Analysis of Alkali Roasting and Leaching on the Extraction of Metallic Values from Titaniferous Minerals

A novel process for the recovery of iron, titanium dioxide and vanadium pentoxide from titanium-bearing waste materials has been developed. The methodology has been designed to address the environmental legacy of tens of million tonnes of waste from vanadium oxide extraction that has accumulated in South Africa and around the world. Simultaneous extraction of the metallic values from the waste materials is not achievable with the existing methods.
Recovery of vanadium oxide dissolved in the titaniferous magnetite (TM) matrix is achieved by two processes that generate two distinct waste materials. Alkali roasting- water leach process solely extracts V2O5 yielding a roast-calcine tailings (Iron-rich tailing, T1). A low-grade TiO2 slag (TiO2 rich, T2) is discarded during co-production of steel and vanadium slag from TM.
T1 and T2 were characterised using scanning electron microscopy (SEM), X-ray diffraction (XRD) and X-ray fluorescence (XRF) to understand the varied and complex mineralogy.
The roast-calcine tailings had a total iron content of 49.7 weight.%, 0.3 weight.% V2O5 and contained up to 11 weight.% of TiO2. Vanadium ions are disseminated in residual magnetite, sodium aluminosilicates and the hematite-ilmenite solid solution observed in the material. By comparison, the low grade TiO2 slag was composed of up to 35 weight.% TiO2 and 1.2 weight.% V2O5. Varied distribution of vanadium ions in refractory magnesium aluminate and magnesium dititanate and the presence of significant quantities of gangue materials makes the state-of-the art TiO2 and V2O5 extraction technologies unsuitable.
Thermodynamic considerations were adopted for analysing the role of alkali carbonate in the physico-chemical separation of the metallic values.
A summary flowsheet is presented below broadly outlining the process steps. Carbothermic reduction in the presence of Na2CO3 was carried out on the roast-calcine tailings (T1) for extracting metallic iron, sodium titanate and water-soluble sodium vanadate in the 1073 K to 1323 K temperature range under reducing atmosphere. Wet magnetic separation of the reduced sample separated the metallic iron into a magnetic fraction, whereas the non-magnetic fraction was dominated by sodium titanate and residual gangue minerals with the sodium vanadate dissolved in the wash water. Water leaching of the non-magnetic fraction was performed to enhance recovery of sodium vanadate and alkali carbonate.
The magnetic fraction was smelted at 1723 K to produce steel micro-alloyed with vanadium and a titania-slag. 0.5M H2SO4 leaching of the non-magnetic fraction was performed to upgrade the TiO2 content to ca. 75 weight.%. T1 acid leach residue was roasted with NaHSO4 at 923 K followed by water leaching to remove water-soluble sulphated impurities. A TiO2 residue of up to 92 weight.% was obtained after water leaching.
Although the TiO2 rich (T2) waste contained a limited amount of iron, up to 10 weight.%, the material was reduced in the presence of Na2CO3 at above 1223 K to decompose phases hosting vanadium ions and form water- soluble sodium vanadate. The reduced product was water leached to dissolve water-soluble compounds for recovery by selective precipitation. Alumina (ca. 75 weight.% Al2O3) was precipitated from the water leach solution after lowering the pH to 8. Addition of ammonium sulphate to the solution followed by solution pH adjustment to 5 resulted in ammonium metavanadate precipitation, which was calcined at 723 K to give vanadium pentoxide of 93 weight.%. The residual solids from water leaching were leached in 0.5M HCl to yield a CaTiO3-rich residue.
The calcium titanate residue was treated using two different methods:
The first involved roasting the calcium titanate residue with NaHSO4 followed by water leaching. As T2 material contained a limited amount of iron it was necessary to add FeSO4 during the water leaching step to ensure dissolution of CaSO4 in order to yield an 86 weight.% TiO2 residue.
The second method utilised sulphuric acid baking at 473 K of the titanate residue to form a porous cake consisting of water-soluble TiOSO4 and CaSO4. Ambient temperature water leaching of the porous cake was carried out to dissolve TiOSO4 and separate it from the CaSO4-rich solid residue. Titanium dioxide hydrate was precipitated from the TiOSO4 solution by hydrolysis at 353 K. The TiO2 precipitate was calcined to produce synthetic rutile with up to 97 weight.% TiO2.
The flowsheet developed for T2 treatment has formed the basis of a pilot-scale plant designed to scale-up the technology and valorise the waste material.