In this thesis work, a high grade chromite concentrate obtained from PinarbaSi, Kayseri region of Turkey was reacted with sulfuric acid solution to determine the optimum conditions of leaching of chromite ores. Conventional methods for producing chromium are still valid in industry. The main process in production includes soda melting and taking it into solution in the form of chemical compounds such as sodium dichromate. Three and six valence states of chromium have importance in production. Trivalent chromium, chrome ore or chromite have no harmful effects. However, compounds of hexavalent chromium show toxic, irritating and corrosive action to people and environment. In the mentioned conventional method, main products consist of hexavalent chromium compounds. In this study, only trivalent chromium remained in the leach solution and did not change to hexavalent state. Obtained product after leaching was chromium(III) sulfate. The maximum extraction of chromium in the absence of perchloric acid was 94.1 % under the conditions of 175 &deg / C, 6 hours and 84.6 wt % sulfuric acid. The maximum extraction of chromium in the presence of perchloric acid was 98.7 % under the conditions of 175 &deg / C, 2 hours, 84.6 wt % sulfuric acid and &frac12 / perchloric acid / chromite ratio. The latter one was also the highest recovery value obtained during the experiments. Moreover, in none of the analyzed samples, appreciable amounts of hexavalent chromium was found during analyses.

In this study, an attempt has been made to hydrometallurgical extraction of cobalt and nickel by atmospheric pressure sulphuric acid leaching and a pug-roast-leach process using two stage roasting for lateritic ore. The ore used in the study was obtained from &Ccedil / aldag Lateritic Ore, Manisa, Turkey. The metal contents of the ore are 2.1 % Ni, 0.12 % Co, 32.45 % Fe, 1.01 % Mn, 2.58 % Cr, 0.78 % Mg and 1.01 % Al. The reserve of lateritic ore deposit is approximately 40 million tonnes. In the study, first sulphuric acid leaching was applied at atmospheric pressure for leaching the &Ccedil / aldag lateritic ore. The effect of various parameters, such as leaching time, leaching temperature, particle size, pulp density and acid strength on Ni and Co extractions were determined. By leaching at 80oC for 40 wt % H2SO4 addition of ore, 1/3 pulp density, Ni and Co extractions were found 44.49 % and 53.03 % respectively, yielded a pregnant solution containing 3.11 g/L Ni and 0.12 g/L Co. But the result of atmospheric pressure sulphuric acid leaching was considered insufficient from the recovery point of view. In the pug-roast-leach process, which is consisted of a two stage roasting followed by water leaching, decomposition temperature differences of sulphates of cobalt, nickel and iron are exploited. In this process, amount of acid, sulphatization and decomposition temperature, sulphatization and decomposition time, leaching temperature and time, solid/liquid ratio, and the effect of water addition during pugging were optimized. Under the optimized conditions (sulphuric acid: 25 wt % of ore / moisture: 20 wt % of ore / sulphatization temperature: 450oC / sulphatization time: 30 minutes / decomposition temperature: 700oC / decomposition time: 60 minutes / leaching temperature: 70oC / leaching time: 30 minutes and solid-liquid ratio: 1/4 by weight), Co and Ni extractions were found 91.4 and 84.4 percent, respectively. A pregnant solution containing 3.084 g/L Ni and 0.185 g/L Co was obtained. These results were considered sufficient for the leaching of lateritic nickel ores.

Titanium diboride was produced both by volume combustion synthesis (VCS) and by mechanochemical synthesis through the reaction of TiO2, B2O3 and metallic Mg. Reaction products were expected to be composed of TiB2 and MgO. However, side products such as Mg2TiO4, Mg3B2O6, MgB2 and TiN were also present in the products obtained by volume combustion synthesis. Formation of TiN could be prevented by conducting the volume combustion synthesis under argon atmosphere. Mg2TiO4 did not form when 40% excess Mg was used. Wet ball milling of the products before leaching was found to be effective in removal of Mg3B2O6 during leaching in 1M HCl. When stoichiometric starting mixtures were used, all of the side products could be removed after wet ball milling in ethanol and leaching in 5 M HCl. Thus, pure TiB2 was obtained with a molar yield of 30%. Pure TiB2 could also be obtained at a molar yield of 45.6% by hot leaching of VCS products at 75oC in 5 M HCl, omitting the wet ball milling step. By mechanochemical processing, products containing only TiB2 and MgO were obtained after 15 hours of ball milling. Leaching in 0.5 M HCl for 3 minutes was found to be sufficient for elimination of MgO. Molar yield of TiB2 was 89.6%, much higher than that of TiB2 produced by volume combustion synthesis. According to scanning electron microscope analyses, produced TiB2 had average particle size of 0.27&plusmn / 0.08 &amp / #956 / m.

The aim of this study was to investigate synthesis of lithium borides by mechanochemical synthesis from oxides. Lithium borides have promising properties in the area of high energy additives and hydrogen storage. Lithium oxide (Li2O), boron oxide (B2O3) and Mg were used to synthesize lithium borides. Experiments were conducted in a planetary ball mill under argon atmosphere. Analyses of the products were done by X-ray diffraction and scanning electron microscopy. Trilithium tetradecaboride (Li3B14) peaks were observed in the product powder. Removal of other phases that were formed during experiments was done by leaching in HCl/water solution. Leaching in 0.5 M HCl/water solution for 10 minutes was found to be sufficient to remove / iron (Fe) and magnesium oxide (MgO). Effects of ball milling parameters such as milling speed, ball to powder ratio, milling duration were investigated and milling for 20 hours with 300 rpm and 30:1 ball to powder ratio was found to be the optimum conditions. Syntheses of other lithium borides (LiB4, Li2B6, LiB13) were also experimented with the same milling parameters. Formation of LiB4, Li2B6 and LiB13 was not observed in the product powders. However, the results of LiB4 and LiB13 production experiments showed also Li3B14 peaks in the product. Li2B6 synthesis experiments resulted in Li2B9 peaks in the product powders.

The aim of this thesis study was to extract nickel and cobalt from Sivrihisar limonitic nickel laterite ore by high pressure acid leaching (HPAL) method under most economical operating parameters. Optimizing the conditions to yield a saleable quality mixed hydroxide product from the pregnant leach solution (PLS) was also investigated. To extract high amounts of nickel and cobalt from the laterite matrix / leaching duration, leaching temperature and sulfuric acid/ore ratio were studied at fixed conditions of -850 &micro / limonitic ore particle size and 40% solids concentration. The Sivrihisar limonitic nickel laterite ore was found to be readily leachable. It was found that 95.4% of Ni and 91.5% of Co were extracted at the optimized conditions of 235oC, 0.23 acid/ore ratio in 60 minutes. The real pregnant leach solution produced at the optimized conditions of HPAL was purified in two iron removal stages under the determined optimum conditions. Nearly all of the Al and Cr were removed from the PLS in the two stages of iron removal. Then, nickel and cobalt were taken out from the PLS in the form of mixed hydroxide precipitates (MHP) in two stages. A MHP 1 product containing 33.41 wt.% Ni, 2.93 wt.% Co was obtained with a Mn contamination of 3.69 wt.% at the optimized conditions of pH=7, 50oC and 60 minutes. The MHP 1 product was also contaminated with Fe (2.83 wt.%) since it could not be completely removed from the PLS without the critical losses of nickel and cobalt during the two iron removal stages.

Philosophiae Doctor - PhD / The catalytic support materials used in the present study are zeolite HBEA and MCM-41. These high silica zeolites were synthesised from coal fly ash (CFA) waste via a novel approach that involved a fusion step, acid assisted silica extraction and removal of Al, Ca and Na from the silica by treatment with oxalic acid. The generated silica was converted to HBEA and MCM-41 via conventional hydrothermal treatment. The metal incorporation onto HBEA was done via two approaches namely; liquid phase ion exchange (LIE) and wet impregnation (WI) while the loading on MCM-41 was only done via WI since the material does not possess exchange sites. The metal solution precursors were AMD and Fe extracted from CFA (FeAsh) via acid leaching followed by pH regulation by concentrated NaOH. This is the first time these solutions were tested as possible metal precursors in catalyst synthesis. / 2021-08-30

A low-cost titanium production process, the CSIR-Ti powder process, which aims to produce titanium powder directly by metallothermic reduction of titanium tetrachloride with lithium, has been under development at the Council for Industrial and Scientific Research (CSIR). Crude titanium powder produced using the CSIR-Ti process is inevitably contaminated with by-products such as lithium chloride, lithium and titanium dichloride. These by-products tend to become sources of impurities in titanium powder, specifically oxygen and chloride impurities. The presence of oxygen and chloride impurities has marked effects on the mechanical properties of titanium finished products. Consequently, for the crude titanium powder to be rendered useful downstream, it must be purified and the by-products reduced to concentrations specified in the commercial standards. The present study was undertaken to examine whether acid leaching could be used to selectively dissolve and prevent hydrolysis of the by-products—specifically excess lithium and unreacted titanium dichloride in the crude titanium powder produced by the CSIR-Ti process. A further objective was to determine whether a purified product that meets both oxygen and total residual chloride content as specified by the standards can be achieved. The effects of key leaching variables and their interaction were also investigated to gain fundamental understanding of these effects on the by-products leaching behaviour. A literature study to select a suitable lixiviant and to establish the aqueous chemistry of the byproducts and their effect on the leaching conditions was undertaken. It showed that of the various acids suggested in the literature, hydrochloric acid was the cheapest and that it was more suited for the CSIR-Ti leaching process than nitric acid, due to the common ion chloride. This simplifies the leachate purification process downstream. The literature study established that Ti(II) has no aqueous chemistry but instead is oxidised to Ti(III) in solution. It was found that Ti(III) is easily oxidised to TiO2+ by dissolved oxygen and water. However, the oxidation rate was slow in hydrochloric acid solutions with the advantage that hydrolysis of the ions could be minimised and the precipitation of the oxides or oxychlorides prevented. It was further revealed that the lithium neutralisation reaction is highly exothermic, with the possibility of raising the leachate temperature to 60°C, resulting in the contamination of the titanium powder particles by the oxide layer and precipitated hydrolysis products. Batch leaching tests were carried out using factorial design of experiments to investigate the effect of initial hydrochloric acid concentration, which was estimated by varying the concentration between 0.032 M and 1 M; particle size, which was varied between −10 mm and +10 mm; and the initial temperature, varied between 14°C and 30°C. The resulting data were modelled and analysed using the analysis of variance statistical method. The solid residues were analysed for oxygen and total residual chloride content. The solid residue was also characterised by scanning electron microscopy (SEM) to examine the morphology of the leached particles. Leaching kinetics model fitting was also conducted. The statistical analysis showed that of the three factors investigated, temperature was the factor with the most statistical significance on both the oxygen and chloride concentration in the purified product, followed by particle size. The effect of acid concentration proved to be minimal, a phenomenon attributed to low concentrations of acid-consuming impurities, specifically excess lithium in the crude product. Thus, the two concentrations of hydrochloric acid investigated were found to be efficient to prevent hydrolysis product formation. Scanning electron micrographs revealed that crushing the crude product with a jaw crusher occluded crude titanium pores, thus locking in some by-products in addition to the pores locked by sintering during the metallothermic reduction. The observation showed that residual chloride impurities in the purified product are not just a consequence of hydrolysis products but also byproducts locked deeper in the pores of the product. Based on the parameter ranges evaluated in the study, a product that satisfied both oxygen and chloride standard specifications was achieved when the crude product was leached in both 1 M and 0.032 M initial HCl concentrations, temperature of 30°C and particle size of +10 mm. The combination of (−10 mm and 14°C) at all concentrations also yielded acceptable oxygen and chloride content levels. Overall, it was concluded from the present work that purification of crude CSIR-Ti product by leaching in dilute HCl is technically feasible. / Dissertation (MSc Applied Sciences (Metallurgy))--University of Pretoria, 2014. / Materials Science and Metallurgical Engineering / MSc (Applied Sciences (Metallurgy)) / Unrestricted

The Council for Scientific and Industrial Research (CSIR) in South Africa is developing a process to produce titanium tetrachloride from a low-grade material such as ilmenite. Titanium tetrachloride can then be used as feed material for titanium metal or pigment-grade titanium dioxide production. Titanium tetrachloride is commercially produced by chlorinating synthetic rutile (<92% TiO2) or titanium dioxide slag (<85% TiO2) at ~900 ˚C. A drawback of chlorination at this temperature is that any constituents other than TiO2 will end up as hazardous waste material. A characteristic step in the CSIR’s proposed process is to nitride titanium dioxide contained in the feed material before it is sent for chlorination. The chlorination of the resulting titanium nitride is achieved at a much lower temperature (~200 ˚C) than that of the existing titanium dioxide chlorination reaction. An added advantage of the low-temperature chlorination reaction is that chlorine is selective mostly towards titanium nitride and metallic iron, which means that any other constituents present are not likely to react with the chlorine. The result is reduced chlorine consumption and less hazardous waste produced. The nitrided ilmenite must, however, be upgraded by removing all iron before it can be sent for chlorination. Commercial ilmenite upgrading processes, called synthetic rutile production, also require the removal of iron and other transition metals before chlorination. A literature review of existing ilmenite upgrading processes revealed four possible process options that could remove iron from nitrided ilmenite. Two of these process options, the Becher and Austpac ERMS SR processes, are proven process routes. The other two are novel ideas – one to passivate iron contained in the nitrided ilmenite against chlorination and the other to use ammonium chloride (as used in the Becher process) as a stoichiometric reactant to produce a ferrous chloride solution. A preliminary experimental evaluation of these process options indicated that the Austpac ERMS SR process is the most viable option for removing iron from nitrided ilmenite. The Austpac ERMS SR process was therefore selected as a template for further process development. A detailed Austpac ERMS SR process review found that two process units in the Austpac ERMS SR process could be used in a process that separates iron from nitrided ilmenite. These are the Enhanced Acid Regeneration System and the Direct Reduced Iron process units. The review also concluded that another leach unit would have to be developed. It was therefore necessary to further investigate the dissolution of nitrided ilmenite in hydrochloric acid. A detailed experimental evaluation of nitrided ilmenite dissolution in hydrochloric acid found that hydrochloric acid could be used as the lixiviant to selectively remove iron from nitrided ilmenite. The dissolution of metallic iron in 90 ˚C hydrochloric acid reached levels of at least 96% after only 60 minutes. An average “combined resistance” rate law was found that could be used to describe this dissolution reaction. The observed activation energy and Arrhenius pre-exponential factor were found to be equal to 9.45 kJ.mol-1 and 30.8 s-1 respectively. The Austpac ERMS SR process review and experimental results described above were then combined and used to propose a process that could be employed to remove iron from nitrided ilmenite. The proposed process was modelled using the Flowsheet Simulation module in HSC Chemistry 7.0 / Dissertation (MEng (Chemical Engineering))--University of Pretoria, 2010. / Chemical Engineering / MEng (Chemical Engineering) / unrestricted

Many literatures have been published recently regarding the recovery of REEs from coal-related materials, such as coal waste, acid mine drainage, and coal combustion ash. The recovery of REEs from coal waste has been investigated by the author in recent years, and it was found that after calcination at 600 ℃ for 2 h, a significant improvement in REE recovery can be achieved. In order to reveal the mechanisms of the enhanced REE recovery after calcination, coal waste samples from two different seams, i.e., Western Kentucky No. 13 and Fire Clay, were selected to investigate the modes of occurrence of REEs. Scanning electron microscopy- and transmission electron microscopy-energy dispersive X-ray spectroscopy (SEM-EDS and TEM-EDS) analyses were conducted to investigate the mineralogy of REEs in two coal waste samples. Totally, 49 and 50 REE-bearing particles were found from the SEM specimens of Western Kentucky No. 13 and Fire Clay coal waste samples, respectively. Based on the elemental composition analyses and TEM-EDS characterization, it was found that apatite, monazite, and crandallite-group minerals were the major light REE (LREE) carriers, while the heavy REEs (HREE) primarily occurred in zircon and xenotime in these two coal waste samples. Further analyzing the REE content and number of REE-bearing particles, it was confirmed that monazite, xenotime, and crandallite-group minerals were the dominant contributors to the total REE (TREE) contents in both materials. In addition to the mineralogy of REEs, the morphology of REE-bearing particles was also investigated. The SEM images suggested that the particle size of most REE-bearing particles was less than 5 μm. Moreover, not only completely liberated particles, but particles encapsulated by the host minerals present in the two coal waste samples. To identify the changes of mineralogy of REEs after recovery, the leaching solid residues of the raw and calcined coal waste samples were also characterized by SEM-EDS analysis. After REE recovery, the same REE mineralogical results were observed from the leaching residues of the raw coal waste samples. However, as for the calcined samples, the crandallite-group minerals disappeared. These results suggested that the crandallite-group minerals were decomposed into easy-to-leach forms after calcination at 600 ℃, thus leading to the improved REE recovery. Moreover, the number of REE-bearing particles (N) found from per area of the calcined leaching residue was confirmed to be larger compared to that of the raw ones. A combination analysis of these results indicated that two mechanisms of the enhanced REE recovery after calcination can be confirmed: (1) decomposing the crandallite-group minerals into more soluble species; and (2) promoting the liberation of the REE-bearing particles encapsulated in the host minerals. The thermal decomposition of crandallite-group minerals was mainly responsible for the enhanced REE recovery from coal waste. However, as a result of the complex isomorphic substitutions and association characteristics, it is difficult to collect a pure endmember of crandallite-group mineral for characterization. Therefore, florencite-(Ce) was synthesized in this study. X-ray diffraction (XRD), SEM-EDS, TEM, thermogravimetric and differential thermal analyses (TGA-DTA), and acid leaching tests were conducted on the synthesized product. The results showed that the variation in Ce leaching recovery corresponded to the phase transformation of florencite. The gradual transformation of florencite from a crystalline mineral into an amorphous phase resulted in the increases in the solubility of Ce. In addition, the thermal transformation of florencite was an independent reaction, which was not interfered by the host materials, such as kaolinite and coal waste. / Doctor of Philosophy / Coal waste has been identified as a promising alternative source of rare earth elements (REEs). This study showed that calcination can significantly improve the REE recovery from coal waste materials. Scanning electron microscopy- and transmission electron microscopy-energy dispersive X-ray spectroscopy (SEM-EDS and TEM-EDS) analyses were conducted to investigate the mineralogy of REEs in two coal waste samples. The results indicated that the REEs mainly present as apatite, zircon, monazite, xenotime, and crandallite-group minerals in the coal waste samples. However, after REE recovery, the crandallite-group minerals disappeared from the calcined coal waste samples. Therefore, it can be confirmed that the calcination treatment resulted in the solubility improvement of crandallite-group minerals in coal waste samples. In order to further investigate the crandallite-group minerals, florencite was synthesized and subjected to a series of characterizations. The results suggested that the thermal phase transformation of florencite from crystalline into amorphous state resulted in the solubility improvement.

Summary Serpentinites from the northern Vermont were examined for the distribution and abundance of As. XRD and electron microprobe showed the samples are composed of antigorite, chromite, magnetite, and carbonate minerals (magnesite, dolomite, calcite). The concentration in As when the samples were dissolved in H3PO4 was 10% of the concentration in As when the samples were dissolved in concentrated HF/HNO3, suggesting that As is mainly incorporated in the structure of antigorite. X-ray absorption near-edge structure spectra showed that the As is As(III) in the samples. Extended X-ray absorption fine structure spectra suggested that the As has a tetrahedral coordination and is located in the Si-site in serpentine.