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Title: Recovery of lithium from china clay waste using a combination of froth flotation, magnetic separation, roasting and leaching
Author: Siame, Edward
Awarding Body: University of Exeter
Current Institution: University of Exeter
Date of Award: 2011
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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%.
Supervisor: Pascoe, Richard Sponsor: Commonwealth Scholarship Commission
Qualification Name: Thesis (Ph.D.) Qualification Level: Doctoral
EThOS ID:  DOI: Not available
Keywords: Flotation; Roasting; Leaching