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1

Effect of oxidation roasting on ilmenite flotation

Mehdilo A., Irannajad M., Rezai B.

Physicochemical Problems of Mineral Processing

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2014

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Vol. 50, iss. 2

493--505

EN

In this study, the oxidation roasting was used to modify the ilmenite chemistry to improve the collector adsorption and ilmenite flotation behavior. The results indicated that the increase of the roasting temperature up to 600 ºC increased the ilmenite flotation recovery. The maximum collector adsorption density occurred on ilmenite roasted at 600 °C which resulted in the increase of ilmenite flotation recovery from 73.5% to 91% at pH 6.3. This improvement was attributed to a greater conversion of Fe2+ to Fe3+ ions and significant decrease in the zeta potential of ilmenite. At the roasting temperatures of 750 °C and 950 °C, ilmenite was converted to rutile and some trivalent iron containing phases such as hematite and pseudobrookite. Under these conditions, the collector adsorption and hence ilmenite flotation improved, however the flotation recovery decreased after the roasting at 950 °C.

2

Flotation of zinc oxide ore using cationic and cationic-anionic mixed collectors

Mehdilo A., Irannajad M., Zarei H.

Physicochemical Problems of Mineral Processing

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2013

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Vol. 49, iss. 1

145--156

EN

Flotation of smithsonite coming from a cerussite flotation circuit was investigated using Armac C and Armac T as cationic collectors and their mixtures with potassium amyl xanthate (KAX) as an anionic collector. Under optimum condition using Armac T with sodium hexamethaphosphate as a de-pressant, a smithsonite concentrate with 42% Zn and 89.6% recovery is obtained while in the presence of sodium silicate depressant Armac C produces a concentrate containing 40.8% Zn with 92.5% recovery. However, Armac T acts a little more selectively than Armac C but their mixture with KAX act inversely. The KAX-Armac C mixed collector improves the Zn grade and recovery of the flotation concentrate while KAX-Armac T does not. At the mixture ratio of KAX - Armac C 2:1, the amine collector consump-tion is decreased and an optimal concentrate with 94% recovery and 43% Zn grade is obtained without any gangue depressant reagents.

3

Characterization and beneficiation of iranian low-grade manganese ore

Mehdilo A., Irannajad M., Hojjati-Rad M. R.

Physicochemical Problems of Mineral Processing

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2013

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Vol. 49, iss. 2

725--741

EN

The mineralogical studies indicated that the Charagah ore deposit contains approximately 17% pyrolusite, 78% calcite and 3–4% quartz. Pyrolusite as a main valuable mineral is found in the forms of coarse and fine pyrolusites. The coarse grains pyrolusite with simple texture is liberated at 180 micrometers. Another kind of pyrolusite with particle size finer than 10 m is disseminated inside gangue phases. This kind of pyrolusite has important effect in beneficiation processes and can affect the manganese grade of the concentrate and its recovery negatively. By jigging machine a pre-concentrate with 20% MnO and a final tailing with about 13% manganese loss are obtained. Using tabling technique or wet high intensity magnetic separation (WHIMS) and also their combination with jigging machine, production of a final pyrolusite concentrate with suitable grade but average recovery is possible. By jigging-tabling a concentrate with – 500+45 m size fraction, 44.3% MnO and 61.3% recovery is obtained while jigging-WHIMS produces a concentrate containing 52.6% MnO with a recovery up to 56.6% and d80 = 180 m.

4

Iron removing from titanium slag for synthetic rutile production

Mehdilo A., Irannajad M.

Physicochemical Problems of Mineral Processing

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2012

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Vol. 48, iss. 2

425-439

EN

Leaching with sulfuric and hydrochloric acids for upgrading titanium slag obtained from the Qara-aghaj ilmenite concentrate to a high grade titanium dioxide (TiO2) was studied. The titanium slag containing 72.7% TiO2 and 7.8% Fe2O3 was leached at varying solid/liquid (S/L) ratio, particle size, leaching time and acid concentration. The optimum amount of S/L ratio, particle size, acid concentration and leaching time using both acids were determined as 1:4, -100 µm, 8% and 2 h. Under optimum conditions, by using sulfuric acid, a titanium dioxide concentrate with 86.8% TiO2 and 1.87% Fe2O3 was produced, while employing hydrochloric acid resulted in a concentrate containing 91% TiO2 and 0.61% Fe2O3. The titanium dioxide concentrate prepared with hydrochloric acid, having acicular texture and by means of elongated separate particles; has less impurities than the one produced by sulfuric acid. The results demonstrate that hydrochloric acid dissolution of titanium slag and removing some impurities such as SiO2 and MgO provide a product with good quality is prepared which is suitable as raw material for TiO2 pigment production through the chloride process.