Fluorapatite differed from 225 nm and 0.08 % for initial sample to 225 nm and 0.09 % and 81.8 nm and 0.63 % for mechanically activated samples with ball to powder ratio of 2 and 15 to 1 respectively.