A large number of unexploited low-grade vanadium-titanium magnetite deposits have been found in the Chao-yang area of China in recent years. The reserves are estimated at more than 20 billion tons. A mineralogical study of raw sample indicated that it was a typical low-grade vanadium-titanium magnetite ore with weathering. Most of vanadium-titanium magnetite was replaced by martite and sphene, which would affect the recovery of valuable elements. The intergrowth relationship between vanadium-titanium magnetite and sphene was very complex, and the grain size of sphene was generally fine, and could not be completely liberated even by fine grinding, whereas the content of vanadium and titanium in sphene was higher than that in vanadium-titanium magnetite, resulting in vanadium-titanium magnetite concentrates with characteristics of high vanadium and titanium content. A magnetic separation process was investigated for the pre-concentration of low-grade vanadium-titanium magnetite ore. The results showed that 73.52% of feed ore were directly discarded for reducing the processing capacity of follow-up processing. A vanadium-titanium magnetite concentrate with 1.14% V₂O₅, 22.22% TiO₂, 42.51% Fe and a rough concentrate of ilmenite with 16.05% TiO₂, 20.77% Fe were obtained from low-grade vanadium-titanium magnetite ore which contains 0.050% V₂O₅, 1.67% TiO₂ and 8.53% Fe.

Vanadium-bearing muscovite is the most valuable component of stone coal, which is a unique source of vanadium manufacture in China. Numbers of experimental studies have been carried out to destroy the carrier muscovite’s structure for efficient extraction of vanadium. Hence, the vanadium location is necessary for exploring the essence of vanadium extraction. Although most infer that vanadium may substitute for trivalent aluminium (Al) as the isomorphism in muscovite for the similar atomic radius, there is not enough experimental evidence and theoretical supports to accurately locate the vanadium site in muscovite. In this study, the muscovite model and optimal location of vanadium were calculated by density functional theory (DFT). We find that the vanadium prefers to substitute for the hexa-coordinated aluminum of muscovite for less deformation and lower substitution energy. Furthermore, the local geometry and relative electronic properties were calculated in detail. The basal theoretical research of muscovite contained with vanadium are reported for the first time. It will make a further influence on the technology development of vanadium extraction from stone coal.

It is important to study the effect of Fe(III) on the positive electrolyte, in order to provide some practical guidance for the preparation and use of vanadium electrolyte. The effect of Fe(III) on the thermal stability and electrochemical behaviour of the positive electrolyte for the vanadium redox flow battery (VRFB) was investigated. When the Fe(III) concentration was above 0.0196 mol l ⁻¹ , the thermal stability of V(V) electrolyte was impaired, the diffusion coefficient of V(IV) species decreased from (2.06–3.33) × 10 ⁻⁶ cm ² s ⁻¹ to (1.78–2.88) × 10 ⁻⁶ cm ² s ⁻¹ , and the positive electrolyte exhibited a higher electrolyte resistance and a charge transfer resistance. Furthermore, Fe(III) could result in the side reaction and capacity fading, which would have a detrimental effect on battery application. With the increase of Fe(III), the collision probability of vanadium ions with Fe(III) and the competition with the redox reaction was aggravated, which would interfere with the electrode reaction, the diffusion of vanadium ions and the performance of VRFB. Therefore, this study provides some practical guidance that it is best to bring the impurity of Fe(III) below 0.0196 mol l ⁻¹ during the preparation and use of vanadium electrolyte.

In this paper, the composite additive of CaF₂/CaO was used to extract vanadium from stone coal, and the effect of roasting and leaching kinetics were studied. The purpose of this manuscript is to realize and improve the vanadium recovery from stone coal using the composite additive. The experimental results indicated that the roasted clinker can be obtained under the conditions of CaF₂/CaO at a mass ratio of 2:3 and a total additive amount of 10 wt %, a roasting temperature 850 °C, and a roasting time of 90 min. The leaching rate of vanadium can reach 86.74%, which increased by 16.4% compared with that of blank roasting under the conditions including a leaching temperature of 950 °C, a sulfuric acid concentration of 15% (v/v), a leaching time of 2 h, and a ratio of liquid to solid of 3 mL/g. The phase transformation analysis indicated that the muscovite structure was effectively destroyed during the roasting process comparing with no additives, which provided the basis for vanadium dissociation. Roasting can promote the formation of calcium vanadate, which is beneficial to the leaching of vanadium. The vanadium leaching kinetic analysis indicated that the activation energy of the acid leaching reaction decreased from 42.50 KJ/mol in the blank roasting to 22.56 KJ/mol in the calcified roasting, and the reaction order, with respect to the sulfuric acid concentration, decreased from 1.15 to 0.85. Calcified roasting has a better mineral activation than blank roasting, which can accelerate the leaching of vanadium and reduce the dependence on high-temperature and high acid levels in the leaching process.

In China, direct-reverse flotation is proved to be applicable to most phosphate ores. However, because the ratio of froth product is generally high, current direct-reverse technology faces challenges in terms of high reagent consumptions and cost. A new gravity and flotation combined process has been developed for the recovery of collophanite from sedimentary phosphate ore from the beneficiation plant of Hubei, China. In this process, 53% of the collophanite was firstly recovered by gravity separation, reducing the mass flow to direct flotation. The gravity tailing was the feed for the direct flotation. The flotation concentrate, mixed with gravity concentrate, was then subjected to reverse flotation. A final concentrate with a grade of 30.41% P₂O₅ at a recovery of 91.5% was produced from the feed analyzing 21.55% P₂O₅. Compared to the conventional direct-reverse flotation 86.1% recovery at 31.69% P₂O₅, it was found that pre-recovery of collophanite by spiral separation could significantly reduce the flotation reagent consumption and lead to improved overall collophanite recovery. The benefits of the new process in terms of cost savings were also discussed.

Phosphate ore is a valuable strategic resource. Most phosphate ore in China is collophane. Utilization of mid-low grade collophane is necessary to maintain social sustainable development. The gravity-flotation combination separation process can be utilized to separate mid-low grade collophane, but the process consumes a large quantity of acid in the reverse stage. Sodium dodecyl sulfate (SDS) was used as a dolomite collector in this study to reduce the acid consumption of collophane flotation. SDS effectively removed dolomite from the gravity concentrate when no other reagents were present. Flotation test results showed that, compared to the conventional gravity-flotation process, the proposed SDS-based process reduced phosphoric acid dosage from 6.1 kg/t to 3.9 kg/t with similar separation results. The SDS action mechanisms on dolomite were further investigated by zeta potential analysis, single mineral flotation tests, infrared spectrum detection, and theoretical analysis. The results indicate that the SDS adsorption on dolomite is mainly physical adsorption, and that favorable separation effects between collophane and dolomite may be attributed to physical adsorption and entrainment. In addition, it also indicates that the physical adsorption can be utilized to remove dolomite from phosphate on account of zeta potential differences when the separate feed is coarse.