Thesis (M.S.)--University of Wisconsin--Madison. / Typescript. eContent provider-neutral record in process. Description based on print version record. Includes bibliographical references.

This study was aimed at recovering lithium from china clay waste using a combination of froth flotation, magnetic separation, roasting and leaching. The china clay waste produced by Goonvean Ltd contains about 0.84% Li2O and 0.36% Rb2O, present in some of the mica minerals. Among the mica minerals, zinnwaldite is the major source of lithium with smaller amounts being contributed by muscovite. The results of the flotation tests showed that the dodecylamine collector dosage had a greater effect on the recovery and grade of mica minerals to concentrate than pH over the range tested. It was found that a mica concentrate containing 1.45% Li2O, 0.55% Rb2O and 4.47% Fe2O3 could be produced at a recovery of 98.6%, 85.2% and 92.8% respectively. Mineralogical analysis of the flotation products showed that the concentrate consisted mainly of muscovite, zinnwaldite and kaolinite with minor amounts of K-feldspar and quartz. The tailing consisted of mainly quartz, K-feldspar and kaolinite with minor amounts of apatite, topaz, zinnwaldite and muscovite. Further upgrading of the concentrate was found to be possible using a wet high intensity magnetic separator producing a magnetic fraction containing 2.07% Li2O, 0.74% Rb2O and 7.42% Fe2O3 with a recovery of 73%, 67% and 77% respectively. A mineralogical analysis of the separation products showed that the magnetic fraction consisted of predominantly zinnwaldite with muscovite as the main contaminant. The non-magnetic fraction consisted of muscovite and kaolinite as the main minerals while zinnwaldite, K-feldspar and quartz were subordinate. Electron-microprobe analysis on individual mica grains have shown that zinnwaldite and muscovite contain on average a calculated Li2O content of 3.88% and 0.13% respectively. Lithium extraction from the concentrate is only possible after the lithium has been converted into a water-soluble compound. Thus, in order to convert the lithium in concentrate into a water-soluble compound, the gypsum and limestone lithium extraction methods together with the new method of using sodium sulphate were investigated. The process involved roasting a predetermined amount of lithium-mica concentrate with either gypsum, limestone or sodium sulphate at various temperatures and subsequently leaching the pulverised materials in water at 85oC. A lithium extraction efficiency of about 84% was obtained using gypsum at 1050oC while rubidium extraction was very low at 14%. It was found possible to extract about 97% Li and 16% Rb if the concentrate was roasted with sodium sulphate at 850oC. Processing the concentrate with limestone resulted in very low lithium extraction. Iron co-extraction was low in all cases. The XRD analysis of the gypsum and sodium sulphate roast-products showed that the water soluble lithium species were KLiSO4 and Li2KNa(SO4)2 respectively. Preliminary tests on the leach solution obtained by using sodium sulphate as an additive have shown that a Li2O3 product with a purity of > 90% could be produced by precipitation with sodium carbonate although more work is required to reach the industrial target of > 99%. The lithium carbonate obtained with Li2CO3 content of about 90% is still suitable for use in the glass and ceramic industries, and as feedstock for the production of high-purity lithium compounds. An economic evaluation of the proposed lithium carbonate production plant has indicated an annual rate of return on the investment before tax of 7.2%.

Thesis (B.S.)--University of Missouri, School of Mines and Metallurgy, 1899. / Year of thesis publication determined from "Forty-First Annual Catalogue, School of Mines and Metallurgy, University of Missouri". The entire thesis text is included in file. Typescript. Illustrated by author. Title from title screen of thesis/dissertation PDF file (viewed March 3, 2010)

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Thesis (B.S.)--University of Missouri, School of Mines and Metallurgy, 1912. / The entire thesis text is included in file. Typescript. Illustrated by authors. Title from title screen of thesis/dissertation PDF file (viewed March 26, 2009) Includes bibliographical references.

Thesis (B.S.)--University of Missouri, School of Mines and Metallurgy, 1907. / The entire thesis text is included in file. Typescript. Illustrated by authors. T. C. Tseung determined to be Tsik Chan Tseung, J. C. Long determined to be James Carter Long, A. P. Fellows determined to be Aubrey P. Fellows, and E. R. Wash determined to be Edwin Richard Wash from "Forty-First Annual Catalogue. School of Mines and Metallurgy, University of Missouri". Title from title screen of thesis/dissertation PDF file (viewed January 26, 2009)

The kinetics of direct reduction of a commercial unagglomerated iron ore, with low-rank coal chars, have been investigated in the temperature range of 800-950°C (1073-1223 K) using a laboratory-size rotary reactor. The variables studied were temperature, coal type, particle size of coal and ore, fixed carbon-to-iron ratio, rotational speed-and percent filling. In addition the effects of a catalyst on the Boudouard reaction and of inert gas flushing on the reduction rate were determined. Mixing studies at room temperature and at reduction temperature yielded the best mixing conditions prior to the kinetics determinations. Agglomeration between the reduced particles was also studied. The mixing experiments at room temperature yielded the following. The degree of mixing depends almost entirely on the coal-to-ore size ratio. Best mixing is achieved with values of this ratio of 1 and smaller, for ore particles larger than 254 µm; for smaller sizes than this, good mixing can be obtained at higher coal-to-ore size ratios. At reduction temperature, improvement in the reduction rate was not obtained either by further increasing the fixed carbon-to-iron ratio from 0.32 to 0.64 or by varying the rotational speed from 7 to 20 r.p.m.. In the kinetics experiments, the overall reduction process was found to be controlled up to 0.5 to 0.8 fractional reduction by the Boudouard reaction, depending on the particle size and temperature; from then on, the kinetics were controlled essentially by the reduction reaction. The activation energies obtained were 224 kJ/mole for the Boudouard reaction, using sub-bituminous coal char, and 264 kJ/mole using lignite coal char; these values correspond to that of a catalyzed reaction. The catalytic effect of the coal ash on the Boudouard reaction was found to be much larger than the respective effect of metallic iron. The presence of a diluent gas extended the fractional reduction over which Boudouard reaction control is exerted. The activation energy obtained for the reduction of wustite by CO is 116.4 kJ/mole. The analysis of the Pco/Pco₂ ratio produced by the reaction proved to be a powerful tool in elucidating the rate controlling step. Smaller ore particles were found to agglomerate considerably more, in the non-catalyzed experiments; the addition of a catalyst for the Boudouard reaction also produced larger agglomerates. In neither case did agglomeration retard the reduction rate to a considerable extent. No accretion growth was observed on the reactor wall. Estimative calculations showed that similar throughputs can be obtained by processing the unagglomerated concentrate, as compared to operations which utilize indurated pellets under the same conditions. An advantage of a process using concentrates is the lower temperature at which it can be operated. / Applied Science, Faculty of / Mining Engineering, Keevil Institute of / Graduate

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Typescript (photocopy). / Digitized by Kansas Correctional Industries

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